Rotary Drilling and Blasting in Large Surface Mines
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Rotary Drilling and Blasting in Large Surface Mines
Bhalchandra V. Gokhale
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Cover Illustration:
Atlas Copco Pit Viper 351E drilling 311 mm holes to a depth of 19 m in the Aitik Mine, Sweden, owned by Boliden Mining Co. Courtesy of Atlas Copco, 2009
CRC Press/Balkema is an imprint of the Taylor & Francis Group, an informa business © 2011 Taylor & Francis Group, London, UK Typeset by Vikatan Publishing Solutions (P) Ltd, Chennai, India Printed and bound in Great Britain by Antony Rowe ( A CPI-group Company), Chippenham, Wiltshire All rights reserved. No part of this publication or the information contained herein may be reproduced, stored in a retrieval system, or transmitted in any form or by any means, electronic, mechanical, by photocopying, recording or otherwise, without prior permission in writing from the publisher. Innovations reported here may not be used without the approval of the authors. Although all care is taken to ensure integrity and the quality of this publication and the information herein, no responsibility is assumed by the publishers nor the author for any damage to the property or persons as a result of operation or use of this publication and/or the information contained herein. Published by: CRC Press/Balkema P.O. Box 447, 2300 AK Leiden, The Netherlands e-mail:
[email protected] www.crcpress.com – www.taylorandfrancis.co.uk – www.balkema.nl Library of Congress Cataloging-in-Publication Data Rotary drilling and blasting in large surface mines / editor, Bhalchandra V. Gokhale. p. cm. Includes bibliographical references and index. ISBN 978-0-415-87878-4 (hardback) 1. Blasting. 2. Rotary drilling. 3. Strip mining. I. Gokhale, B.V. (Bhalchandra V.) TN279.R68 2011 622’.292–dc22 2010045162 ISBN: 978-0-415-87878-4 (Hbk) ISBN: 978-0-203-84139-6 (eBook)
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Dedication
Dedicated to, My father, who gave me his genes, My mother, who brought me up and showered me with love till my 50s, Both my elder brothers, who supported me after the early demise of my father, My wife, who has been bearing with me for four decades, My daughter, son in law, son and daughter in law who have been making my life worth living, My grand daughter, who has kept me active.
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Contents
Foreword Preface Credits for illustrations 1 Surface rock excavation scenario 1.1 1.2
1.3 1.4 1.5
1.6
Introduction Need for surface rock excavation 1.2.1 Recovery of minerals or crushed rock 1.2.2 Creation of unobstructed space Strength of rock mass and classes Drill and blast operations Options to drill and blast 1.5.1 Water jetting 1.5.2 Machine digging 1.5.3 Ripping 1.5.4 Rock breaking Oilwell drilling
2 An overview of blasthole drilling 2.1 2.2 2.3 2.4 2.5
Introduction Basic concepts of drilling Peculiarities of blasthole drilling Comparison of drilling methods Choice of a blasthole drilling method
3 Properties of rocks 3.1 3.2 3.3
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Introduction Earth and its interior Geological cycle 3.3.1 Igneous rocks
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Contents
3.3.2 3.3.3
3.4
3.5 3.6 3.7
Sedimentary rocks Metamorphic rocks 3.3.3.1 Contact metamorphism 3.3.3.2 Regional metamorphism 3.3.3.3 Dynamic metamorphism 3.3.3.4 Impact metamorphism Formation of rock mass 3.4.1 Volcanic activities 3.4.1.1 Extrusive structures 3.4.1.2 Hypabyssal structures 3.4.2 Plutonic structures 3.4.3 Surface activities Soils, rocks and their profile Occurrence of minerals Rock specimen properties 3.7.1 Density 3.7.2 Compressive strength 3.7.3 Tensile strength 3.7.4 Hardness 3.7.4.1 Shore hardness 3.7.4.2 Vickers hardness 3.7.5 Toughness 3.7.6 Brittleness 3.7.7 Coefficient of internal friction 3.7.8 Swell factor 3.7.9 Abrasiveness 3.7.9.1 Cerchar abrasiveness test 3.7.9.2 Miniature drill test 3.7.9.3 Bit wear test
4 Brief history of rotary blasthole drilling 4.1 4.2 4.3 4.4 4.5
Introduction Era of shot hole drilling Rotary drilling Rotary blasthole drilling Truck mounted rotary blasthole drills
5 Rotary blasthole drilling bits 5.1 5.2
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Introduction Drag bits 5.2.1 Blade bits 5.2.2 Claw bits 5.2.3 Feed force and rotation of drag bits
36 37 38 38 38 38 40 40 40 40 40 43 43 44 46 47 47 50 52 53 54 54 54 55 57 57 57 58 60 61 61 61 63 71 80 83 83 83 83 85 85
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Contents
5.3
5.4
5.5
5.6 5.7
Tricone bits 5.3.1 Types and nomenclature 5.3.2 Geometrical design aspects 5.3.2.1 Fluid circulation 5.3.2.1.1 Air cooled tricone bits 5.3.2.1.2 Sealed bearing bits 5.3.2.2 Size and shapes of teeth 5.3.2.3 Cone size 5.3.2.4 Number of teeth 5.3.2.5 Cone orientation 5.3.2.6 Top connections 5.3.3 Metallurgy 5.3.3.1 Materials 5.3.3.2 Heat treatment 5.3.4 Summary of design features 5.3.5 Bit load and rotation of tricone bits 5.3.6 IADC code Dull bit analysis 5.4.1 Bit observation 5.4.1.1 Observation during drilling operation 5.4.1.2 Observation after bit withdrawal 5.4.1.3 Observation at final bit rejection Bit records 5.5.1 General data 5.5.2 Formation data 5.5.3 Blasthole drill and accessories details 5.5.4 Bit performance details 5.5.5 Bit discard details Bit failure analysis and remedy Discarding drill bits
6 Rotary blasthole drilling accessories 6.1 6.2
6.3
6.4
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Introduction Drill pipes 6.2.1 Integral drill pipes 6.2.2 Fabricated drill pipes 6.2.3 Choice of a drill pipe Stabilizers 6.3.1 Replaceable sleeve stabilizers 6.3.2 Welded blade stabilizers 6.3.3 Integral blade stabilizers 6.3.4 Roller stabilizers Crossover subs
ix
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Contents
6.5
6.6
6.7
Shock absorbers 6.5.1 External shock absorber 6.5.2 In the hole shock absorbers Miscellaneous accessories 6.6.1 Thread protectors 6.6.2 Lifting bails and hoisting plugs 6.6.3 Blasthole plugs 6.6.4 Rotary deck bushing 6.6.5 Drill stem wrench 6.6.6 Bit breaker 6.6.7 Blasthole inclinometer 6.6.8 Blasthole camera 6.6.9 Feed force measuring kit 6.6.10 Blasthole dewatering pump 6.6.11 Recovery tools 6.6.12 Blasthole sampler 6.6.13 Laser measuring instruments Miscellaneous safety items 6.7.1 Close fitting shirt and long pants 6.7.2 Safety glasses 6.7.3 Safety toed shoes 6.7.4 Ear plugs or ear muffs 6.7.5 Safety vest 6.7.6 Respirator 6.7.7 Rain coat 6.7.8 Face shield
7 Rotary blasthole drills 7.1 7.2 7.3
7.4 7.5
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Introduction Assemblies in rotary blasthole drills 7.2.1 Assembly placement Types of rotary blasthole drills 7.3.1 Special purpose rotary blasthole drills 7.3.2 Conventional rotary blasthole drills 7.3.2.1 Crawler mounted rotary blasthole drills 7.3.2.2 Carrier mounted rotary blasthole drills 7.3.2.3 Comparison of crawler vs. carrier Layouts of rotary blasthole drills Details of assemblies 7.5.1 Assemblies in the lower group 7.5.1.1 Carrier undercarriage 7.5.1.2 Crawler undercarriage
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Contents
7.5.2
7.5.3
7.5.4
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7.5.1.2.1 Slewing crawler undercarriage 7.5.1.2.2 Pivoted crawler undercarriage 7.5.1.2.3 Crawler track components Assemblies in the upper group 7.5.2.1 Main frame 7.5.2.2 Leveling jacks 7.5.2.3 Prime mover 7.5.2.4 Compressor 7.5.2.5 Hydraulic system 7.5.2.6 Bit lubricating system 7.5.2.7 Automatic lubrication system 7.5.2.8 Radiators and oil coolers 7.5.2.9 Machinery house 7.5.2.10 Dust control system 7.5.2.10.1 Wet dust control 7.5.2.10.2 Dry dust control 7.5.2.11 Operator’s cab 7.5.2.12 Driver’s cab 7.5.2.13 Fuel tanks 7.5.2.14 Transformer and cable reel Assemblies in the mast group 7.5.3.1 Mast 7.5.3.2 Mast raising cylinders and mast braces 7.5.3.3 Mast ladder 7.5.3.4 Pipe changer 7.5.3.4.1 Carousel pipe changers 7.5.3.4.2 Single pipe changers 7.5.3.5 Bit changer 7.5.3.6 Angle hole attachments 7.5.3.7 Rotary head 7.5.3.8 Feed mechanism 7.5.3.8.1 Double drum and wire rope 7.5.3.8.2 Hydraulic cylinder and wire rope 7.5.3.8.3 Hydraulic cylinder and chain 7.5.3.8.4 Fixed sprockets and chain 7.5.3.8.5 Rack and pinion with chain 7.5.3.8.6 Chainless rack and pinion 7.5.3.9 Auxiliary winch 7.5.3.10 Tool handling jib 7.5.3.11 Centralizer Special purpose items 7.5.4.1 Language name plate 7.5.4.2 Lighting
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Contents
7.6
7.7
7.5.4.3 Fast fuel fill system 7.5.4.4 Depth indicator 7.5.4.5 Fast retract system 7.5.4.6 Tow hook 7.5.4.7 Fire extinguishers 7.5.4.8 Automatic fire suppression system 7.5.4.9 Welding outlet 7.5.4.10 Automatic leveling system 7.5.4.11 Remote propel control 7.5.4.12 Air conditioners 7.5.4.13 Heaters 7.5.4.14 Hydraulic test station 7.5.4.15 Video camera system Extreme cold operation devices 7.6.1 Engine starting aid 7.6.1.1 Glow plugs 7.6.1.2 Manifold flame heater 7.6.2 Batteries 7.6.3 Double wall machinery house 7.6.4 Machinery house heater 7.6.5 Operator cab heater 7.6.6 Other enclosures and heaters 7.6.7 Protective coating on steel components Comparison of types of blasthole drills
8 Compressed air and air compressors 8.1 8.2
8.3
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Introduction Compressed air 8.2.1 Gas laws 8.2.1.1 Boyle’s law 8.2.1.2 Charles’ law 8.2.1.3 Guy Lussac’s law 8.2.1.4 Joule’s law 8.2.1.5 Poisson’s law 8.2.1.6 Amagat’s law 8.2.1.7 Avogadro’s law 8.2.1.8 General gas law 8.2.2 Power required for air compression Flow of compressed air 8.3.1 Compressed air flow in steel pipes 8.3.1.1 Simplified formulae 8.3.1.1.1 Harris formula 8.3.1.1.2 Engineering toolbox formula
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Contents
8.3.2 Compressed air flow in hose pipes 8.3.3 Compressed air flow in pipe fittings 8.3.4 Effect of ups and downs in the flow path 8.3.5 Effect of curvature in the flow path 8.3.6 Compressed air flow through nozzles 8.3.7 Leakage of compressed air 8.4 Compressors used on blasthole drills 8.4.1 Sliding vane rotary compressors 8.4.2 Rotary screw compressors 8.4.3 Discharge and pressure control 8.4.3.1 Load/no load control 8.4.3.2 Modulation control 8.4.3.3 Load/no load and modulation control 8.5 Measurements of compressed air 9 Mechanics of rock fracture under a drill bit 9.1 9.2
9.3
9.4
9.5 9.6
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Introduction Rock fracture in drilling 9.2.1 Modes of fragmentation 9.2.2.1 Mechanical energy 9.2.2.1.1 Crushing 9.2.2.1.2 Impact crushing 9.2.2.1.3 Scratching 9.2.2.2 Heat energy 9.2.2.2.1 Spalling 9.2.2.2.2 Fusion and vaporization 9.2.2.3 Chemical energy Basic theory of solid fracture 9.3.1 Mohr-Coulomb theory of shear failure 9.3.2 Griffith theory of tensile failure Fracture of rock in drilling 9.4.1 Penetration of an indenter 9.4.2 Penetration of a wedge 9.4.3 Formation fracture below a drill bit 9.4.4 Specific energy Drillability of rocks Estimation of penetration rate 9.6.1 Protodyakonov approach 9.6.2 Paone and Bruce approach 9.6.3 Bauer, Calder and Workman approach 9.6.4 Specific energy approach 9.6.5 Microbit test approach 9.6.6 Indenter test approach
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Contents
10 Flushing the blasthole 10.1 10.2 10.3
10.4
10.5 10.6
10.7 10.8 10.9
Introduction Why compressed air? Schematics of flushing 10.3.1 Gravitational force 10.3.2 Buoyant force 10.3.3 Drag force 10.3.4 Terminal velocity 10.3.5 Laws for settlement of particles in fluid Formulation of desired bailing velocity 10.4.1 Size of fragments 10.4.2 Density of fragments 10.4.3 Roundness of fragments 10.4.4 Roughness of fragments 10.4.5 Rate of fragmentation 10.4.6 Annular space 10.4.7 Inclination of the blasthole 10.4.8 Quantity of water injection 10.4.9 Other approaches Formulation of compressed air pressure 10.5.1 Oil injection Choosing a drill pipe 10.6.1 Drill pipe dimensional parameters 10.6.2 Drill pipe surface treatment 10.6.3 Size and shape of drill cuttings 10.6.4 Bailing velocity 10.6.5 Drill pipe wall thickness Discarding a drill pipe Choosing nozzles for tricone bits Choosing the right compressor
11 Effect of altitude and severe weather 11.1 11.2
11.3
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Introduction Atmosphere 11.2.1 Effects of altitude on atmosphere 11.2.2 Effects of humidity on atmosphere Effect of scarce air in drilling 11.3.1 Diesel engines 11.3.2 Electric motors and transformers 11.3.3 Hydraulic system 11.3.4 Air compressors 11.3.5 Blasthole flushing 11.3.6 Drill lubrication
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Contents
11.4
Effect of severe weather 11.4.1 Stormy winds 11.4.2 Heavy rainfall 11.4.3 Heavy snowfall 11.4.4 Extreme hot weather 11.4.5 Extreme cold weather
12 Computers in rotary blasthole drilling 12.1 12.2
12.3 12.4
12.5
12.6
12.7
12.8
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Introduction Drill operation 12.2.1 Tramming 12.2.2 Leveling 12.2.3 Drilling and pipe handling Modes of blasthole drill computerization Drilling knowledge systems 12.4.1 Formation logging units 12.4.2 Analog drilling recorders 12.4.3 Electronic drilling recorders 12.4.4 Parameters of drilling knowledge data Positioning by GPS 12.5.1 Location recognition by GPS 12.5.2 Movement of drill by using GPS Computerized drill systems 12.6.1 Measurements while drilling 12.6.1.1 Independent parameters 12.6.1.1.1 Time 12.6.1.1.2 Depth 12.6.1.1.3 Rotary speed 12.6.1.1.4 Weight on the bit 12.6.1.1.5 Torque 12.6.1.1.6 Vibrations 12.6.1.1.7 Compressed air pressure 12.6.1.2 Calculated parameters 12.6.1.2.1 Rate of penetration 12.6.1.2.2 Blastability index 12.6.2 Automation systems Hardware for drill computerization 12.7.1 System unit 12.7.2 Display monitor 12.7.3 Input device 12.7.4 Software Advantages of drill computerization 12.8.1 Advantages of drilling knowledge system
xv
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Contents
12.8.2 12.8.3 12.8.4 12.8.5
Advantages of drill automation system Advantages of GPS positioning system Advantages of integrated mining system Advantage of added safety 12.8.5.1 Blasthole depth limiting input 12.8.5.2 Tram interlock for pipe in hole 12.8.5.3 Centralizer damage prevention 12.8.5.4 Leveling jack interlock 12.8.5.5 Carousel damage prevention 12.8.5.6 Breakout wrench damage prevention
13 Concepts for rotary blasthole drill design 13.1 13.2 13.3 13.4 13.5 13.6 13.7 13.8 13.9 13.10 13.11
375
Introduction Primary requirements of the drill Size and weight of the drill Rotary speed and torque Feed force and speed Pullout force and speed Breakout torque Hydraulic leveling jacks Ground load bearing Compressor discharge and pressure Engine power
375 375 376 378 383 384 384 385 385 386 386
14 Cost analysis of rotary blasthole drilling
389
14.1 14.2 14.3 14.4
14.5
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369 369 371 372 372 372 372 373 373 373
Introduction Working pattern of a blasthole drill Factors related to blasthole drilling cost Owning cost 14.4.1 Purchase expenditure 14.4.2 Yearly taxes, duties and levies 14.4.3 Salvage value 14.4.4 Calculation of owning cost 14.4.4.1 Single installment 14.4.4.2 Multiple installments Operating cost 14.5.1 Cost of maintenance and repairs 14.5.2 Cost of consumables 14.5.2.1 Cost of power 14.5.2.2 Drill lubricants 14.5.2.3 Oils 14.5.2.4 Water
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Contents
14.6 14.7 14.8 14.9
14.5.2.5 Tires 14.5.2.6 Power cable 14.5.3 Cost of accessories and bits 14.5.3.1 Drilling accessories 14.5.3.2 Tricone bits 14.5.4 Cost of operating labor Overhead cost Total blasthole drilling cost General considerations in cost estimation Example of cost estimation
15 Procurement of rotary blasthole drills 15.1 15.2
15.3
Introduction Pre procurement considerations 15.2.1 Which drilling method? 15.2.2 What type of drill? 15.2.3 How many blasthole drills? Procurement methodology 15.3.1 Draft of specification for the drill 15.3.2 Erection and commissioning of the drill 15.3.3 Drill manuals 15.3.4 Operation and maintenance training
16 Tips for operating rotary blasthole drills 16.1 16.2 16.3 16.4 16.5 16.6 16.7 16.8 16.9 16.10 16.11
Introduction Knowing the environment Moving the rotary blasthole drill Setting up the rotary blasthole drill Operating a rotary blasthole drill Controlling dust emissions Transporting the drill Maintenance activities Periodical check ups Drilling records Safety precautions
17 Mechanics of blast fragmentation 17.1 17.2 17.3 17.4
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Introduction Explosion process Formation of a crater by a blast Blast fragmentation mechanism
xvii
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Contents
17.4.1 17.4.2 17.4.3 17.4.4 17.4.5 17.4.6 17.4.7 17.4.8
Crushing of rock Radial cracking Circumferential cracking Spalling Radial push by gases Cracking by flexure Collision of fragments Separation of beds
18 Properties of explosives 18.1 18.2 18.3
Introduction Explosives Properties of explosives 18.3.1 Velocity of detonation 18.3.2 Detonation pressure 18.3.3 Blasthole pressure 18.3.4 Strength 18.3.4.1 Weight and bulk strength 18.3.4.2 Absolute and relative strength 18.3.5 Heat of explosion 18.3.6 Specific gas volume 18.3.7 Sensitivity 18.3.7.1 Sensitivity to shocks 18.3.7.2 Sensitivity to friction 18.3.7.3 Sensitivity to heat 18.3.7.4 Sensitivity to detonator strength 18.3.7.5 Gap sensitivity 18.3.8 Handling and transport safety 18.3.9 Brisance value 18.3.10 Charging density 18.3.11 Toxic fumes 18.3.12 Water resistance 18.3.13 Hygroscopicity 18.3.14 Storage life 18.3.15 Volatility 18.3.16 Material coexistence 18.3.17 Minimum hole diameter
19 Thermochemistry of explosives 19.1 19.2 19.3
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Introduction Chemical nature of explosives Reactions of explosive chemicals
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Contents
19.3.1 19.3.2 19.3.3 19.3.4
Oxygen balance Volume of products of explosion Heat of explosion Strength of explosive
20 Explosives – their history and composition 20.1 20.2 20.3
20.4
Introduction Brief history of explosives Types of explosives in a blasthole 20.3.1 Detonator (blasting cap) 20.3.2 Primary explosive 20.3.3 Booster 20.3.4 Main explosive 20.3.4.1 Nitroglycerin 20.3.4.2 Nitrocellulose 20.3.4.3 Ammonium nitrate 20.3.4.4 Sodium nitrate 20.3.4.5 Fuel oil 20.3.4.6 Wood pulp 20.3.4.7 Sulfur 20.3.4.8 Antacid Explosive mixes used in mine blasts 20.4.1 Dry blasting agents 20.4.1.1 Poured ANFO 20.4.1.2 Packaged ANFO 20.4.1.3 Heavy ANFO 20.4.2 Slurry 20.4.2.1 Bulk slurry 20.4.2.2 Cartridged slurry 20.4.3 Emulsions 20.4.4 Dynamite 20.4.4.1 Granular dynamite 20.4.4.2 Gelatin dynamite
21 Tests on explosives 21.1 21.2
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Introduction Tests for measurements of VOD 21.2.1 D’Autriche method 21.2.2 Chronograph method 21.2.3 Fiber optic sensor method 21.2.4 SLIFER method 21.2.5 Optical measurement method
xix
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Contents
21.3
21.4
21.5
21.6
Tests for measurement of strength 21.3.1 Ballistic mortar test 21.3.2 Cratering test 21.3.3 Underwater test 21.3.4 Plate dent test 21.3.5 Traulz lead block test 21.3.6 Cylinder compression test Tests for measurement of sensitivity 21.4.1 Shock sensitivity test 21.4.2 Heat resistance test 21.4.3 Cap sensitivity test 21.4.4 Gap sensitivity test 21.4.5 Friction sensitivity test 21.4.6 Electrostatic discharge sensitivity test Tests related to storage of explosive 21.5.1 Effect of wetness test 21.5.2 Internal ignition test 21.5.3 Material compatibility test 21.5.4 Vacuum stability test Miscellaneous tests 21.6.1 Critical diameter test 21.6.2 Critical height test 21.6.3 Bullet impact test 21.6.4 Koenen test
22 Blasting consumable and accessories 22.1 22.2
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Introduction Blasting consumables 22.2.1 Initiation transmission line 22.2.1.1 Safety fuse for transmitting fire 22.2.1.2 Electric transmission wires 22.2.1.3 Detonating cord for transmitting detonation 22.2.1.4 Detonating cord for transmitting shock 22.2.1.5 Hercudet tube for transmitting fire 22.2.2 Detonator 22.2.2.1 Delay element 22.2.2.2 Electric detonators 22.2.2.2.1 Solid pack electric detonator 22.2.2.2.2 Fusehead electric detonator 22.2.2.2.3 Exploding bridgewire detonator 22.2.2.3 Non-electric detonators 22.2.2.4 Electronic detonators
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Contents
22.3
22.4
Blasting instruments 22.3.1 Testing instruments 22.3.1.1 Blaster’s multimeter 22.3.1.2 Blaster’s ohmmeter 22.3.1.3 Blaster’s tagger 22.3.2 Initiating instruments 22.3.2.1 Safety fuse initiator 22.3.2.2 Electric detonation initiator 22.3.2.3 Generator type blast initiator 22.3.2.4 Capacitor discharge type blast initiator 22.3.2.5 Detonation wave initiator 22.3.2.6 Shock wave initiator 22.3.2.7 Hercudet system initiator 22.3.2.8 Electronic blast initiator 22.3.3 Measuring instruments 22.3.3.1 Hot hole meter 22.3.3.2 Burden measuring instrument 22.3.3.3 Seismograph 22.3.3.4 VOD meter 22.3.3.5 High speed camera Blasting tools and miscellaneous items
23 Hazards to and from blasting 23.1 23.2
23.3
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Introduction Hazards to blasting process 23.2.1 Presence of ground structure 23.2.2 Seepage of ground water 23.2.3 Hot blastholes 23.2.4 Hole deviation Hazards of blasting process 23.3.1 Misfires 23.3.1.1 Faulty safety fuse installation 23.3.1.2 Faulty electric blasting circuits 23.3.1.3 Faulty detonating cord circuits 23.3.1.4 Faults of exploder or faulty operation 23.3.1.5 Unnoticed ground water inflow 23.3.2 Ground vibrations 23.3.2.1 Nature of ground vibration waves 23.3.2.2 Prediction of ground vibration levels 23.3.2.2.1 Langefors formula 23.3.2.2.2 Scaled distance formula 23.3.2.3 Damage by ground vibrations
xxi
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Contents
23.3.3
23.3.4 23.3.5 23.3.6
Air blast 23.3.3.1 Prediction of air blast pressure 23.3.3.2 Damage by air blast pressure Fly rocks Air pollution Environmental changes
24 Properties of rock masses 24.1 24.2
24.3
24.4
Introduction Rock specimen properties and blasting 24.2.1 Influence of rock strength on blasting 24.2.2 Influence of rock density on blasting 24.2.3 Influence of rock porosity on blasting 24.2.4 Specimen blastability Properties of rock masses and blasting 24.3.1 Voids 24.3.2 Folds, unconformities and bedding planes 24.3.2.1 Shooting with the dip 24.3.2.2 Shooting against the dip 24.3.2.3 Shooting along the strike 24.3.3 Faults and joints Classification of rock masses 24.4.1 Classification by visual observation 24.4.2 Classification by index 24.4.2.1 RQD based classification 24.4.2.2 RMR based classification 24.4.2.3 RTQI based classification 24.4.3 BI index based classification 24.4.3.1 Index proposed by Hansen (1968) 24.4.3.2 Index proposed by Hainen and Dimock (1976) 24.4.3.3 Index proposed by Ashby (1977) 24.4.3.4 Index proposed by Langefors (1978) 24.4.3.5 Index proposed by Lilly (1986) 24.4.3.6 Index proposed by Ghose (1988) 24.4.3.7 Index proposed by Gupta (1990) 24.4.3.8 Index proposed by JKMRC (1996) 24.4.3.9 Index proposed by Han, Weiya and Shouvi (2000)
25 Methods and patterns of charging blastholes 25.1 25.2
Book.indb xxii
Introduction Mechanized blasthole charging system
545 546 547 549 550 552 553 553 553 554 554 555 555 555 556 556 558 558 559 560 560 560 560 562 563 564 564 565 566 566 566 566 568 568 569 569 571 571 571
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Contents
25.3
25.4
25.2.1 Trucks for loading ANFO or heavy ANFO 25.2.2 Trucks for loading slurry or emulsion 25.2.3 Safety features of bulk delivery system Blasthole charging pattern 25.3.1 Type and placement of primer 25.3.2 Direction of propagation of detonation 25.3.2.1 Top priming 25.3.2.2 Bottom priming 25.3.2.3 Multi point priming 25.3.3 Continuous side initiation 25.3.4 Air decking 25.3.5 Priming under special rock mass conditions 25.3.5.1 Well defined ore layer 25.3.5.2 Heavy water seepage in blasthole 25.3.5.3 Hard boulder in soft bed 25.3.5.4 Cavities in the rock mass Drilling and firing patterns 25.4.1 Drilling patterns 25.4.2 Firing patterns 25.4.2.1 “V” pattern 25.4.2.2 Echelon delay pattern 25.4.2.3 Flat face pattern 25.4.2.4 Channel delay pattern 25.4.2.5 Sinking hole pattern
26 Design of a surface blast 26.1 26.2 26.3
26.4
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Introduction Blastholes in a mine bench Types of blasting in surface mines 26.3.1 Conventional bench blasting 26.3.2 Secondary blasting 26.3.3 Cast blasting 26.3.4 Presplitting 26.3.5 Snake hole blasting 26.3.6 Rip rap blasting What is involved in design of a blast 26.4.1 Powder factor 26.4.2 Blasting direction 26.4.3 Number of free faces 26.4.4 Blasthole diameter 26.4.4.1 Desired rate of production 26.4.4.2 Desired fragmentation
xxiii
572 573 574 575 576 580 581 582 582 583 583 586 586 586 587 587 588 589 592 592 593 593 593 595 597 597 597 598 598 599 599 599 601 603 603 603 604 604 605 605 606
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Contents
26.5 26.6 26.7 26.8
26.4.4.3 Properties of rock specimen 26.4.4.4 Presence of geological structures 26.4.5 Blasthole depth 26.4.5.1 Combination of shovel and dumper 26.4.5.2 Walking dragline 26.4.5.3 Combination of wheel loader and dumper 26.4.6 Blasthole inclination 26.4.7 Burden 26.4.8 Blasthole spacing 26.4.9 Subdrilling 26.4.10 Stemming height 26.4.11 Size of the blast Calculation of burden Relationship of blasthole parameters Design of conventional surface blast Design of other types of blasts 26.8.1 Rip rap production blasting 26.8.2 Snake hole blasting 26.8.3 Cast blasting
Appendix 1 Appendix 2 Appendix 3 Appendix 4 Appendix 5 Appendix 6 Appendix 7 Appendix 8 Appendix 9 Appendix 10 Appendix 11 Appendix 12 Appendix 13 Appendix 14 Appendix 15 Index
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Properties of atmospheric air at high altitudes Air properties at various temperatures Bailing velocities in blasthole annulus Air pressure loss in steel pipes Air pressure loss in hose pipes Air pressure loss in pipe fittings A note on tungsten carbide and other hard metals Scales of hardness – measurements and their conversion Manufacturers of rotary blasthole drilling equipment Details of rotary blasthole drills manufactured all over the world Details of drill pipes used for rotary blasthole drilling Details of stabilizers and other miscellaneous items Details of shock absorbers Properties of some rocks Bibliography
606 606 609 609 610 611 612 616 617 618 620 620 621 625 627 630 630 630 632 639 643 647 651 657 661 667 675 679 687 693 699 707 709 717 737
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Foreword
I joined Bucyrus-Erie in 1974 after finishing my graduate studies. BE assigned me working on the design and test of rotary blast hole drills in the U.S., Canada and South Africa. In 1976 we had offered BE 45R (and also 40R which was soon to be discontinued) rotary blasthole drill to Malanjkhand Copper Project in India. To follow up this potential business I, as a young Sales Manager of BE, visited the mining site near Balaghat, with another young sales executive Mr. B.V. Gokhale from our Indian distributor. I vividly remember, Mr. Taneja, the director of the project, stating in our last meeting, “Mr. Johnson, we can’t purchase 45R because it is too large and too expensive for our job and your 40R does not have angle hole drilling attachment which is so important for our job. However, I must say that you and Mr. Gokhale have certainly taught us many aspects of rotary blasthole drilling in our discussions. Many thanks for that.” BV and I traveled throughout India to sites operated by Hindustan Copper, KIOCL, and Coal India at a later date. We had discussion on different types of equipment. I found BV not only had a superior knowledge of blast hole drills but also rigid frame haul trucks, dragline, electric mining shovels etc. When BE were considering starting manufacture of their shovels in India, we found BV as the most suitable individual to work as consultant to BE. Based on some of his work we were able to form a joint venture for electric mining shovels with Heavy Engineering Corporation and had signed a MOU for blast hole drill manufacturing. BV then conducted an in depth study of the Indian Coal sectors and equipment demands for the next 20 years. His study involved visit to all sites in the country for coal mining and accessing the precise long term demand for coal for power generation. His study convinced BE to establish a subsidiary operation in India. I was convinced beyond doubt that knowledge of BV in the domain of blasthole drilling was of highest caliber. Sometime during 1984 he mentioned to me that he is going to write a book on blasthole drilling. I knew he had all the necessary talent and knowledge for the venture. I wished him best luck and was really eager to see the book because there was no book on the subject of blasthole drilling. The nearest to the subject was “Drilling of Rocks” by McGregor but was grossly outdated within a decade after its publication in 1967.
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xxvi
Foreword
I was naturally pleased to see his book “Blasthole Drilling Technology” published in 2003. It does contain lot of useful material on the subject. But as the book was published by him, without backing of an international publisher, it has remained somewhat unknown. Now he has come up with this “Rotary Drilling and Blasting in Large Surface Mines”. I have read a few chapters of the book and I find that BV has put many innovative ideas on various topics. I can whole heartedly recommend this book to everyone directly or indirectly involved in blasthole drilling. It covers every aspect of rotary blasthole drilling. To a mining engineering student it gives complete knowledge of the machine and methods of its use without spending hours of field work. For a driller it will give deep insight into the various factors that affect the drilling process and several associated topics. To a mining superintendent it shows how to choose machine drill pipes and other high value consumables. For him the elaboration of cost calculations is of great importance. To a design engineer with a drill manufacturer it reveals several concepts that form the basis of the actual drilling process, and guiding principles for the design of a rotary drill. In short, by reading this book everyone involved with rotary blasthole drilling, will get a deep insight in the subject and will become acquainted with many ideas that can be used to get better results in drilling operations. Later in the book BV has also covered blasting techniques. Whereas I don’t have adequate knowledge to comment on the blasting part, after quickly reading it, I certainly feel that he has described the explosives, the process of rock fracture in blasting, thermochemistry of explosive and the design of a blast for surface mining operations in most appropriate manner. With a worldwide sales network CRC Press/Balkema are well placed to spread the knowledge of relatively untouched subject of rotary blasthole drilling and blasting through this book. I wish grand success to both BV and CRC Press/Balkema. Dean Johnson, Ex VP Terex Corporation. Currently Capital Goods Consultant for Global Sales and Distribution of Open Pit Mining and Construction Equipment.
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Preface
Nearly seven years ago I wrote and published my book titled ‘Blasthole Drilling Technology’. Many readers of the book, found it very useful because it contained more detailed information about rotary and rotary percussion drilling than almost all other books. Some of the readers, however, did express that I should have written two separate books – one on rotary blasthole drilling and the other on rotary percussion blasthole drilling. When I thought over this suggestion, I did realize that these two modes of blasthole drilling are vastly different from each other in terms of equipment and techniques. The only commonality is that both can be effectively used for the blasthole drilling application. Even for such application rotary percussion drilling is best suited for drilling small diameter blastholes, whereas, rotary drilling is best suited for drilling larger blastholes. Therefore, initially I decided to write a book only on rotary blasthole drilling. As I thought further over the matter I realized that in ‘Blasthole Drilling Technology’ blasting was covered only through one chapter and hence gave only rudimentary knowledge of it to the readers. In other words the book was subject-oriented and not application-oriented. In reality blasthole drilling and blasting are so inseparable that every book on blasthole drilling must cover blasting to a considerable extent. Considering all these aspects I have written this book. The book is named ‘Rotary Drilling and Blasting in Large Surface Mines’ because rotary blasthole drilling is almost exclusively carried out in large surface mines. The book also covers blasting to sufficient depth. Quite a bit of text matter as well as illustrations in this book have been repeated from ‘Blasthole Drilling Technology’ but there have been many additions to the text matter and illustrations. This surely makes it worthwhile to purchase this book even if one has a copy of ‘Blasthole Drilling Technology’. This book is surely not a work of art but a book pertaining to specific techniques. The purpose of the book is to make the readers understand the techniques of drilling large diameter blastholes by rotary drilling method, subsequently filling them with appropriate explosives in apt manner and finally blasting them to get desired fragmentation of the rock mass. When it comes to understanding, a picture is always worth thousands of words. Keeping this saying my in mind I have included nearly 400 illustrations in this book. Almost 60% of these illustrations are based on the images that I have retrieved from internet websites. As it was necessary to change these images to suit the context.
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xxviii
Preface
I have modified almost all the images. I have endeavored my best efforts in obtaining permission for us of the original image. In this regard I must honestly confess that without use of such images this book would not have been complete. On pages immediately following the table of contents of this book in a section titled ‘Credits for Illustrations’ I have given due credit to the organizations or individuals for initial use of the mother image. In case some name has been missed please consider such errors to be inadvertent. Readers are sincerely requested to bring all such error to my notice through a letter, so appropriate action will be taken at the time of preparing the next edition of this book. During the early years of my interest many manufacturers and suppliers fed me with their product and technical literature to turn me towards the field of drilling. I am thankful to them. Some names that I vividly remember in this regard are, Tamrock, Atlas Copco, Ingersoll Rand, Joy Manufacturing Co., Gardner Denver. I worked in Greaves Cotton Co. Ltd., Voltas Ltd., Bucyrus Erie Co. and Harnischfeger Corporation. As a product manager my job was to understand technological aspects of different types of construction and mining equipment and answer technical queries of customers. Without the generous attitude of organizations mentioned in earlier paragraphs this book would not have seen the light of day. From late 1990s Internet became the source of my information. It is ever expanding. There are many individuals, without whom this book would not have been possible. In this regard the first name that comes to my mind is Ms. Germaine Seijger, Senior Acquisition Editor of CRC Press/Balkema. Apart from her there are many individuals from CRC Press/Balkema who have taken lot of efforts in greatly elevating the standard of this book. Next in the line is Mr. Dean Johnson, Ex VP Terex Corporation. Right from the days in 1977 when our friendship began, till now he has been sending me very valuable articles, books etc. on drilling so as to be well informed in the subject. Perhaps more important is the fact that he has always kept my enthusiasm burning, I have personally not met him since 1986 but on his own he has gifted me two most recently published books on blasthole drilling. I cannot forget the names of Prof. A.K. Ghose, Former Director of ISM Dhanbad, Dr. A.K. Chakraborty and Mr. G.K. Pradhan, Joint Director of PCRA. After reading my book ‘Blasthole Drilling Technology’ these respected persons were largely instrumental in spontaneously awarding me with a Lifetime Achievement Award for Outstanding Contribution in Drilling Engineering through Indian Mining Journal at RockEx 2005 held in Nagpur. This was a great moral booster for me. It kept my interest in the field alive to culminate in this book. A technical book is always backed by technical data. While preparing the table of appendix 10 many persons gave the needed information promptly. I am thankful to them. Ms. Amanda Finn – Measurement Devices Ltd., Mr. Brian Fox – Atlas Copco Drilling Solutions, Mr. K.S. Ramchandran – Revathi Equipment Ltd., Mr. Tharmarajah Ramchandran – Sandvik Mining and Construction, Mr. Greg Scott – Bucyrus International Inc., Mr. Fred Slack – Schramm Inc.
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Preface
xxix
I believe, that every technical book written by any author must be read by at least one external editor who can make the book easily “digestible” to the readers. For my book this was very ably done by Dr. James Rainbird. He not only polished my language but with his superb memory and “hawk” eye went on to point out a few anomalies as distantly placed as chapters 3 and 22. I am greatly indebted to him. Finally, I thank my wife Suneeti, son Amit, daughter in law Anjali for smilingly bearing the neglect shown towards them while writing the book. B.V. Gokhale Mumbai, August 8, 2010
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Credits for illustrations
Literally hundreds of illustrations have been included in this book so as make the contents easily understandable to the readers. Nearly 60% of the illustrations are based on images gathered by me from various sources viz. books, periodicals and particularly internet websites of many manufacturers and suppliers. I have made changes in the images to suit the contents in the book. Thanks are due to all these individuals or organizations for such indirect contribution. Following is a list of individuals or organizations who originally have published the mother images.
Name of individual or organization
Figures for which credit given
AB Nordbok Acker Drill Co. Allis Chalmers American Coldset Corporation Aquila Mining System Atlas Copco
4.1 5.1 8.11-8.14 6.8 12.5, 12.6, 12.9, 12.10, 12.11, 12.13 2.2-2.5, 2.7, 4.10, 5.5, 5.6, 5.8, 6.5, 6.9, 7.1, 7.4, 7.6, 7.28, 7.31, 7.34-7.36, 7.41, 7.43, 7.57, 7.58, 8.5, 10.4, 10.5, A12.1, Table 5.13 7.23 4.12, 4.13 22.11, 22.20 22.20 5.2, 5.3 2.8, 4.8, 4.18, 4.19, 7.2, 7.8, 7.9, 7.22, 7.25, 7.38-7.40, 7.42, 7.50, 7.59, 6.2, 6.4, 8.3, 13.1, 26.3, 26.7 1.9, 7.10, 7.15 3.10 7.51 12.3, 26.14 26.27-26.30 4.14, 4.16 2.6, 6.1, 6.6 6.14, 6.15 22.7, 22.12, 22.14, 22.15 3.7
Autolub Lubin Baroid Inc. Blasting Accessories Inc. Blasters Tool and Supply Co. Inc. Blue Demon Bucyrus International Inc. Caterpillar Christine Rasch+Dr. Wolfgang Kollenberg Christopher Bills CR Books D’Appolonia Dave Lang Drillco Smithtool Duraquest Dyno Nobel ELE International
(Continued )
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xxxii
Credits for illustrations
Name of individual or organization
Figures for which credit given
Elvex Corporation Explosiva Fluid Control Products Inc. Geolograph GeoSonic GeoVision Gulf Publishing Hamlyn Publications implosionworld.com International Drilling Services J. Trent Bruce Joy Manufacturing Co. Keith Haddock Kennmetal Laser Measuring Instruments Inc Laval Measurement Surveys LIM SA Lincoln M&P Flange and Pipe Protection Mammoth Technologies Measurement Devices Ltd. Mining Technology International Modular Mining System Nordex Explosives Ltd NTNU – Norwegian Univ. of Sci. and Tech. Oil Tool International Services Pvt Ltd Olympus Orica P&H Mining Prof Hustrulid Pulsar Measuring Co. R. Smith and Sons Inc. Research Energy of Ohio Inc. Sandvik Schlumberger Schramm Inc. SDS Corporation Security BDS Smith Drilling Co. SA Smithtool Steve (Hanna Equipment Co.) Thomas Instrument Inc. ThyssenKrupp Fordertechnik TLC Engineering Solutions + Vibronics Transtronics Trumph Drilling Tools Twin Disk US Army US Dept. of Interior+US BM+OSHA Ultra Enviro-Systems (P) Ltd. United Nations Environmental Program
6.38 25.3 7.52 12.2 22.17 6.26 4.2-4.5, 4.7, 4.11 3.1, 3.2, 3.5 22.13 A12.2, A12.5 8.15 4.15, 7.5, 7.7 4.6, 4.20 5.4 6.33 6.27 6.23, 12.4 7.24 8.17 6.21 6.31, 6.34-6.37 6.12, 6.13 12.8 25.4 3.11, 3.14, 17.10 A12.8 22.19 25.1 7.26, 7.30, 7.32, 7.37, 7.47, 7.56, 8.6, 26.10, 26.11 25.2 6.25 6.18 22.11 4.21, 7.16, 7.48, 23.7, 26.6, A7.1 6.22 7.13, 7.14, 7.27, 7.49 6.32 5.9, 6.28, 8.18 6.16 5.7, 5.11, 5.12, 5.14, 5.17, 5.18, 5.29, 5.21 1.8 22.11 1.7 21.2 6.24 A12.7 7.21 25.20 7.54, 9.16, 9.24, 20.1, 20.2, 22.2, 22.18 22.18 1.12 (Continued )
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Credits for illustrations
Book.indb xxxiii
Name of individual or organization
Figures for which credit given
University of Western Australia Untraced Uralmash Vermeer Corporation Vibronics Inc. World Mining Equipment www.all-security-system.com www.geology.ebr.state.nc.us www.soundwaves.usgs.gov
26.14 23.4, 23.9, 23.10, 25.18 5.10 1.10 21.2 7.20 7.53 1.11 1.6
xxxiii
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Chapter 1
Surface rock excavation scenario
1.1
INTRODUCTION
Excavation of rock at the surface of the ground becomes necessary for several reasons. Drill and blast is the most commonly adopted method for such excavations. In essence, this consists of drilling an array of holes in the rock mass, filling them with explosives and then initiating an explosion so that the release of enormous energy breaks the large rock mass into small pieces. This may appear quite simple, but the techniques of drilling the holes and blasting them have advanced very greatly through the experience gained in the past four centuries over which this excavation method has been used. Excavation of rock below the surface of the ground is also essential for various purposes. The techniques, methods and equipment used for underground excavation differ considerably from those used for surface rock excavation. Apart from drill and blast, other alternative techniques and corresponding equipment are used for surface rock excavation. This chapter is devoted to essential background knowledge of surface rock excavation techniques.
1.2
NEED FOR SURFACE ROCK EXCAVATION
Surface rock excavation can mean an excavation carried out by means of equipment that stands on the surface of the ground. However, in this book its meaning is taken as the excavation of the rock lying below the surface of ground at shallow depth. The need for surface rock excavation is felt for three basic reasons as follows. 1 2
Recovery of minerals or crushed rock Creating an unobstructed space These points are elaborated further below.
1.2.1
Recovery of minerals or crushed rock
Many minerals of commercial importance, called ores, are located at shallow depths below the ground surface. Mineral deposit is the term used for a rock mass that
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2
Rotary drilling and blasting in large surface mines
contains a mineral of commercial importance in economically large proportions. A mineral deposit is usually surrounded by other minerals that have no economic importance. These are called waste. Since they often lie above the mineral deposits, as in surface coal mines, they are also called overburden. When mineral deposits lie near the surface of the ground, as illustrated in Figure 1.1, the volume of waste to be removed is comparatively less. In such cases, the whole rock mass is fragmented by one of the rock excavation methods. It is very easy to visually differentiate and pick up ore from such a fragmented rock mass with loading equipment and to transport it by hauling equipment to the place designated for further processing of the ore. To ensure continuity of operation of ore recovery, it also becomes necessary to transport the waste to a dumping place. This is sometimes carried out by using different types of equipment. All the activities narrated in the above paragraphs are together called surface mining, or open pit mining, or opencast mining. Large pieces of rocks are used in the construction of dams, breakwaters and the substrate of road and railway embankments. Smaller crushed rock pieces of different sizes are used in large quantities for making concrete used for the construction of objects such as dams, bridges, harbor wharves, buildings etc. For making very strong and durable structures, the concrete must have great strength. This is possible only when the pieces of rock used in making concrete have a high compressive strength. Therefore, crushed rock is recovered from rock masses in specific areas. The activities carried out for getting crushed rock are similar to those adopted in surface mining. Since no specific mineral is involved and the whole rock mass is of importance, all these activities together are called quarrying rather than mining.
Overburden
Ore Bodies
Figure 1.1 Ore bodies below a hillock.
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Surface rock excavation scenario
1.2.2
3
Creation of unobstructed space
Instances in which it is necessary to make suitable space, by removing obstructions and unevenness of the ground, are met in virtually every type of construction. Some common examples are described below. Huge skyscrapers constructed in many cities have to stand and remain stable on the ground for over a hundred years. A sufficiently competent rock mass, capable of bearing the enormous load imposed by such buildings, is met only at depths in the ground. This makes it necessary to excavate a large rock mass for construction of the foundations. Foundations of bridges have to reach to great depths and have to penetrate into competent rock to a certain depth, because only at such depth is the load-bearing capacity of the ground sufficiently high. The danger of the whole foundation drifting under heavy water currents is also minimized when foundations are constructed in this manner. Roads and railways have to pass through hilly regions at several places. Since locomotives and vehicles have limited grade ability, at many places it is safer and economical to make a long cutting or to construct a tunnel along the road or railway route rather than make a long circuitous route. Depending upon the topography and the width of the road or number of railway tracks, the cutting can be of enormous size. Cuttings of very large dimensions are required to be excavated for water supply or navigation canals. Concrete dams have to be anchored well in the rock mass beneath the ground surface in the river basin as well as at both sides. For properly anchoring a large concrete dam, excavation can spread over a very large area and to great depths. In all the examples mentioned above, surface excavation of the rock has to be carried out by one of the excavation methods. Often a rock mass measuring several million cubic meters is removed in the excavation process.
1.3
STRENGTH OF ROCK MASS AND CLASSES
The word ‘rock’ will be repeatedly used in forthcoming chapters. The meaning of this word is viewed differently by geologists and engineers. To a geologist, every aggregate of minerals means rock, because he is more concerned with the formation, stratification, mineral contents etc. of the rock, rather than its strength. A civil engineer looks at the rock from the viewpoint of strength, load-bearing capacity etc., as he is more concerned with either making something by using rock pieces or breaking a rock mass to create space and construct a structure at the place. A mining engineer is concerned with both the above viewpoints. The properties of rocks from an engineering viewpoint have been elaborated in chapter 3 because they are more closely connected to drilling. The properties of rock masses assume larger importance in the realm of blasting and are, therefore, detailed in chapter 23. The properties of rock masses are however, for the better understanding of various aspects, presented in the first and second chapters, so the concept of the strength of rock must be introduced here first.
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4
Rotary drilling and blasting in large surface mines
Table 1.1 Strength based rock classification. Qualitative name
Strength in MPa
Strength in psi
Extremely Hard Very Hard Hard Moderate Soft Very Soft Extremely Soft
More than 345 207 to 345 103 to 207 69 to 103 34 to 69 7 to 34 Less than 7
More than 50000 30000 to 50000 15000 to 30000 10000 to 15000 5000 to 10000 1000 to 5000 Less than 1000
When referring to the strength of rock, one is most commonly referring to its compressive strength. Let us presume that a cylindrical sample of rock, having a length of 100 mm and diameter of 60 mm is subjected to a steadily increasing axial compressive load and breaks when the load reaches 200 kN. In such cases the strength of the rock can be calculated by dividing the breaking force by the cross sectional area of the sample. For the specific values of the parameters mentioned here, the compressive strength Sc will be: Sc = 200*103/(π*602/4) = 70.735 N/mm2 i.e. Sc = 70.735 MPa For giving some idea of the strength of rocks, adjectives such as soft, hard etc. are arbitrarily applied before their names. The arbitrary nature of these terms can be removed to a good extent by following the nomenclature in Table 1.1.
1.4
DRILL AND BLAST OPERATIONS
Drill and blast operations have a unique place in mining and civil engineering. When the volume of hard rock mass to be excavated is very large, the use of mechanical cutting methods is snail-slow and ineffective. Blasting, however, can quickly fragment even the hardest known rocks over a large area. Breaking of rock mass into fragments requires energy. For any specific weight of rock mass the surface area is inversely proportional to the size of the fragments formed in fragmentation. When rock mass is blasted by the drill and blast method, the confinement of the explosive in the blastholes generates extremely high pressure and a huge quantity of heat. When this pressure and heat passes on to the surrounding rock mass, the rock mass just behind the wall of the blasthole gets pulverized into very fine particles. This happens because in that region new cracks are developed due to high energy concentration. However, at greater distances only the most vulnerable pre-existing cracks expand and join together to form a parting plane. Many such planes give rise to very large fragments in comparison to those formed by the other methods of breaking rocks.
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Surface rock excavation scenario
5
Naturally, blasting proves to be the most economical method of primary rock fragmentation. Figure 1.2 shows the effectiveness of different rock fragmentation methods. The lower and upper curves of the shaded area indicate operations of high and poor efficiency respectively. For a typical surface rock excavation, many blastholes are drilled systematically in a vertical or near vertical direction, filled with explosives and then blasted. These blastholes have specific diameter, depth and distance between them. After removal of the fragmented rock, depending upon the extent of excavation, many step-like benches, as shown in Figure 1.3, are formed.
Energy Required to Break in kW/kg
60 Medium Rotary Blasthole Drill (Chips)
52.5
Extra Large Rotary Blasthole Drill (Chips)
45
Road Header
37.5
Surface Miner
30
Full Face Tunneling
22.5
Ripping
15
Blasting
7.5 0
0
25
50
75 100 125 150 175 200 225 250
Mean Fragment Size in mm
Figure 1.2 Average fragment size achieved by different fragmentation methods.
Bench 1
Bench 2 Bench Height
Figure 1.3 Benches formed in mining process.
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6
Rotary drilling and blasting in large surface mines
Various parts of a bench and its surroundings are shown in Figure 1.4. The top openings of blastholes are in a straight line parallel to the crest of a bench; this line is called the row. Only one row of blastholes is shown in Figure 1.4 but in actual practice 3 to 8 rows of blastholes are drilled and filled with explosive to carry out a blast. More detailed nomenclature of a bench is given in Figure 1.5
Backbreak
Ideal Crest Line Crest or Bench Edge Overburden Bench Top
Blastholes
Overbreak Overhang Bench Face
Sockets
Toe Bench Floor Stump
Burden Below Grade
Figure 1.4 Bench nomenclature.
First Row
Spacing Bench Edge
Bench Top Burden
Bench Height = K Bench Floor Blasthole Bottom Level Reduced Level A Blastholes
Reduced Level of the Bench
Mean Sea Level
Figure 1.5 Bench and blasthole terminology.
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Surface rock excavation scenario
1.5
7
OPTIONS TO DRILL AND BLAST
When all is said and done, blasting is a hazardous and dangerous operation. Due to its harmful effects like vibrations, flying rocks, poisonous gases etc., it is prohibited in areas that are in proximity to human populations, important structures etc. Even if drilling and blasting is carried out in a remote area, it remains a dangerous operation for those who work on the site or live in the vicinity. Before concluding that blasthole drilling and blasting is needed, the necessity and feasibility of other ‘no blast’ methods of rock excavation have to be thoroughly examined. Excavation methods that require no blasting are: 1 2 3 4
Water Jetting Machine Digging Ripping Rock Breaking
In general, if the formation can be described in one or more of the following terms, in most of the cases it can be excavated by one of the above ‘no blast’ methods. Soil, Loess, Talus, Gravel, Peat, Loam, Sandy Clay, Alluvium, Sand Bed, Soft Coal, Shingle, Weathered Limestone, Weathered Rock.
1.5.1 Water jetting Water Jetting is used for excavation in semi-consolidated or loosely cemented mineral masses, that can be termed as soil or extremely soft rocks. In this method a high pressure water jet is directed at the rock mass. The soft rock simply crumbles under the impact of water. Broken rock fragments flow with the water. Gold mines, which are often located in the alluvial areas near a river, or soft sedimentary rock formations on the slopes of hills, use this excavation method. Since the water used for breaking the rock mass cannot easily be recovered, a huge quantity of water is needed for such operations. A typical water jetting operation on hill slopes is shown in Figure 1.6.
1.5.2
Machine digging
If well consolidated and cemented rocks, e.g. iron ore-bearing laterite, have low strength binding between the particles, they can be directly dug by means of rope or hydraulic shovels, provided a wall-type face is available for such digging. Deposits of lignite coal are often horizontal and have a layer of soft sandstone above them. Bucket wheel excavators are preferred for excavation in such lignite mines. Figure 1.7 shows a bucket wheel excavator in action. Buckets are fixed to the rotating wheel of the excavator. As the wheel rotates, the buckets scrape or rip the soft overburden which fills the buckets. As the buckets tilt, the material collected in the bucket falls on the conveyor belt system of the excavator. The conveyer belts then carry the material to the dumping area. Horizontal beds of lignite are thus brought into the open. This method proves economical and effective due to its continuity in excavation.
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8
Rotary drilling and blasting in large surface mines
Figure 1.6 Water jetting in gold mining.
Figure 1.7. A giant bucket wheel excavator.
1.5.3
Ripping
When the rock mass is too hard for excavation by water jetting or machine digging, a third available alternative is ripping. Ripping has been used in the farming industry for thousands of years under the name plowing. Ripping means deep scraping the ground with one or many teeth attached to a moving machine. This method proves very economical in soft and medium-hard formations with compressive strength less than about 70 MPa.
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Surface rock excavation scenario
9
There are many ways in which ripping is carried out. Most of the crawler dozers can be equipped with an optional single shank or multi shank ripper attachment. This attachment is fixed at the back of the dozer and is operated hydraulically to force the tooth or teeth in the ground to the desired depth. This depth is suitably chosen according to machine power and formation hardness. Once this is done the dozer is moved. As the dozer moves, the tooth or teeth break the surrounding rock like a plow in farming. A single tooth behind a large dozer can penetrate even 1 m depth in medium soft ground. Light duty ripping can be carried out with motor graders by using a ripper attachment. Figure 1.8 illustrates a dozer, equipped with a three shank ripper attachment, busy ripping in a medium soft formation. Nowadays large dozers, equipped with powerful engines and strong single or multi-shank ripper attachments, are encroaching on the traditional drill and blast territory. Formation rippability can be decided on the basis of the seismic wave velocity. A well known chart universally adopted for the purpose, is reproduced in Figure 1.9. If formations composed of other types of rocks are encountered, verification of their rippability, either by means of field tests or reliable information on ripping in similar formations, is desirable. If the formation is highly fissured and jointed, ripping becomes easier. During the process of ripping some large size boulders are often encountered. Such boulders can be broken with hydraulic hammers attached to a hydraulic excavator. For a long time, machines having a rotating drum with long teeth have been used for scraping the top bituminous layers of roads. This is also an example of ripping. Recently, similar but very heavy and powerful machines have been introduced in the mining industry under the name ‘Surface Miner’. Some small mines and civil engineering contractors opt for such surface miners. One such surface miner is shown in Figure 1.10.
Figure 1.8 Three shank ripper attachment behind a crawler dozer.
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Rotary drilling and blasting in large surface mines
Seismic Wave Velocity in m/s 0
1000
500
1500
2000
2500
3000
3500
4000
4500
Topsoil Clay Glacial Till Igneous Rocks Granite Basalt Trap Rock Sedimentary Rock Shale Sandstone Siltstone Claystone Conglomerate Breccia Cliche Limestone Metamorphic Rocks Schist Slate Minerals and Ores Coal Iron Ore Rippable
Marginally Rippable
Non Rippable
Figure 1.9 Formation rippability chart.
Figure 1.10 A surface miner.
Surface miners are equipped with a rotating drum on their front side. The drums have tungsten carbide teeth fitted into the fixtures provided for the purpose. As the drum is pushed into the ground by means of hydraulic cylinders provided for the purpose, and is rotated by means of the hydraulic motors within the drum, the teeth
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Surface rock excavation scenario
11
scrape or break the ground. As the teeth of surface miners are of small size, compared to those of a ripper, the broken rock fragments are also of small size. Compared to dozers with a ripper attachment these machines are less effective on hard rocks but more effective on soft rocks.
1.5.4
Rock breaking
A hand held paving breaker is a very commonly used tool for breaking pavements in cities. In essence it is a percussion hammer. When the rock mass to be broken is hard and large in volume, the energy output of the paving breaker falls very short of requirements. Therefore, when large scale hard rock excavation is to be carried out, very large size percussion hammers mounted on the booms of hydraulic excavators are used, as shown in Figure 1.10. Hydraulically driven percussion hammers require very low energy input due to their efficient power transmission as compared to the compressed air hammers. A hydraulic excavator is chosen for mounting these hammers because the hydraulic excavator is easily available, works on hydraulic energy, is very versatile in placing the hammer at any alignment and can reach tall heights. All that has been said above may induce readers to draw a conclusion that in soft and medium-hard rippable formations no blasthole drilling and blasting is necessary. This is not true. Overburden in some coal mines can have marginal rippability. Coal can also have marginal rippability. Even if both overburden and coal in coal deposits are rippable, blasthole drilling and blasting are often practiced for overriding considerations as follows. 1
2
3
4
A sizable part of the sedimentary formation may be non-rippable. In such cases, if bucket wheel excavators are used they may either prove unsuitable or the hard portion has to be blasted by drilling closely spaced small diameter blastholes so that the resulting rock fragments are sufficiently small, and their excavation and transportation can be done by using the bucket wheel excavators. It is to be noted that by their design and construction, bucket wheel excavators are suitable only in easily rippable formations. When the formation in a coal mine is difficult to rip and the coal seam is very thick, it is often more advantageous to use a dragline for overburden removal and shovels for coal removal. In such an event blasthole drilling and blasting becomes necessary in both overburden and coal. Dozer ripping can be done in relatively hard formations. However, when such ripping is done the broken formation remains in a horizontal layer. Loading of waste from such broken layers into the hauling equipment, without unacceptable air pollution, is difficult. Moving the fragmented rock mass by using dozers over a long distance is often impracticable.
One important factor, to be taken into account before resorting to drilling and blasting or ripping, is the cost of crushing. Fragment size distribution achieved by drilling and blasting is far different than that by machine digging or ripping.
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12
Rotary drilling and blasting in large surface mines
Figure 1.11 Hydraulic hammer attached to an hydraulic excavator.
Small size fragments can be obtained by ripping with multi shank rippers. Similar adjustments can also be made in blasting where the diameter of the holes as well as burden (the distance between two adjacent rows), and spacing ( the distance between two adjacent blastholes in a row), can be reduced to give smaller blast fragments. In any case for large scale excavation operations, where the rock is non-rippable, drilling and blasting is the only available alternative. Needless to say nearly 80% of the excavation in mining activities is carried out by drilling and blasting.
1.6
OILWELL DRILLING
One more important and very useful mineral that is abundantly available in the crust of earth is crude petroleum oil. In terms of the importance attributed to the mineral obtained, oil gets far more preference than other minerals that are in solid form. Oil is retrieved from the earth’s crust by means of drilling holes into the oil reserves. Since oil is in fluid form, it can be easily pumped by pumps that are placed on the ground. For this reason, there is no necessity of blasting in the oilwell drilling industry. Rotary drilling techniques are invariably used in current oilwell drilling practice. The reasons are: 1 2 3
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Oil reserves are very often at great depths and it is almost impossible to reach to such depths by using other drilling methods. Progress achieved by other drilling methods is very slow. Many techniques such as hole deviation, well cementation, blowout prevention etc. are not possible without using many accessories that can be used only with rotary drilling.
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Surface rock excavation scenario
13
Due to the economic preference given to the oil industry over the mining industry, far more rotary drilling research is carried out by the oilwell industry than is carried out by the mining industry. However, rotary oilwell drilling vastly differs from rotary blasthole drilling. Similarities between the two are limited only to the fact that both use rotary drilling Table 1.2 Comparison of oil well drilling and rotary blasthole drilling.
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Aspect
Oil well drilling
Blasthole drilling
Hole Depth
Holes are very deep. Depth often exceed 5000 m.
Holes are shallow. Depth rarely exceeds 50 m.
Hole Diameter
Holes may start with large diameter, sometimes even 600 mm, but the diameter reduces to about 200 mm at maximum depth.
Holes have uniform diameter from top to bottom. Diameter ranges between 152 to 445 mm.
Hole Spacing
There is no defined pattern of hole placement.
Holes are always drilled in a well defined rectangular or triangular pattern.
Hole Straightness
Holes are rarely straight. Often they are purposefully deviated.
Holes are required to be straight. Hole deviation is undesirable and avoided.
Need for Casings
Holes are always cased to considerable depth. Casing within casings i.e. telescoping casings are usual. Huge quantity of tubular steel is required for casings.
Holes are never cased.
Drilling Rig
Rig is in modular form. Several pieces of equipment such as diesel engines, derrick, drawnwork, mud pump etc. are assembled at the site of the hole and after completion of the hole, all the pieces are dissembled. Moving from one hole to the other requires 20 to even 100 truckloads and takes 15 days upward.
Rig is composite and is termed a rotary blasthole drill. It contains everything that is needed for drilling the hole. It is moved from one hole to the other in about two minutes and does not require any other carriers. Drill rods are added to the drill string automatically by pipe changer.
Flushing Medium
Water is a usual medium of flushing. Often many different additives are mixed with water for specific purposes. In very few cases gas is also used as medium.
Air is invariable flushing medium. Small quantity of oil and water is mixed for bit cooling, bit lubrication as well as dust suppression.
Drill Pipes
Mostly integral drill pipes are used. High tensile strength is the most important property. Pipes are external upset and internal flush.
Mostly fabricated drill pipes are used. High surface hardenability is the most important property. Pipes are external flush.
Testing in the Hole
Many tests are carried out in the hole at various depths by pausing the drilling.
No test is carried out by pausing the drilling. In fact there are no tests.
Drilling in Sea
Holes are often drilled in sea from specialized ships, jackups, submersibles etc.
Holes are always drilled on land by the drill standing on the surface of ground.
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Rotary drilling and blasting in large surface mines
Figure 1.12 Scenario of a typical oil well drilling operation on land.
techniques and tricone bits attached to the drill string composed of drill rods. Even these accessories, used in these two domains of drilling, are not interchangeable. Of several points of dissimilarities only a few are mentioned in Table 1.2. Due to such differences, the research carried out in rotary drilling by the oilwell industry is of very limited use for the mining industry. A typical scenario of oilwell drilling operation is shown in Figure 1.12. Oilwell drilling equipment usually comprises many components that are carried to the worksite in several trucks and assembled for drilling operations.
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Chapter 2
An overview of blasthole drilling
2.1
INTRODUCTION
The most effective, and hence most usually adopted, method of destruction of an object is by blasting. This method invariably aims at breaking the object into several small pieces. Explosives are used for the purpose of breaking the object. An explosive is a substance that generates a very high volume of gases and a huge quantity of heat within a very short time span of a few milliseconds as soon as it is ignited. When gases are generated by ignition in this way, very high pressure is exerted on the surroundings. At the same time, the strength of the surrounding material also reduces considerably due to the heat. As a result of this, cracks develop in the material and the material gets fragmented. When the place of generation of gases is confined, the gases cannot escape to atmosphere and the intensity of the pressure increases to a considerably higher level. This results in maximum destruction. For the most effective destruction of surrounding material, a hole is drilled in the material to be destroyed. After completion of drilling, the hole is charged (i.e. filled with explosive material) in a certain predefined manner. A detonator, connected by two copper wires, is kept in the hole while the explosive is filled. The hole is then sealed from the top. Once the appropriateness of all the precautionary measures is verified, an electric current is sent through the wires. The electric current heats the wire in the detonator and the explosive material in the detonator explodes. This explosion causes further detonation in the explosive filled in the hole. The hole drilled for filling explosives in this manner is called a blasthole. The process of drilling such holes is called blasthole drilling. A machine used for drilling the hole is called a blasthole drill or often simply a drill. Except cases like breaking a boulder in field etc., a blasthole is seldom drilled in singular. They are drilled in numbers, perhaps in hundreds, one after another, then charged and blasted simultaneously. The fragmented rock mass formed by blasting is almost always moved away, either to create space, as in civil engineering projects or to extract a mineral of interest, as in mining or quarrying. If the sizes of the fragments are small they can be moved easily. In order to have small size of rock fragments after a blast, blastholes are required to be located in a properly planned manner. This requires careful design to determine the diameter, depth and inclination of the holes, and also layout and other relevant parameters of the holes. Rotary drilling is one of the two methods currently used for drilling blastholes in the ground. The other method is called rotary percussion drilling. This book deals with rotary blasthole drilling and covers it from the theoretical as well as practical aspects.
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16
2.2
Rotary drilling and blasting in large surface mines
BASIC CONCEPTS OF DRILLING
In actual practice, drilling is carried out by using a drill bit connected at the bottom of a drill string that consists of drill pipes, as shown in Figure 2.1. A drill string often consists of some other items, not shown in the figure, that have specific purposes. The drilling bit is meant to interact with the rock. When energy in some form is passed to the rock from the drill bit the rock fractures. The drill string enables a continuous supply of energy to the drill bit so that rock fracture continues at progressive depths. As the bit interacts with rock and breaks it, rock cuttings are formed. These cuttings are moved away by means of a circulating fluid that is pumped to the bit through the central bore in all the drill string components. When fluid escapes from the bit and starts moving up to the ground surface through the annular space, as shown in the figure, it drags the cuttings along with it and deposits them on the ground surface. Thus, the bit keeps on breaking fresh portions of the rock.
Drilling Fluid Moving Up through the Annular Space Between Hole Walls and Drill Pipes
Drill Pipe
Drilling Fluid Moving Down through the Central Bore of the Drill Pipes
Drill Bit
Figure 2.1 Drilling Schematics.
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An overview of blasthole drilling
17
The drill string is also given rotary motion so the bit rotates and its cutting points cover a complete cross sectional area to form a cylindrical blasthole. The drilling process can commence, continue, and be complete only when the following essential technical requirements are fulfilled. 1
2 3 4
5
6 7 8 9
The drilling bit, meant for breaking the rock formation, must have adequate hardness, heat resistance, strength, toughness, etc., so it can keep on breaking the formation over a long period of time. Energy, in one of the many possible forms, must be supplied to the bit to cause rock disintegration. The drill bit must be rotated to ensure a circular hole, irrespective of bit cross section. Components in the drill string must also have specific properties to effectively transfer feed force, rotary torque, percussion blows etc., without undue deformations. The rock cuttings formed in the process of formation disintegration must be taken out of the hole as soon as possible, so they do not waste energy in getting crushed once again. The drill should be able to supply sufficient energy to the drill bit through the drill string components to cause formation fracture. The drilling bit must be advanced in the intended direction of the drillhole. The walls of the hole, formed as the drilling progresses, must be prevented from collapsing inside the hole. After attaining the desired drillhole depth, it must be possible to withdraw the drill string composed of drill bit, drill pipes and other accessories.
Combinations of means of fulfilling these essential requirements give rise to different drilling methods. Besides the above technical requirements, it is also necessary to fulfill certain other environmental and economic needs. These are, I Cuttings of very small size – in the form of dust – generated during the process of drilling, should be prevented from mixing with atmospheric air and polluting it. II The process of drilling should continue at optimum speed so as to keep the cost of the operation to a bare minimum. III Noise generated by the drilling equipment must be kept to an acceptable low level. Theoretically, there are many forms of energy that have been applied to a rock mass for its disintegration in laboratories, but in current field practice only mechanical energy is used. Two forms of mechanical energy, viz. percussive and rotary, are used for drilling. In either case the drill string and the drill bit are rotated. Accordingly, the drilling methods are called rotary percussion drilling and rotary drilling. A schematic of both these drilling methods is in Figure 2.2. In rotary percussive drilling the energy generated at the drill is in the form of repeated blows like those of a hammer used in manual chiseling. These blows create
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18
Rotary drilling and blasting in large surface mines
Heavy Feed Force Without Percussive Blows
Low Feed Force With Percussive Blows
High Torque Low Torque
Large Diameter Drill Pipes
Small Diameter Drill Rods Drill Bit With Rotating Cones
Drill Bit Without Moving Parts
Figure 2.2 Schematics of percussion and rotary drilling.
shock waves in the drill string. The shock waves that travel to the bit are passed on to the rock mass through the cutting edges or points on the bit. As a result, cracks are formed in the rock mass and result in formation of rock chips. In rotary drilling the drill pushes the bit into the rock mass with very heavy force. When this heavy feed force is transmitted to the rock mass through the cutting points of the drill bit, cracks develop and rock chips are formed. In both rotary and rotary percussion drilling, cuttings are removed from the hole by a fluid circulated from bottom to top. A drill contains many assemblies meant for various purposes. In a rotary percussion drill the assembly used for generating blows is called the hammer. Hammers are of two types viz. top hammers and down-the-hole hammers. Figure 2.3 shows rotary percussion drilling with these hammers. A top hammer is so called because it always remains at the top of the drill string and hence out of the hole. It creates impacts by causing to and fro movement of a piston contained within it. In addition, it also generates rotary motion through a motor within it. The piston movement or rotary motion are caused by hydraulic fluid or compressed air (now obsolete) depending upon the type of drill. Top hammers give fastest penetration rates in hard rocks. They are mostly used for drilling holes for different purposes on construction sites, small opencast mines,
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An overview of blasthole drilling
19
Feed Frame Feed Motor
Rotary Head
Boom Hoisting Cylinder
Top Hammer Feed Frame Hoisting Cylinder Boom
Drill Pipe Crawler Mounting DTH Hammer
Drill Rod Drill Bit
Figure 2.3 Schematics of top hammer and down the hole hammer drilling.
quarries and underground mines. The diameter of holes that can be drilled with top hammers is usually within 38 to 152 mm. The efficiency of top hammer drilling reduces rapidly as more drill rods are added to reach greater depth. Drilling holes deeper than 30 m with a top hammer is difficult because of the energy lost at the connections of the drill rods. Some top hammers are rated for drilling holes to 50 m depth by use of special drill string components. A down-the-hole hammer, usually abbreviated as DTH hammer, is attached to the bottom of the drill string. A drill bit is attached to the lower end of the DTH hammer. A DTH hammer always goes inside the hole along with the bit. The movement of the piston in the DTH hammer is caused by compressed air. The rotary movement for drilling is generated by the rotary head in the drill and is passed on to the drill bit through drill string components and DTH hammer. DTH hammers give fast penetration rates in hard rocks. They are also used for drilling holes for different purposes on construction sites, opencast mines, quarries and underground mines. The diameter of holes that are usually made with DTH hammers is within 100 to 200 mm. Many large hammers, used with specially built drills, are capable of drilling holes of diameter up to 900 mm and in case of specially built hammers even to 1500 mm. Holes of depth 1000 m or even more can be drilled by DTH drilling as the impact energy used for drilling is generated right above the bit and is, therefore, not lost in transmission.
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Rotary drilling and blasting in large surface mines
Drill bits used with top or DTH hammers are made from alloy steel. On one side they have tungsten carbide inserts. For most of the bits used now, the inserts are in the form of buttons because such bits are found to be superior to chisel shaped inserts arranged in cross or X shape. On their other side, top hammer bits have female threads that enable them to be attached to the drill rods. DTH bits have splines and collars for attachment to the hammers. Tungsten carbide insert buttons are shrink fitted in the holes made in the steel body, whereas the chisel-shaped inserts are brazed into pre-machined grooves by means of a suitable brazing alloy. Top hammer drilling bits are small in size and are firmly coupled to the drill rods by means of threads. DTH bits are loosely coupled to the DTH hammers and have splines to ensure rotation as well as a collar at their top end to ensure that the bits are not detached from the hammer while drilling is carried out or when they are retracted after completion of the hole. Figures 2.4 and 2.5 show some bits used in top and DTH hammer drilling respectively. Schematic of rotary blasthole drilling is shown in Figure 2.6.
Button Bit
Cross Bit
X Bit
Figure 2.4 Bits used in top hammer drilling.
Splines
Collar Cross Bit
X Bit
Button Bit
Figure 2.5 Bits used in down the hole hammer drilling.
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An overview of blasthole drilling
21
Mast Drill Head Shock Absorber Mast Braces Machinery House Operator’s Cab
Crawler Mounting
Drill Pipes Stabilizer Drill Bit
Figure 2.6 Schematics of rotary blasthole drilling.
In a rotary blasthole drill, the rotary motion is generated at the drill head by hydraulic or electric motors. The head is pushed down through the feed mechanism to exert the necessary feed force. Drills used for rotary blasthole drilling are very large and heavy, as they have to exert a very high feed force and high torque on the bit through the drill string. Crawler mounting is best suited for the large, tall and heavy drills and for traveling in rough ground. The components, from top to bottom, in a rotary drill string are a shock absorber, a saver sub, drill pipes, a stabilizer and a drill bit. Rotary drilling is carried out by using tricone roller bits in soft to very hard rock mass. In a soft rock mass, drag bits or claw bits are also used. These three types of bits are shown in Figure 2.7. All the three types of bits use tungsten carbide inserts fixed onto a steel body either by shrink fitting, brazing, or mechanical pin or circlip fitting. Tricone bits have three rotating cones fixed into the bit legs through roller and ball bearings. Unlike rotary percussion bits, they have moving parts. Under the heavy force exerted onto the bit, the buttons or teeth on the rotating cones of the bit penetrate into the rock mass. With the heavy torque, they crush, scrape or plow the rock mass to create cuttings. These cuttings are moved out by drilling fluid circulated through the central bore of the drill string components. When rotary or rotary percussion drilling methods are used for blasthole drilling, it is customary to use compressed air as the drilling fluid because it is far more efficient in cutting removal as compared to water or other types of drilling fluids.
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22
Rotary drilling and blasting in large surface mines
Blade Bit
Claw Bit
Tricone Bit
Figure 2.7 Rotary drilling bits.
2.3
PECULIARITIES OF BLASTHOLE DRILLING
Compared to the other purposes of drilling, such as waterwells, oilwells, exploration etc., blasthole drilling has some peculiarities as follows, 1
2
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Blastholes are drilled at the same location. In the case of huge mining projects there may be a need to drill many million blastholes within an area of say ten square kilometers. Most of such mining projects last for fifty or more years. Some mines are productive even for a century. When a set of blasthole drills becomes inefficient after years of use it is replaced. In many mines, not only one or two but several such blasthole drill generations are replaced one after another. In the case of tunnel construction all the blastholes lie along the alignment of the tunnel. Here too, millions of blastholes are localized between the two ends of a tunnel. For other purposes of drilling, such as oil or water wells only a few holes are drilled at one location. The drilling work may last for a few weeks or months at the most. The location of the next drill hole may be far away, sometimes even thousands of kilometers. Blastholes are very near to each other. In surface mining operations several hundred blastholes are drilled at each of the benches. The distance between the blastholes is often from 3 to 12 m. In the case of underground mining or tunneling Operations, many blastholes are drilled at the face of tunnel. The distance between two blastholes can be as low as half meter or even lesser. Therefore, a drill is not required to be moved but the feed frames of the drill can be moved and suitably aligned for drilling each of the blastholes. Movement of the drill becomes necessary for drilling blastholes in the next face after the blasted fragments are removed. This distance is often from 3 to 5 m. In rock faces formed after blasting, blastholes are drilled in a very specific pattern, that is more or less the same as previous. For other purposes of drilling, even if few holes are required to be drilled at the same location, the distance between holes can be very large. Further, in many cases there is no definite pattern for drilling such holes.
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An overview of blasthole drilling
3
4
5
6
7
23
Blastholes are shallow in depth. In tunneling, the blastholes are horizontal and rarely exceed 6 m in length. In underground mining blastholes may be reach 30 m in length or depth. In the case of surface mines, blasthole depths vary from 10 m to 50 m depending upon the mining method and the machinery used for moving the rock. For other purposes of drilling, such as oil or water wells, the hole depth is commonly in terms of hundreds of meters. Particularly in oilwell drilling, hole depth can be between one to seven kilometers or even more, the deepest being over 10 km. Blastholes are drilled in rock masses that have a high degree of uniformity. In surface mining the blastholes are drilled in a more or less uniform rock mass. This rock mass may be stratified but the properties of the strata do not vary greatly. In underground mining, and particularly in tunneling, the rock mass characteristics may greatly differ from one place to the other but the variation within a group of, say, 100 blastholes is very small. For other purposes of drilling, the variety of rock masses can be very large. In very few cases, if any, the rock mass at two locations of the hole is the same. Blastholes are drilled in the same environment. In surface or underground mining as well as tunneling, the holes are required to be drilled at almost the same depth and surroundings. For other purposes of drilling, two holes may be in radically different surroundings. For example, one waterwell may be in a busy locality near a sea port, whereas the next hole may be at a hill station at high altitude. No testing is done in blastholes. In the case of blastholes, no testing is involved. The holes are drilled, charged with explosives and fired. After removing the fragmented rock by using loading and hauling equipment, the process of drilling is continued once again over the freshly exposed area. In a few instances it may be essential to carry out a few tests in blastholes but such tests are carried out only after the hole is drilled and the drill is moved to next blasthole location. For other purposes of drilling, a lot of ‘in hole’ testing may have to be done in the midst of the drilling process. For hole path orientation test, permeability test, borehole logging etc. special tools have to be lowered into the hole. Blastholes are always straight. Barring unavoidable deviation, blastholes are intended to be in a straight line. In other purposes of drilling, holes may be purposefully deviated and so are on a curved path.
In view of all these peculiarities of blasthole drilling, the drills, components and tools are designed with an objective of giving fastest economical output in terms of meters drilled.
2.4
COMPARISON OF DRILLING METHODS
Rotary percussion and rotary drilling methods described above have been compared in Table 2.1. The comparison is made on a technological basis from the blasthole drilling viewpoint.
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Book.indb 24
Table 2.1 Comparison of blasthole drilling methods on the basis of essential requirements. Rotary percussion blasthole drilling
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Criteria of comparison
Top hammer drilling
Down the hole hammer drilling
Rotary blasthole drilling
Means of Achieving Formation Disintegration
By means of blows generated at the drilling machine located above the ground.
By means of blows generated in the down the hole hammer placed just above the drill bit in the drill hole.
By means of static pressure given by the feed force exerted by the drilling machine placed on ground.
Type of Drill String
Relatively slender as it comprises of drilling rods coupled together by means of sleeves.
Relatively rigid because it comprises of large diameter drill pipes that have male and female threads on their ends. No sleeve is required for coupling the pipes.
Rigid because it comprises of even larger diameter and heavier drill pipes that have male and female threads on their ends. No sleeve is required for coupling the pipes.
Transfer of Impacts and Rotary Motion
Both impacts and rotary movement are transferred through drill string.
Only the rotary movement is transferred through drill string. Impacts are generated at the hammer in the hole just above the bit.
Very heavy static feed force instead of impacts and rotary movement are transferred through the drill string.
Practical Hole Dia. Range
38 mm to 150 mm.
89 mm to 250 mm.
150 mm to 445 mm.
Practical Hole Depth Range and Susceptibility to Hole Deviation
0 to 18 m. Highly susceptible to deviation at higher depths due to high slenderness ratio of the drill rods.
0 to 100 m. Hole deviation is relatively small because drill pipes have low slenderness ratio.
0 to 60 m. Hole deviation is relatively small because drill pipes have low slenderness ratio.
Book.indb 25
Possibility of Angle Hole Drilling
As holes are drilled with light equipment they can be drilled at any angle.
As holes are drilled with medium equipment drilling holes in direction other than vertical or near vertical is difficult.
Drilling equipment is very heavy.Vertical or slightly inclined holes are possible.
Effect of Depth on Penetration Rate
Penetration rate is considerably reduced with depth because of loss of percussion energy. It is absorption by drill rods.
No significant loss of penetration rate as the hole depth increases because no energy is lost in transmission.
Penetration rate is unaffected by increase in the depth of the hole because impact energy is not used.
Means of Cutting Removal from the Hole
By means of compressed air circulated through the drill string. Sometimes water is also used as circulation medium.
By means of compressed air circulated through the drill string.
By means of compressed air circulated through the drill string.
Magnitude of Feed Force Required on Drilling Tool
Small force is essential to ensure that the bit is in contact with the formation.
Small force is essential to ensure that the bit is in contact with the formation.
Very heavy force is essential to ensure that the bit remains in contact with the formation and crushes the formation beneath it.
Effect of Extreme Cold Weather
As hydraulic oil is the medium of power transfer, great difficulties are experienced due to oil viscosity changes.
As compressed air is the medium of power transfer, difficulties in cold weather are easily surmountable.
Electric drills operate satisfactorily. In hydraulic drills difficulties are experienced due to oil viscosity changes.
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Rotary drilling and blasting in large surface mines
Table 2.2 presents a qualitative comparison between top hammer and downthe-hole hammer drilling methods in typical limestone quarries. Faster penetration rate is one of the most important criteria of a drilling method. Figure 2.8 illustrates the penetration rates experienced while drilling about 150 mm blastholes in quarries by Rotary, DTH and Top Hammer drilling methods. The variation of penetration rates is of a qualitative nature.
Table 2.2 Qualitative comparison of TH and DTH drilling. Criteria of comparison
TH drilling
DTH drilling
Diameter of Blastholes in mm Penetration Rate Achieved Straightness of the Blasthole Suitability for Drilling Deep Blastholes Production Capacity per shift Low Fuel Consumption Economic Service Life of Drill String Suitability for Drilling in Difficult Conditions Suitability for Drilling in Good Conditions Simplicity of Drilling From Operator’s Viewpoint
75–127 ♦♦♦♦ ♦♦♦ ♦♦♦ ♦♦♦♦ ♦♦♦♦♦ ♦♦♦ ♦♦♦ ♦♦♦♦♦ ♦♦♦♦
127–300 ♦♦♦ ♦♦♦♦♦ ♦♦♦♦♦ ♦♦♦ ♦♦♦ ♦♦♦♦ ♦♦♦♦ ♦♦♦♦ ♦♦♦♦♦
Hard Monolithic Rock Mass DTH
Rotary Drifter
Increasing Penetration Rate
Increasing Penetration Rate
Soft Rock Mass
DTH Rotary
Increasing Rock Strength
Drifter
Increasing Rock Strength
Increasing Penetration Rate
Hard Fractured Rock Mass
DTH Drifter Rotary
Increasing Rock Strength
Figure 2.8 Penetration rates experienced in drilling about 150 mm blastholes in limestone.
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An overview of blasthole drilling
27
It must be borne in mind that the trend of penetration rates shown in Figure 2.8 does change when the rock to be drilled is different and particularly when the holes are of larger or smaller diameter.
2.5
CHOICE OF A BLASTHOLE DRILLING METHOD
Once it is decided that blasthole drilling is to be carried out in preference to other ‘no blast’ methods of rock excavation described in chapter 1, the next logical question is “Which should be the blasthole drilling method?”. The main factors that influence the choice of drilling method are: 1 2 3 4
Period over which operations continues. Diameter of the hole. Depth of the hole. Characteristics of the formation to be drilled.
The most important factor that affects the choice of method is the continuity of operations. In the case of civil engineering, the quantum of blasthole drilling to be carried out is relatively very small as compared to mining. To be specific, a civil engineering contractor may rarely have a need for continuous blasthole drilling at one location for more than two years. Thus, for a civil contractor it is wise to select many smaller machines rather than few large machines for his civil engineering projects, because after completion of the project the smaller machines can be distributed on many separate small jobs. Blastholes in civil engineering are rarely larger than 127 mm in diameter and almost never larger than 200 mm. The depth of a blastholes is also limited to about 20 m. Civil engineering firms, therefore, usually choose top hammer drilling for blastholes of smaller diameter and DTH drilling in case of larger blastholes. In mining choosing the right drilling method is very critical as the blasthole drilling operations continue for decades and blasthole drills specifically designed for a method are to be procured before commencing the operation. Choice of incorrect drills can be disastrous as reselling drills at a fair price is difficult. When blasthole drilling is to be carried out it in the soft sedimentary overburden of opencast coal mines, the diameters of the hole range between 152 to 381 mm and depths range between 15 to 55 m. In soft thick coal beds blasthole diameters range between 152 to 200 mm and depths between 10 to 20 m. For soft sedimentary formations suitability of a blasthole drilling method is illustrated in Figure 2.9. In many opencast metal mines the formations are hard and the variation in hardness is over a very wide range. Here both overburden and ore are removed by using shovels and dumpers. For this reason the blasthole depths lie between 10 to 20 m and diameters can be from 100 to even 445 mm. For all such cases the blasthole drilling method can be chosen on the basis of the illustration in Figure 2.10. In underground coal or metal mines, blasthole diameters are between 51 to 100 mm but in certain cases can be as high as 152 mm. The depths of blastholes can be from 2 to 30 m.
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Rotary drilling and blasting in large surface mines
Top Hammer Rotary with Drag or Tricone Bit Rotary with Tricone Bit (Rarely Needed) Rotary Drag Bit
Rotary with Tricone Bit
Soft Very Soft 50
0
100
150
200
250
300
350
400
450
Blasthole Diameter in mm
Figure 2.9 Suitability of blasthole drilling method in very soft and soft formations. TH Drilling DTH Drilling TH or DTH Rotary with Tricone Rotary with Drilling Bit or DTH Tricone Bit Extreme Hard Very Hard Hard Medium Hard 0
50
100
150 200 250 300 Blasthole Diameter in mm
350
400
450
Figure 2.10 Suitability of blasthole drilling method in medium to very hard formations.
In coal mines it is very common to use rotary drag bit drilling. In hard coal formations top hammers are often used for drilling blastholes. When holes of diameter 152 mm to 30 m depth are to be drilled, as in some metal mines, DTH hammer is the most usual choice because the size of the drilling machines used for DTH drilling is small. In tunneling or shaft sinking operations, diameters of blastholes are 38 to 64 mm. The depths of the blastholes, in a very few instances, can be as much as 25 m but in most of the cases depths are between 2 to 6 m. The top hammer rotary percussion drill is almost exclusively used in all such operations. Two important points must always be remembered while choosing the right drill for blasthole drilling. Rotary drilling requires exerting a heavy feed force in the downward direction on the drill string. Hence the drills are very heavy. If they are not heavy, their ability in drilling large diameter blastholes remains limited and must be investigated cautiously. Percussive drills need not exert heavy feed force, but when drilling upward holes the weight of the piston acts adversely to the blow energy. Percussive drill performance must be judged in the light of this fact.
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An overview of blasthole drilling
29
The performance of blasthole drilling equipment can be judged by means of carrying out certain laboratory tests. However, due to many influencing factors, it cannot be totally relied upon, while tackling an important issue like choosing a blasthole drill. Field trials by using hired equipment before taking a final decision about the suitable blasthole drilling method is a far better alternative, even if it is expensive. In this chapter both rotary percussion and rotary drilling methods have been described briefly. As the subject of this book is rotary blasthole drilling, in all forthcoming chapters all the elaborations are presented from the a viewpoint of rotary blasthole drilling.
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Chapter 3
Properties of rocks
3.1
INTRODUCTION
Blasthole drilling is a part of earth excavation processes. Needless to say, some basic knowledge of earth science, and particularly the properties of rock and rock masses, is very essential to a person engaged in blasthole drilling. Earth science is a very wide subject and has many branches within it. Branches that are important in the context of blasthole drilling are geology, mineralogy, meteorology and rock mechanics. This chapter presents essential knowledge of Geology and mineralogy. Topics on rock mechanics are covered in chapter 9. Meteorology is elaborated in chapter 11. The properties of rock masses have been covered in chapter 23 as they are more relevant to blasting than drilling.
3.2
EARTH AND ITS INTERIOR
Earth is one of the planets revolving around the sun in an elliptical orbit, having a mean diameter of 1.496 × 106 km. The shape of the earth is almost spherical. Earth rotates around its axis and completes one rotation in one day. The two points at which the earth’s rotational axis crosses its surface are called poles. According to the magnetic properties of the earth these poles are named south and north. An imaginary circle, perpendicular to the earth’s rotational axis, that lies on the surface of the earth and thus divides the earth into two equal hemispheres, is called an equator. The equatorial diameter of the earth is 12756.6 km and the polar diameter is 12713.8 km. The volume of the earth is 1.08323 × 106 km3. The total surface area of the earth is calculated to be 510 × 106 km2 of which land covers only 149 × 106 km2 i.e. 29.22%. The remaining 70.78% area is covered with water. Average land height is estimated to be about 840 m above the mean sea level. The average depth of the water below the mean sea level is 3804 m. Various experiments have proved that the average density of the earth is 5517 kg/m3 and accordingly the mass of the earth works out to be 5.976 × 1024 kg. The interior of the earth consists of three zones as illustrated in Figure 3.1. The innermost zone, spherical in shape with a radius of about 3400 km, is called the core. The core itself is subdivided into an inner core and an outer core. The outer core is
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Rotary drilling and blasting in large surface mines
Inner Core Outer Core Inner Mantle Crust
Outer Mantle
Figure 3.1 Interior of the earth.
believed to be composed of molten iron and nickel and has a density of 10000 kg/m3. The inner core is believed to be in a solid state and has a density of 17000 kg/m3. The next outer zone, called the mantle, has a thickness of about 2900 km. The mantle is again subdivided into two zones viz. inner and outer mantle. The inner and outer mantle have a density of 5700 kg/m3 and 3300 kg/m3 respectively. The outer mantle consists of rock-forming minerals. The outermost zone is called the crust. It is contained in a shell of thickness varying between 5 to 40 km. As illustrated in Figure 3.2, the thickness of the crust is smaller below the oceans and seas and is much larger below the middle portion of the land—particularly mountains. These portions of the crust are called oceanic crust and continental crust. Earth’s crust is also divided in sial and sima. Sial, i.e. outer shell of the crust, is mainly composed of silica and aluminum and hence the acronym sial. Sial is present below the continental shelf and shallow sea/ocean floor. Its thickness varies between 0 to 13 km. Sial is absent below the deep ocean floors. The average density of sial shell is 2700 kg/m3. It is mostly composed of granitic rocks mainly formed from quartz and orthoclase. Sima is the inner shell of the crust. As it comprises silica and magnesia, it is named by the acronym sima. Sima is present everywhere i.e. below the sial as also the ocean floors. Sima thickness increases towards the center of the continental crust. Average density of sima is about 2940 kg/m3. It is mostly composed of basic rocks e.g. gabbro and basalt. Most of the elements found in earth’s crust are in compound form of which oxides are the main type. The proportions of various oxides found in the earth’s crust is given in Table 3.1. The proportion of elements that form these compounds is given in Table 3.2. Oxides chemically combine with each other, or with other substances, to form minerals that are the main constituents of rock. A mineral is an inorganic substance having definite chemical composition and properties. Earth’s crust contains more than 2000 minerals.
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Properties of rocks
Mountains
33
Oceanic Crust Ocean
Sial
Sima Mantle
Figure 3.2 Earth’s crust.
Table 3.1 Oxide proportion in earth’s crust. Common name of oxide
Chemical formula
% Wt. in earth’s crust
Silica Corundum Calcium Oxide Ferrous Oxide Sodium Oxide Magnesium Oxide Ferric Oxide Potassium Oxide Water Titanium Dioxide Other Oxides
SiO2 Al2O3 CaO FeO Na2O MgO Fe2O3 K2O H2O TiO2 –
59.1 15.2 5.1 3.7 3.7 3.5 3.1 3.1 1.3 1.0 1.2
Table 3.2 Element proportion in earth’s crust.
Book.indb 33
Element name
Chemical symbol
% Wt. in earth’s crust
Oxygen Silicon Aluminum Iron Calcium Sodium Potassium Magnesium Titanium Hydrogen Other Elements
O Si Al Fe Ca Na K Mg Ti H –
47.0 27.0 8.0 5.0 3.5 3.0 2.5 2.0 0.5 0.17 1.33
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3.3
GEOLOGICAL CYCLE
One of the fundamental laws in respect of matter states that matter can neither be created nor destroyed but can only be altered. Accordingly, all the land that has come into existence, is actually as a result of transformation through many activities. Such activities are going on even now and will never stop. Figure 3.3 shows a flow diagram which indicates all such activities. These activities are geologic in nature i.e. they commence, continue and complete within the earth and its surroundings. Being cyclic in form they are called the ‘Geological Cycle’. Magma is the most primary matter in the geological cycle. It is a very hot, viscous, siliceous rock melt containing many minerals, water and gases. In addition to silica and oxygen, that are the major constituents of magma, it also contains elements such
Magma Mixing of Magma with Water in Volcanic Activity
Mixing of Magma with Air in Volcanic Activity
Solidification of Magma After Igneous Intrusions and Extrusions
Matter Suspended in Air
Matter Suspended in Water
Biological Activity
Biological Activity
Sedimentation and Deposition Sedimentation and Deposition
Igneous Rocks
Weathering and Transportation
Anatexis i.e. Regeneration of Magma by Fluxing of Pre Existing Rocks Biological Activity
Metamorphism Biodeposits
Biological Activity
Sediments and Soils
Diagenesis
Diagenesis
Weathering and Transportation Anatexis i.e. Regeneration of Magma by Fluxing of Pre Existing Rocks
Weathering and Transportation
Weathering and Transportation
Sedimentary Rocks
Weathering and Transportation
Weathering and Transportation
Metamorphism
Metamorphic Rocks
Figure 3.3 Geological cycle.
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Properties of rocks
35
as potassium, sodium, calcium etc. Magmas are accumulated at great depth below the surface of the earth. Due to superheated steam and hot gases, magma always exerts very high pressure on its surroundings. Such high pressure sometimes exceeds the limit of strength of the surrounding mass, resulting in formation of cracks in the surrounding mass. Being in a fluid state, magma flows through such cracks and either intrudes below or extrudes above the surface of the earth. Depending upon the magnitude of the pressure and the rate at which the magma flows, violent volcanic eruptions or slow intrusions take place. When magma cools down, it transforms into different types of rocks. Such rocks, formed after magma intrusion or extrusion, are called igneous rocks. Igneous rocks are found at great depths below all the surface of earth i.e. land or oceans. Various weathering agents e.g. water, wind, sea, river, glacier, biological activities, trees etc. cause continuous disintegration of igneous rocks. Resultant materials, that are in unconsolidated granular form, are called sediments. Sediments are often moved away by transporting agents such as streams, rivers, wind. These sediments eventually get deposited at other locations. When these sediments get cemented and/or consolidated, they turn into a hard and strong sediment mass that is called sedimentary rock. In volcanic activities, sometimes the sedimentary and igneous rocks get heated to very high temperature and are also subjected to very high pressure. Under such conditions the igneous and sedimentary rocks undergo transformation on a large scale to form metamorphic rocks. In some very large scale volcanic activities, all the three kinds of rocks are thrown from the place of their original formation to great depths below the earth’s crust. Inflow of enormous heat from the surroundings increases the temperature of these rocks to such a level that they melt and transform into magma. This process of magma formation is called anatexis.
3.3.1
Igneous rocks
Igneous rocks are composed of minerals that have silica as one of the main constituents. These minerals are quartz, orthoclase, plagioclase, hornblende, mica, augite and olivine. A mineral composition and texture-based genetic classification of igneous rocks is generally followed in geology. According to this basis of classification, there are four classes of igneous rocks, viz. acidic, intermediate, basic and ultrabasic. Acidic rocks have higher quartz contents than other igneous rocks. Quartz crystals can be separately observed in such rocks. The other major constituent of acidic rock is orthoclase. Besides this plagioclase, mica and hornblende are also present. Intermediate rocks may have some quartz and orthoclase but usually less than 10% each. The main constituents of intermediate rocks are plagioclase, mica and hornblende. Basic rocks do not contain any quartz. Their main constituent is plagioclase. Other constituents are hornblende, augite and olivine. Ultrabasic rocks are mostly formed from olivine and have some quantities of augite and plagioclase. Acidic rocks are generally lighter in color, whereas intermediate, basic and ultrabasic rocks are increasingly darker in that order.
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Rotary drilling and blasting in large surface mines
The density of acidic rocks is about 2600 kg/m3. For intermediate, basic and ultrabasic rocks it increases gradually to finally reach the level of 3300 kg/m3. The names of some igneous rocks are: granite, basalt, andesite, syenite, dolerite, gabbro, granodiorite, peridotite, ignimbrite, rhyolite, diorite, pumice, ash, eclogite, pitchstone. Figure 3.4 illustrates different classes of igneous rocks mentioned above. Igneous rocks are also classified on the basis of their texture i.e. the relationship between the grains forming the rock. It depends upon grain size, grain shape, degree of crystallinity and grain contact relationship. In this type of classification specific names of the rocks are used. Unfortunately the names of rocks are colloquial in nature and, therefore, there is no unification or standardization in the names of the rocks. Some rocks and their texture-based classification is shown in Table 3.3.
3.3.2
Sedimentary rocks
In the transportation of fragments formed by weathering and erosion, the solid materials get deposited. These deposits are in the form of layers and are called sediments. Sedimentary rocks are formed from these sediments. Therefore all sedimentary rocks are layered. Consolidation, cementation and crystallization turn sediments into sedimentary rocks.
Approximate Percentage of Principal minerals
Acidic
Intermediate
Basic
Ultrabasic
100
75
Quartz Plagioclase
50 Orthoclase
Mica
25
Augite
Olivine
Hornblende 0
Figure 3.4 Composition of igneous rocks. Table 3.3 Classification of igneous rocks. Nature of composition Grain and crystal
Rock type
Coarse Grain, Large Crystals Plutonic Medium Grain, Large Crystals
Acidic
Granite Porphyritic Granite Hypabyssal Granophyre
Medium Grain, Medium Crystals Fine Grained, Medium Crystals Fine Grained, Small Crystals Extrusive
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Intermediate Basic
Ultrabasic
Diorite Porphyrite
Peridotite
Quartz Porphyry Rhyolite Andesite
Gabbro
Dolerite
Basalt Trachyte
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Properties of rocks
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Often particles in a sediment are of very small size. Due to their predominantly cohesive nature, these particles remain bonded together. Very small pore spaces within these particles are filled with water. When pressure gets exerted on such sediments the water from the pore spaces gets oozed out. In absence of water all the pressure is borne by the particles so the cohesive bonding between particles increases. Eventually a solid hard mass is formed. This mass is called sedimentary rock. The process of oozing out of water and the formation of a solid mass is called consolidation. In certain cases the sediments have particles formed out of very fine materials of a particular nature—mainly calcareous in their chemical composition. When these materials come in contact with water, a chemical reaction takes place and the particles stick together with a very high bonding force. The result is the formation of a hard mass. This process, called cementation, also gives rise to sedimentary rocks. Occasionally, the cementing materials occupy spaces between larger fragments deposited earlier. In such sedimentary rocks two distinct particle sizes are easily noticeable. The process of crystal formation from dissolved minerals often occurs in the shallow parts of the sea or in lakes, in desert areas etc., where evaporation is much higher than precipitation. The sea or lake contains dissolved minerals such as calcium bicarbonate and calcium sulfate. As evaporation takes place, water is lost and the dissolved minerals form crystals. These crystals settle on the bottom of the sea or lake. As evaporation continues, more crystals from and accumulate on the sea or lake floor. After some time interval, they transform into sedimentary rocks. Sedimentary rocks are found near the surface of the earth. They may also be composed of many different types of minerals. Their thickness is usually less than 1 km but can be as high as 10 km. The names of some sedimentary rocks are: breccia, conglomerate, sandstone, limestone, coal, lignite, chalk, dolomite, arkose, greywacke, shale. Sedimentary rocks are classified on the basis of particle size as shown in Table 3.4, or on the basis of their chemical composition as shown in Table 3.5.
3.3.3
Metamorphic rocks
Metamorphic rocks are formed by recrystallization or transformation of already existing rock masses or sediments. The process is called metamorphism. It requires high temperature or pressures or both. Table 3.4 Classification of sedimentary rocks on the basis of their particle size. Class of rock
Grain size
Name of the fragment class
Names of some of the rocks in the class
Rudaceous
Larger than 256 mm
Boulders
Conglomerate (Rounded Fragments) or Breccia (Angular Fragments )
Between 64 and 256 mm
Cobbles
Between 2 and 64 mm
Peebles
Between 1/16 and 2 mm
Sand
Arenaceous
Argillaceous Between ½56 and 1/16 mm Silt Less than ½56 mm
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Clay
Sandstones Siltstones Mudstones, Shale
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Rotary drilling and blasting in large surface mines
Table 3.5 Classification of sedimentary rocks on the basis of their chemical composition. Composition
Chemical sediment
Organic sediment
Calcareous Carbonaceous Siliceous Ferruginous Aluminous Phosphatic Saline
Oolithic Limestone, dolomite – Flint, Chert Oolithic Limestone, Laterite Bauxite Phosphorite Rock Salt, Gypsum, Anhydrite
Reef and Shelly Limestone, Chalk Peat, Lignite, Coal Diatomite, Radiolarite Bog Iron Ore – Guano, Bonebeds –
Metamorphism is of four different types as follows. 3.3.3.1
Contact metamorphism
In contact metamorphism very little pressure is involved. The metamorphism is mainly due to heat that transforms matter into metamorphic rocks. During the process of magma intrusion, the heat from magma is given to the adjoining rock. Very high temperatures at the contact between the intrusive body and surrounding older formation causes recrystallization. In this manner a layer of metamorphic rocks is formed around an intrusive body. In the case of small intrusive bodies, the thickness of such layers can be very small but for large intrusive bodies it can even be 2 to 3 km. 3.3.3.2
Regional metamorphism
Changes in the lower crust and mantle exert high pressures and medium temperatures in some parts of the earth’s crust. If the disturbances are very large, existing rocks may be thrown as deep as 10 to 15 km, where pressure and temperatures are very high. These conditions result in metamorphism in that region. The metamorphic rocks formed in this manner are not in layered form but in massive bodies lying within that region which often have dimensions exceeding tens of kilometers. 3.3.3.3
Dynamic metamorphism
Changes in the crust and mantle of the earth can develop cracks in the ground. Very often the layers on one side of the crack are moved up or down relative to others. Such movement generates very high pressure and temperature at the crack. These give rise to metamorphic rocks in thin layers. As the reason of the metamorphism is movement, it is called dynamic metamorphism. 3.3.3.4
Impact metamorphism
When a large meteorite hits the earth’s surface, the impact generates heat and pressure that results in the formation of metamorphic rocks in the immediate vicinity of the impact area.
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Properties of rocks
39
Names of some metamorphic rocks are: slate, schist, gneiss, marble, quartzite, skern, hornfels, amphibolite, mylonite, granulite, epidorite. Classification of metamorphic rocks is less rigidly adhered to than rocks of the other two types. A few rocks are recognized on the basis of their distinct textural and mineralogical features. Composite names are used to achieve greater precision in describing metamorphic rocks. In such composite names, the main constituent of the rock is usually mentioned first, e.g. Garnet-Mica Schist. Table 3.6 describes the generally accepted classes of metamorphic rocks. These classes are based the original rock from which they were formed. Table 3.6 Classification of metamorphic rocks on the basis of original rocks and formation process. Original rock Argillaceous Sediments
General description
Slate
Very fine grained with flat platy cleavages. Often dark in color. A product of low grade metamorphism.
Phyllite
Fine grained with undulating, lustrous cleavage planes. Commonly greenish in color. Exhibits higher metamorphism than slate.
Mica Schist
Medium to coarse grained with rough often puckered cleavage parallel to mica foliation. Middle to high grade of metamorphism.
Gneiss
Medium to coarse grained. Foliation of streaky bands of mica, quartz and feldspar. High grade metamorphism.
Granofels
Fine to coarse but even grained, unfoliated, often quartofeldsparthic. A product of medium to high grade metamorphism.
Arenaceous Sediments
Quartzite
Highly interlocking quartz grains. Usually lacks foliation. Occasionally contains mica crystals. Mostly white in color. High grade of metamorphism.
Calcareous Sediments
Marble
Contains interlocking calcite and dolomite grains. Sometimes contains calcium and silicate minerals. Usually white, gray or buff color. A product of high grade metamorphism.
Basic Igneous Rocks
Antinolite and Chlorite Schist
Greenish rocks with high degree of foliation. Undulated or puckering cleavage. Low to middle grade of metamorphism.
Amphibolite
Medium to coarse grained rocks. Usually found in blocks. Usually foliated and banded. Hornblende and plagioclase are the main constituents. High degree of metamorphism.
Eclogite
Medium to coarse grained rocks. Greenish red color due to red garnet set in a matrix of green pyroxene. Exhibits high grade of metamorphism.
Hornfels
Mostly fine grained rocks. Formed by contact metamorphism. Dark in color. Unfoliated.
Mylonite
Very fine grained flinty rocks. Thinly banded. Usually dark colored. Formed as a result of mechanical crushing and grinding.
Mixed Sediments or Acidic IgneousRocks
Various Sediments
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Metamorphic rock
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40
3.4
Rotary drilling and blasting in large surface mines
FORMATION OF ROCK MASS
Rock mass is formed at or near the surface of the earth by different activities that take place near the surface of the earth. Main activities in this regard are 1 2 3
Volcanic Activities Plutonic Activities Surface Activities
These have been briefly accounted for in earlier parts of this book. The following is a larger elaboration.
3.4.1 Volcanic activities Volcanic activities are those in which magma, which exists at depths within earth under very high pressure and temperature, finds a way to flow above or near the surface of earth. In this activity different types of rock masses are formed. In geological terms they are called structures. Two classes of volcanic structures are known. They are, 1 2
Extrusive Structures Hypabyssal Structures
3.4.1.1
Extrusive structures
Extrusive structures are formed when magma erupts above the ground surface. Such magma, that erupts on the surface of ground, is called lava. Some extrusive structures are lava plateaux, lava cones, hornitos, craters, calderas etc. Descriptions of these extrusive structures are given in Table 3.7. Usually lava cools very rapidly at the surface of the earth. Therefore, rocks found in these structures are fine grained. 3.4.1.2
Hypabyssal structures
When structures are formed by magma intrusion at shallow depths, they are termed hypabyssal structures. Some hypabyssal structures are volcanic plugs, dikes, laccoliths, sills etc. Descriptions of these hypabyssal structures are given in Table 3.8. As the cooling rate of magma in these structures is not very high, the rocks formed in them are medium grained. Occasionally, when the earth surface covering these structures gets eroded, they become visible on the ground surface.
3.4.2
Plutonic structures
Plutonic structures are formed by magma intrusions at great depth below the surface of the earth. Usually these structures are of very large size that ranges from ten to even many hundred kilometers. The rate of cooling of the magma is very slow. More often than not, these structures have layers of rocks. Due to the extremely slow process of
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Properties of rocks
41
Table 3.7 Extrusive structures. Name
Description
Lava Plateau
Lava plateau are formed on the surface of the earth by flow of magma through fissures. Fissures are thin but long and deep seated gaps at the surface of the earth. Magma that flows through these fissures is low in viscosity and flows over a very large surface. It cools rapidly and rocks formed in the process are basalt.
Lava Cone
Lava cones are formed when lava flows to the surface of earth through medium sized pipelike structures. Usually such lava has somewhat intermediate viscosity and, therefore, spreads less easily on the surface of the earth. In the process a hillock with gentle slope on all the sides is formed.
Hornito
Hornitos are formed when highly viscous lava flows to the surface of earth through large structures. Hornitos are tall but small diameter structures with steep slopes on the sides.
Crater
Craters are depressions in the ground. They are formed at the mouth of a volcano. Craters can also be formed when a meteorite hits the earth surface.
Caldera
Caldera are very large depressions on the surface of the earth. They are formed by coalescence of several small craters or repeated explosions of a volcano at adjacent places or ground subsidence of a large portion of earth surface in a huge hollow formed just below the surface of the earth during a violent volcanic eruption. Calderas can be very large in diameter – often tens of kilometers.
Table 3.8 Hypabyssal structures. Name
Description
Volcanic Plug
During volcanic intrusions magma often flows to the ground surface through circular feed channels. After the volcanic activity comes to an end such channels remain filled with solidified magma. Later the surrounding structure may get eroded, leaving behind the magma channel in vertical or near vertical position. This is called a volcanic plug.
Dike
Often magma flow cuts across the bedding planes of the older formations and forms sheet like structures. These are called dikes. Many different types of dikes – based on their shapes – are known e.g. parallel dikes, radial dikes, ring dikes etc.
Laccolith
Intrusive magma often takes the shape of a dome having flat underlying surface and arch like top surface. In appearance this structure looks like a mushroom. These are called laccolith. Laccoliths are called composite when they are formed from two different sources of magma. Similarly two or more laccoliths formed from the same magma source are called multiple laccoliths. Cedar tree laccoliths are the laccoliths lying one above another and resemble to the shape of a cedar tree. All these laccoliths are formed from the same magma source.
Sill
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When magma intrudes in between two bedding planes and forms a sheet like structure, of nearly same thickness and negligible concavity, it is known as a sill. Sills, like laccoliths, can be multiple or composite.
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Rotary drilling and blasting in large surface mines
cooling, the fluid state of magma lasts for a long time. Acidic magma, which has lower density, floats at the top and the basic magma settles down to greater depth. In this condition the upper rocks are coarse grained acidic—granites—and the lower rocks are coarse grained basic—gabbro. Various plutonic structures are batholiths, stocks, lopoliths, phacoliths etc. Descriptions of these extrusive structures are given in Table 3.9. Table 3.9 Plutonic structures. Name
Description
Batholith
Batholiths are gigantic structures having a cross sectional area of more than 100 km2. The structure goes on enlarging horizontally with increasing depth. There is no apparent bottom to the batholith. It can be considered as a part of inner crust. Stocks are similar to batholiths but much smaller in their physical dimensions. Structures smaller than 100 km2 area are classified as stocks. When the magma intrusions take a saucer like shape with the upper and lower concave surfaces as seen from the top, the resulting structure is called a lopolith. By their nature the lopoliths are formed by magma intrusions between two concave layers. The path of magma is somewhat narrow. Therefore, the type of magma that intrudes in lopoliths is basic in nature. Naturally, rocks formed in the lopoliths are coarse grained basic rocks. When magma intrudes in between two convex beds, as seen from the top, the resultant structure is called phacolith. In other words phacoliths are similar to lopoliths but inverted upside down.
Stock Lopolith
Phacolith
Table 3.10 Surface structures. Name
Description
Unconformity During the deposition of the sediments that give rise to the beds, there is a time interval during which no deposition takes place. Therefore, a surface formed during this time interval separates the old beds from new beds. On many occasions, due to change in the mode of deposition or change in the type of sediment, such surface is very distinct and is called unconformity. Angular, Parallel, Non Depositional and Hetrolithic are the four different types of unconformities. Joints During tectonic activities different types of stresses are developed on the beds. Rocks are capable of withstanding high compressive stresses but are weak in tension. Due to this, cracks i.e. joints develop at various places in the rock mass. The most important property of joints is that there is no apparent movement on two sides of the joint with respect to each other. A series of parallel joints is called as a set of joints. Joints are classified on the basis of mode in which they were formed. Three classes of joints viz. Shrinkage, Tectonic and Sheet are generally recognized. Faults Fractures in formation beds, along which there has been observable movement of the formation on each of the sides of the fracture plane, are termed as faults. Normal, Hinge, Pivot, Tear and Reverse are the five types of faults Folds During tectonic movements the plane beds are often converted into a wave form. Such structures are called folds. The trough like portion of a fold is termed as syncline, whereas, the crest like portion is named as anticline. There are nearly ten different types of folds.
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Properties of rocks
3.4.3
43
Surface activities
Denudation is a general term used to refer to the processes that alter the surface of the earth. It therefore consists of weathering, transportation and erosion. Weathering means natural disintegration and/or decomposition of the rock mass. Transportation means shifting of the disintegrated or decomposed material away from the place at which they were initially formed. While the broken materials are transported, they come in physical contact with other fresh uneroded formations. The friction developed between the two causes weathering of the fresh formation. This process is termed erosion. As all the above processes take place on the surface of the earth, they are called surface processes. Almost all the surface processes result in deposition of the material on the earth surface in layers and the structures formed in the process are unconformities, joints, faults, folds. A description of different surface structures is given in Table 3.10.
3.5
SOILS, ROCKS AND THEIR PROFILE
The term soil originates from the farming industry. In civil engineering it is defined as an unconsolidated, uncemented mineral mass containing air, water and other impurities of organic nature. Soil is also a product of weathering. It usually lies at or near the surface of the ground. When an excavation below the surface of ground is carried out in the form of a trench or vertical cut, different layers having different constituents and colors are visible on the wall. The composite structure of such layers is called a soil profile. Soil profiles depend upon many factors. Different soil profiles are seen in different regions. Some typical soil profiles are shown in Figure 3.5. In modern nomenclature the ‘O’ horizon is the topmost. It consists of fresh litter in its upper, and decomposed organic matter in its lower, part. It retains soil moisture. The ‘A’ horizon, having a dark blackish brown color, underlies the ‘O’ horizon and consists of a mixture of minerals and organic material. Leaching, i.e. chemical decomposition, removes minerals like Ca, Fe as well as decomposed organic matter to lower horizons. The ‘E’ horizon underlies the ‘A’ horizon. Usually it is a light-colored horizon with ongoing activity of leaching, or the removal of clay particles, organic matter, and/or oxides of iron and aluminum. In some forest regions the ‘E’ horizon has a high concentration of quartz giving the horizon an ash-gray appearance. Beneath the ‘E’ horizon lies the ‘B’ horizon. Most of the fine matter leached in the upper horizons is accumulated in the ‘B’ horizon. The accumulation of fine material leads to the creation of a dense layer in the soil. In some soils the ‘B’ horizon is enriched with calcium carbonate in the form of nodules or as a layer. This occurs when the carbonate precipitates out from capillary action of soil or water moving downward. The ‘C’ horizon is the next deeper horizon. It mainly consists of parent rock in weathered form. The lowermost horizon is the ‘D’ horizon and represents the parent rock in intact form. In some places some horizons are absent.
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44
Rotary drilling and blasting in large surface mines
Tundra Profile
A - Peat
Terra Rossa Profile A - Red Clay
O (Contains organic matter) A (Rich in humus)
B - Clay
B - Limestone
B (Rich in Mineral salts)
C (Subsoil and weathered rock)
Laterite Profile
Desert Profile A - Sand (No Dunes) A - Clay
B - Concentration Rich in salt D (Solid rock)
C - Leached Zone D D - Solid Rock
Figure 3.5 Soil profile under different environments.
3.6
OCCURRENCE OF MINERALS
A rock mass is formed by the assemblage of several minerals. Some of these minerals are of great economic importance because they are useful in many ways e.g. for extraction of metals, for use in different manufacturing processes, for medical use, for ornamental use etc. Mineral deposit is a term used for an assemblage of minerals where one or more, but not all, the minerals in the assemblage are of value. When a granite formation is used for production of tiles for house flooring, kitchen platform etc., it is not called a mineral deposit since it does not involve any specific mineral. But when a granite formation is excavated for obtaining pyrite contained in it, the granite formation is called mineral deposit. An ore body is the specific portion in a mineral deposit that fully or largely contains the mineral of economic importance. The remaining portion of the mineral deposit is considered as waste and is usually called overburden because it lies above the ore body. It has to be removed to gain access to the ore body. When an ore body contained in a mineral deposit has to be accessed it becomes essential to break the overburden or the surrounding rock mass and move it away. The ore itself has to be crushed before subjecting it to further chemical processes. All such activities together are called mining. When the rock mass is excavated as a whole, the activity is called quarrying. Examples of this are the rock mass excavation in broken form e.g. large stone pieces
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Properties of rocks
45
or crushed stone required in the construction industry or in dimensional block form for cutting tiles for house flooring etc. In nature, mineral deposits occur in many different forms. During the formation of igneous rocks, the magma often contains small quantities of minerals of copper, lead, zinc etc. These minerals do not find a place in the main rock mass that contains quartz, feldspar, mica etc. Therefore, these minerals get separated into ore bodies. Water superheated to very high temperatures is very active and can dissolve many economically important minerals to form a hydrothermal fluid. When an igneous rock mass forms, the hydrothermal fluid usually accumulates at the top of the rock mass as it has low density. After a certain time interval the water itself gets evaporated and the compounds crystallize to form a mineral deposit. On many occasions, the hydrothermal fluid mentioned above enters and spreads through cavities such as joints, fissures etc. This gives rise to vein minerals in the form of large crystals. Tourmaline, topaz, cassiterite etc. are some of the important vein minerals. In certain cases the hydrothermal fluid intrudes the existing rock bodies and forms a mineral deposit called a replacement mineral deposit. Minerals of lead and zinc are found in such deposits. The process of magmatic differentiation during the formation of intrusive and extrusive igneous bodies gives rise to pyritic deposits containing iron and sulfur. With the magma intrusion, hot gaseous substances flow in the magma mass. As the magma cools these gaseous substances deposit certain minerals, which till then are in a volatile state. The process of formation of such deposits is called pneumatolysis. The most important mineral contained in such deposits is Wolframite. At times, due to different degrees of metamorphism, not all the minerals change their form. In the process some minerals become liquid in their state and easily flow away. What remains is a skeleton of fibrous silicates commonly known as Asbestos. In certain instances, the formations in which metamorphism takes place involve carbon in the form of coal or carbonaceous gases. These get converted into graphite or diamonds. Mineral deposits are also associated with sedimentary rocks. The flow of water in a river may carry minerals of low specific gravity due to the limited velocity of water. Such minerals are mostly silicates. What remains in their place is a mineral mass of high specific gravity. These are called placer deposits. Gold, platinum, antimony etc. are associated with such placer deposits. Sea water contains many dissolved minerals. The proportion of these is about 31 parts per thousand. When sea water enters certain depressions at the time of high tide, it may remain entrapped in the depression for a long time. Evaporation of the water leads to formation of a thin layer of mineral deposit. When the process of such evaporation gets repeated with the next equal high tide the thickness of mineral deposit grows and eventually the deposit acquires a thickness of considerable magnitude. Such deposits are called evaporites. Potash, magnesium, gypsum etc. are associated with such deposits. Laterite is also a form of sedimentary mineral deposit. Laterite deposits are formed in tropical regions in the rainy season. Sediments formed are transported to flat surfaces where they penetrate into the ground. In the dry season some minerals
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Rotary drilling and blasting in large surface mines
containing the leached ions are drawn to the surface of the ground by capillary action. In the next rainy season these minerals are washed away and fresh sediments are redeposited. In this manner higher density minerals remain deep seated and form a large mineral mass. Iron ore or bauxite are the minerals found in such deposits.
3.7
ROCK SPECIMEN PROPERTIES
The objective of blasthole drilling is to drill a hole in the rock at the fastest possible penetration rate, while giving maximum life to the blasthole drill and components like drill bits, drillstring accessories etc. The progress of a blasthole is within a very small cylindrical path in a rock mass. Therefore, when it comes to blasthole drilling, the characteristics of the whole rock mass are less important as compared to the properties of a rock sample or specimen. In the rotary blasthole drilling process, the intact rock at the bottom of the hole is crushed or scraped and chips formed are blown out along with circulating compressed air. If any discontinuities, like joint, fracture etc. are perpendicular to the blasthole alignment they usually help the blasthole drilling process and result in faster penetration rates. However, if they are at an angle with the blasthole alignment and if the strength properties of the rock on two sides of the cracks or joints are far different, the blasthole can deviate from its intended alignment to an unacceptable level. From the standpoint of blasthole drilling, the most important property of intact rock is drillability, and from the viewpoint of blasting it is blastability. Drillability is a complex combination of several factors amongst which engineering properties of the rock specimen play a very important role. Some of these are density, compressive strength, tensile strength, hardness, toughness, brittleness, coefficient of internal friction, and abrasiveness. Drillability is treated in separate chapters in this book but the other engineering properties are briefly dealt with in the following subsections. Since intact rock specimens from different depths are usually obtained in the forms of cores in diamond core drilling, these cylindrical cores are used for most of the tests in a rock testing laboratory. The cores are carefully prepared before testing so that they have appropriate length and diameter. The ends of the test specimens are also carefully cut so the circular surface is exactly perpendicular to the core axis and is in a plane so that the specimen is subjected to uniform load and not point loads. When hard rock cores are to be shaped, diamond tools are used for the shaping. To prevent diamonds from burning, water is used as a coolant. The specimens so prepared are dried before testing. For shaping cores of soft rock that contain clay and other constituents that may be affected by water, tungsten carbide tools are used. Such tools do not require water as a coolant. In order to get truly a representative value of the property, many tests are carried out and the mean value of the property is taken as the correct representation. A few formulae are given in the following elaboration. As these formulae are only for the purpose of understanding the relation of various quantities, no units of measurements have been given for the quantities. In any case these formulae are unified and will give desired results when consistent units for all the terms and appropriate values of the constants in the formulae are used.
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Properties of rocks
3.7.1
47
Density
Density of rocks, usually measured in terms of kg/m3 or lb/ft3, is not so important from the viewpoint of formation fracturing, but is of great importance when it comes to flushing. As per Stoke’s law, discussed in one of the later chapters, chips of the rocks with low density, such as shale, coal etc., are bailed out from the blasthole very easily but chips of the rocks with high density such as iron ore, uranium ore etc., are bailed out with relatively great difficulty. Fortunately, the density of rocks is usually within a small range. When the bailing velocity of compressed air is properly selected by use of an appropriate compressor, the penetration rate does not reduce on account of the density of rocks. Densities of many different rocks are given in one of the appendices at the end of the book.
3.7.2
Compressive strength
In plain words, compressive strength means ability to resist compression. When a cylindrical rock sample is subjected to increasing axial compressive force as shown in Figure 3.6, its axial length goes on reducing and at the same time its diameter goes on increasing. Stress exerted on rock sample is calculated as σ = F/A = F/(π ∗ d2/4) The strain that results from the test is calculated as ε = ΔL/L Explanation of the symbols F, d, L is contained in Figure 3.6 itself. At certain values of compressive stress the rock sample fails. Thereafter the compressive strain keeps on increasing even when the axial load is not increased, or in fact even when it is decreased, resulting in unchanged or reduced compressive stress. Eventually, the rock sample crumbles into pieces. The maximum value of stress to which the rock specimen is subjected in such a test process, is called compressive strength. In most of such tests
ΔL
Compression Force F
Original Shape of the Rock Sample is in Gray Color
Top Plate Averaged Shape After Compression L Likely Failure Plane ΔL1
Bottom Plate
Figure 3.6 Uniaxial compression test.
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Rotary drilling and blasting in large surface mines
the rock sample is subjected to compressive pressure only in the axial direction and no pressure is applied on its cylindrical side except the atmospheric pressure. Therefore, the strength is termed as uniaxial compressive strength and abbreviated as UCS. Sometimes UCS is also taken as an acronym of unconfined compressive strength. UCS is measured in the laboratory in a machine that exerts force on the material sample as shown in Figure 3.7. The machine is called a Universal Testing Machine. It consists of special servo-controlled mechanisms for measurements. Rock samples used for the test are usually cylindrical in shape and the circular surfaces of the sample are very carefully polished to avoid point loading. The force exerted on the sample is increased at a certain rate and the material sample eventually fails. Brittle materials, like rock, get crushed into pieces. The compressive strength of the material is calculated by dividing the breaking force by the cross sectional area of the sample. UCS of a dry rock sample is always more than that of a wet rock sample. For some sedimentary rocks the difference is very high. The commonly adopted procedure for determination of UCS is as per American Standard ASTM D2938. Stress-strain curves plotted from the observations of unconfined compression tests on samples of quartzite, hornfels and marble are shown in Figure 3.8. Rock is said to be elastic when the strain is linearly proportional to stress, whereas when the relationship is non linear the rock is called plastic. Quartzite shows perfectly elastic behavior throughout up to a point B, and thereafter shows slightly plastic behavior up to the ultimate compressive strength at point C.
Figure 3.7 Rock core sample being subjected to UCS test.
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Properties of rocks
300 B
Stress in MPa
250 200
B1
49
C Quartzite
C1 Hornfels
150 100
2
C2
50
3
Marble
0
10
20 30 40 Strain in mm ×106
50
60
Figure 3.8 Stress strain curves for some rocks.
Since the length of the curve OB is several times longer than the length BC and even the zone BC is only slightly plastic, it is safe to say that quartzite is an elastic rock. Hornfels shows perfectly elastic behavior up to point B1 and thereafter significantly plastic behavior up to the ultimate compressive strength at point C1. Hornfels is considered an elastoplastic rock because the length of plastic zone B1C1 is of significant magnitude as compared to elastic zone OB1 and the plasticity in zone B1C1 is also of considerable magnitude. In the case of marble there is no elastic zone at all. Marble, therefore, is considered a plastic rock. When rock is subjected to compression, failure occurs due to shear stress at an inclined plane as shown in Figure 3.6. Some important properties of rock determined from the readings of uniaxial compression tests are the ‘Modulus of Elasticity’, ‘Modulus of Plasticity’ and ‘Poisson’s Ratio’. If rock shows elastic behavior then the modulus of elasticity within elastic zone is given by the equation E = (σ1 − 0)/(ΔL1 − 0) Modulus of plasticity in the plastic zone is calculated by the equation EPL = (σcom − σ1)/(ΔLcom − ΔL1) The meaning of the symbols used in the above two equation can be easily understood from Figure 3.9. Poisson’s ratio is the ratio of lateral strain to axial strain. In case of Figure 3.6 it equals μ = ΔL1/ΔL
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Rotary drilling and blasting in large surface mines
Stress in MPa
50
C1
σcom B1
σ1
0 0
ΔL1
Strain
ΔLcom
Figure 3.9 Stress strain curves for elastoplastic rocks.
In all the compression or tension tests carried on the rocks, the rate at which stress is applied has a great bearing on the ultimate compressive or tensile strength. All rocks show very high strength when the compressive force is applied rapidly. UCS of rocks has a reasonably good correlation with point load index and shore hardness. Often rock specimens (not rock masses) are classified on the basis of their UCS as indicated by Table 3.11. UCS of some commonly occurring rocks is given in Table 3.12 and also in Appendix 14. A special test, called triaxial compression test, is also sometimes carried out on soft rock samples. In such tests the rock sample is subjected to pressure on the cylindrical side of the sample. The value of compressive strength obtained in triaxial compression tests is higher than the UCS. The triaxial compression test has a far better resemblance of in-situ behavior of the rock, because in actual practice a portion of rock in the alignment of the blasthole is always surrounded by an adjacent portion of rock mass. Such rock mass invariably exerts horizontal pressure on the portion of rock in blasthole alignment. The commonly followed procedure for determination of CS by triaxial compression test is as per American Standard ASTM D2664.
3.7.3 Tensile strength The concept of tensile strength is almost identical to that of compressive strength, but the direction of the axial force acting upon the sample is exactly opposite. Here, as the sample is subjected to increasing tensile force, the diameter of the rock sample goes on reducing and the axial length goes on increasing. Tensile stress and tensile strain are observed in the tests in the same manner as those in compressive stress and strain. At a certain value of tensile stress, the rock sample fails. At this point the rock sample usually breaks apart in the case of elastic rocks. In some rocks, that show high plasticity, the tensile strain keeps on increasing even when the axial force is not increased. The value of the stress at such point is called tensile strength.
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Properties of rocks
51
Table 3.11 Rock classification on the basis of unconfined compressive strength. Uniaxial compressive strength in MPa
Point load Schmidt index in hardness by MPa L – hammer
Method of strength in the field
R5 – Extremely >250 Strong
>10
50–60
R4 – Very Strong
100–250
4–10
40–50
R3 – Strong
50–100
2–4
30–40
R2 – Medium Strong
25–50
1–2
15–30
R1 – Weak
5–25
–
501 kN)
Today, in 2010, rotary blasthole drills are being made by many companies spread over many countries. A broad idea of their product range is shown by Table 4.1. In the late 1970s the list of manufacturers of rotary blasthole drills included Atlas Copco, Bucyrus Erie, Chicago Pneumatics, Driltech, Gardner Denver, Hausherr, Ingersoll Rand, Joy Robbins, Marion, Reedrill, Rudgormash and Schramm. The manufacture of rotary blasthole drills is not a very lucrative business. Many companies built rotary blasthole drills in the past, but very limited demand for the product has forced them either to close or sell the manufacturing rights and lose their identity. Bucyrus kept their leadership in large rotary blasthole drills virtually till the end of the last millennium, through acquisition of Marion. In the meanwhile in 2003 Atlas Copco, well known all over the world for their top hammer and down-the-hole hammer drills acquired and amalgamated the Ingersoll Rand drilling equipment line with their percussion and DTH drill line. With this move they seem to have overtaken Bucyrus. However, Bucyrus have recently
Table 4.1 Rotary blasthole drill manufacturers and their range. Manufacturing range
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Manufacturer
Country of mfg.
Small
Medium
Large
E. large
Atlas Copco Bucyrus Intl. Changsha Keda CMCL Emico Elecon Globe Drill Hausherr Heavy Eng. Corp. KLR LMP Precision NKMZ P&H Mining Revathi Equipment Rudgormash Sandvik Driltech
USA USA China India Australia Germany India India India USSR USA India USSR USA
Yes Yes Yes Yes
Yes Yes Yes Yes
Yes Yes Yes
Yes Yes Yes
Yes
Yes Yes Yes Yes Yes
Yes
Yes Yes
Yes Yes
Yes Yes Yes
Yes Yes Yes Yes Yes
Yes
Yes
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Rotary drilling and blasting in large surface mines
countered this move by acquiring Terex Mining who, about three years ago had acquired and amalgamated Reedrill. Whether Bucyrus regain their leadership is yet to be seen. Other major players in large rotary blasthole drills are P&H Mining with inheritance from Robbins and Gardner Denver, and Sandvik with inheritance from Driltech. Very heavy machines, weighing more than 150,000 kg and capable of exerting bit loads more than 550 kN are now being manufactured by Bucyrus International, P&H Mining and Atlas Copco. Some of them can drill 20 m deep holes of diameters in excess of 406 mm in one single pass. The largest amongst these is the Bucyrus International 59HR that weighs up to 193,000 kg, can exert a bit load up to 734 kN and is rated for drilling blastholes of 445 mm diameter. It is shown in Figure 4.18. Most of the manufacturers mentioned above are of US origin, but the Russians had also independently developed electric blasthole drills for drilling blastholes of 320 mm diameter to 32 m depth some forty years back. They are often found in East European and some other Asian countries. One of the deep-rooted ideas in the minds of blasthole drill designers was a drill with twin mast, meant to drill two blastholes simultaneously. Marion Power Shovel Company developed one such drill and put it to field trials at Squaw Creek Mine. It did work and produced results but not to the level of satisfaction that the customer could order more of its kind. The drill is shown in Figure 4.19. All through the history of rotary blasthole drilling, the objective in improving machinery has been to introduce safety, efficiency, reduced maintenance and automation. The earliest step towards this was pipe handling. Carousel pipe changers were used even in the early 1950s. Single pipe changers, developed in the 1960s, were better at handling long and heavy drill pipes. Drilling recorders were introduced in blasthole drills in the early 1960s. In the early 1980s they became electronic. Automatic fire extinguishing systems and automatic lubricating systems were incorporated into blasthole drills in the late 1960s. In 1971 blasthole drills entered an era of automation, when Bucyrus Erie adapted a primary logic circuit to control some of the drilling operations on their blasthole drill model 60R. Early electric drills featured Ward Leonard controls for accurately controlling the DC electric motors in the rotary head. Some manufacturers adopted AC motors and frequency variation to control the motors in the early 1980s. However, advances in Static or Thyristor Control for DC motors have still retained DC motors on the rotary heads of blasthole drills. In the 1990s integrated circuits revolutionized the blasthole drills to a very great extent. They made it possible to precisely measure and monitor several functions in the drilling operation. Some features offered by such systems were useful in the following manner. 1
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With the drilling recorder the driller could record the values of as many as 6 parameters of the drilling activities. With this facility the operators could moni-
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Brief history of rotary blasthole drilling
2
79
tor the drilling process very effectively. Later, when digital systems were introduced the number of parameters went to as much as 12 and their values could be seen on the screen in the operator’s cab. Blasthole drills could be automatically leveled within an accuracy of 0.25°.
Figure 4.18 The largest, Bucyrus International’s 59HR, 2003.
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Rotary drilling and blasting in large surface mines
Figure 4.19 Marion twin mast rotary blasthole drill, 1986.
3
4
5
With specialized drill process software, stored in the computerized system of the blasthole drill, it became possible to control drilling operations in the most appropriate manner by automatically adjusting feed force, feed rate, rotary speed, compressed air supply and so on. Radio control for propelling operations of the blasthole drills made it possible to move the drill to the next blasthole by an operator who stood on the ground near the new hole position rather than the operators cab of the drill at the position of the already drilled hole. Very high positioning accuracy was achieved through such remote control propelling. With the introduction of Global Positioning Systems (GPS) it became possible not only to navigate the blasthole drill even more accurately, but also to integrate the drilling operation with a comprehensive mine operation system to increase productivity.
Technological advances in the new millennium have pushed automation further by introducing GPS. More details on these are given later in a separate chapter that deals with computers in rotary drilling. 4.5 TRUCK MOUNTED ROTARY BLASTHOLE DRILLS For almost a decade after the first world war, self propelled cable tool drilling rigs were being manufactured by manufacturers like Keystone, Star, and Armstrong. However when trucks became easily available in the early 1930s, many drill manufacturers started making their drills on a frame which could be simply mounted at the time of delivery onto the truck chassis purchased from the truck manufacturers.
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Figure 4.20 Sandvik truck mounted rotary blasthole drill, 2008.
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Rotary drilling and blasting in large surface mines
This move was beneficial to both manufacturers and purchasers. Drill manufacturers because it considerably reduced their inventory, and purchasers because it enabled them to get spare parts for half of their machine far more quickly from the truck manufacturers. The choice of truck-mounted rotary blasthole drills for the main mining operations is made very rarely for several reasons, explained in one of the next chapters, but many mining companies often keep one rotary blasthole drill in their fleet of equipment because: 1 2 3
It can be taken from one bench to another in much less time. It can be used for drilling deep holes often required in mine benches. If the main crawler blasthole drill fails and needs a long time for repairs, the truck-mounted blasthole drill can temporarily carry out the blasthole drilling work as a stopgap arrangement.
In the early 1970s many manufacturers made truck-mounted rotary blasthole drills, but currently only a few of them have continued their manufacture. Truck-mounted rotary blasthole drills are now available from Atlas Copco, Sandvik, Schramm. Most of the truck-mounted rotary blasthole drills can handle four to six drill pipes of length 7.6 m and diameters up to 127 mm, and are suitable for drilling 200 mm blastholes. One such drill is shown in Figure 4.20. Their bit load capability is of the order of 133 kN. Larger truck mounted drills, capable of drilling blastholes of diameters up to 311 mm with 219 mm drill pipes of length 12.2 m each, are also available. These drills have bit loading capacity of 311kN.
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Chapter 5
Rotary blasthole drilling bits
5.1
INTRODUCTION
The drill string used in rotary drilling comprises many items, but the most important component in it is the drill bit. The drill bit disintegrates the formation under its teeth or cutting edges, and enables progress of the blasthole in the intended direction. In this process the bit undergoes very heavy wear and the penetration rate eventually drops down below an acceptable level. Thereafter the bit is required to be replaced. As a drill bit is an expensive and critical item, it is thoroughly dealt with in this separate chapter. Other accessories used in a rotary drill string are elaborated in chapter 6. Two types of drilling bits, viz. drag bits and tricone bits are used in blasthole drilling.
5.2
DRAG BITS
In the early days, drag bits made from cast steel were used for drilling blastholes in soft and very soft formations. Ever since tungsten carbide inserted drag bits made entry into the field of blasthole drilling, the use of cast steel bits stands almost discarded. Today, tungsten carbide inserted drag bits are used everywhere but in the near future drag bits, made with thermally stable diamond compacts, may change the whole scenario of rotary blasthole drilling. Two types of tungsten carbide inserted drag bits, viz. blade bits and claw bits, are currently used.
5.2.1
Blade bits
Blade bits have two types of cutting edges viz. Chevron Type and Step Type. A Bit with chevron cutting edge is shown in Figure 5.1. Such a cutting edge is usually found on small diameter bits used for drilling blastholes in underground mines. The step type cutting edge, shown by line drawing in Figure 5.2, is adopted for larger diameter bits that are used for drilling blastholes in surface mining operations. Step type drag bits are less prone to blasthole deviation. Blade bits have two, three or four wings. At their cutting edges they have flat, blade-like tungsten carbide inserts brazed to the main cast steel body.
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Figure 5.1 Chevron blade type drag bit.
Figure 5.2 Step blade type drag bit.
Experiments at various research centers have indicated that the shear strength of the material used for bonding the carbide inserts to the steel body of the blade bit should be more than 240 MPa for the bit to work satisfactorily. If blade type bits are used with air flushing, they attain high temperatures, and the possibility of losing a tungsten carbide insert increases considerably due to weakening of the brazing material that joins the carbide insert to the bit body. Therefore, blade type bits are either used with water circulation or water is injected in the stream of compressed air used for flushing.
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Rotary blasthole drilling bits
85
Since most of the soft formations, such as alluvial overburden in coal mines, coal seams, potash deposits etc., that require blasthole drilling are non-abrasive, the body of the bit remains intact for a long time. In such a situation drag bits with replaceable blades, as shown in Figure 5.3, can be used with advantage. Details of some blade type drag bits used in blasthole drilling are given in Table 5.1. More data can be obtained from the manufacturers.
5.2.2
Claw bits
A typical claw bit is shown in Figure 5.4. Depending upon the bit diameter, it has five or more replaceable teeth and a replaceable pilot. All the teeth are fixed in a forged, machined and heat-treated bit body by means of steel pins or circlips. Since the inserts in the claw bits are not brazed, the bit can be used with air circulation along with water injection. The performance of claw type drag bits is better, as penetration of the teeth of the bit in the soft formation is deeper. For this reason, claw bits require higher torque than the blade bits. Claw bits are chosen for drilling in formations having UCS ranging from 35–70 MPa. In any case, while drilling blastholes with drag bits, a large quantity of water is required to be injected with the compressed air for bit cooling. Details of some claw type drag bits used in blasthole drilling are given in Table 5.2.
5.2.3
Feed force and rotation of drag bits
Due to its low importance, the process of drag bit drilling has not been researched or recorded as deeply as have tricone bits. Field practice in using drag bits suggests that blade bits should be used with a bit weight of 100–200 N per mm bit diameter and rotated at 20–50 rpm, whereas claw bits should be subjected to a feed force of 3000–5000 N per tooth of the bit and rotated at 10–40 rpm.
Connector
Reaming Stabilizer Holder
Bit Shank
Blades
Figure 5.3 Replaceable blade drag bit.
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Table 5.1 Details of some blade type drag bits. Nominal bit diameter mm
Inches
Connection
48–64 70 73 70–89 92–120 127 130 133–140 143 152–171 178–203 178–311
1.7/8–2.1/2 2.3/4 2.7/8 2.3/4–3.1/2 3.5/8–4.3/4 5 5.1/8 5.1/4–5.1/2 5.1/4–5.1/2 6–6.3/4 7–8 7–12.1/4
AW or A Rod AW or A Rod AW or A Rod NW or N Rod 2.3/8" API Reg. 2.3/8" API Reg. 2.7/8" API Reg. 2.7/8" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 4.1/2" API Reg.
Configuration – (chev./step) and number of wings C3 C3 C3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, S3 C3 S3
Figure 5.4 Claw type drag bit.
5.3 TRICONE BITS Today, only about 2% of rotary blasthole drilling is carried out with drag bits. The remaining 98% is carried out with tricone bits. Tricone bits with meshing cones were introduced in oilwell drilling practice in 1933. Initially tricone bits were used only for oilwell drilling. When bits with a facility for compressed air circulation through their bearings were developed in 1949,
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Table 5.2 Details of some claw type drag bits. Nominal bit diameter mm
Inches
Connection
Number of teeth
98–101 105–108 114–120 127–130 140–149 152–165 178–190 203–216 250 270 280 311
3 7/8–4 4 1/8–4¼ 4½–4¾ 5–5 1/8 5 ½–5 7/8 6–6½ 7–7½ 8–8½ 9 7/8 10 5/8 11 12
2.3/8" API Reg. 2.3/8" API Reg. 2.3/8" API Reg. 2.3/8" API Reg. 2.7/8" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 4.1/2" API Reg. 4.1/2" API Reg. 6.5/8" API Reg. 6.5/8" API Reg. 6.5/8" API Reg.
4 4 4 4 5 6 6–7 8 9–10 10–12 12 14
Pilot bit diameter mm 40 40 40 40 40 40 50 50 50 50 50 50
tricone bits were also introduced into blasthole drilling practice. Tungsten carbide inserted bits were initiated in 1951, and sealed bearing bits made their appearance in 1959. Sealed journal bearing bits were launched in 1968.
5.3.1 Types and nomenclature Drag bits, as shown in Figure 5.1 to 5.4, do not have any moving parts. Tricone bits, however, are assembled by using many parts, some of which move while the drilling operation goes on. The reasons for having three cones on a bit are: 1
2
The contact between the cone of a bit and the horizontal bottom of the blasthole is always linear. In two-cone bits these two lines are actually collinear. In threecone bits the three lines are radial. Therefore, a blasthole drilled with two-cone bits is far more prone to deviation. Tricone bits exert the bit weight in three vertical planes in a far better balanced manner and reduce the hole deviation to a great extent. The balanced bit load distribution also gives longer life to a tricone bit as compared to a two-cone bit. Bits with more than three cones can give better weight distribution but in such case the size of the cones has to be reduced and so the bearings within them have to be smaller. This leads to very weak bits that are susceptible to early bearing failure.
Primarily two types or tricone bits are made, viz. milled tooth bits and tungsten carbide insert bits. Milled tooth bits are also called steel tooth bits. In milled tooth bits the cones are completely made from alloy steel and the teeth are formed by milling – a mechanical metal cutting process.
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Tungsten carbide inserted bits have different types of carbide inserts firmly fixed into the holes drilled in the cones by another manufacturing technique called shrink fitting. Since tungsten carbide is a much harder and abrasion-resistant material, the carbide inserted bits give far better performance – often 3 to 9 times – than that of a comparable milled tooth bit operating under similar conditions. Milled tooth and tungsten carbide insert tricone bits are shown in Figure 5.5. Tungsten carbide inserted bits are more expensive than milled tooth bits due to the higher price of tungsten carbide inserts and added manufacturing costs. Some manufacturers make composite bits where the cones have steel teeth on all the portion except the outer edges, where tungsten carbide inserts are fitted. The design principles of a milled tooth bit also apply to a tungsten carbide insert bit. Three cones of a milled tooth tricone bit, when seen from bottom of the bit, look like those shown in left side picture in Figure 5.6. No. 1 cone is the cone which has a spear point i.e. the teeth at the apex of the cones. The next cone, i.e. cone no. 2, has teeth on a circle at a small distance from the apex of the cone. The row of these teeth is called the nose row. The nose row on cone no. 3 is at a slightly larger distance from its apex compared to cone no. 2. The diameter of a tricone bit is equal to the inner diameter of the smallest circular ring through which the bit can pass. This dimension is treated as the nominal hole size of a new bit. Even though used for blasthole drilling, tricone bits are made to the standards laid down by the American Petroleum Institute. Acceptable tolerances in the diameter of tricone bits as per API Standards are given in Table 5.3.
Figure 5.5 Milled tooth and tungsten carbide insert tricone bits.
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Rotary blasthole drilling bits
Gage Teeth
Bit Leg
Cone No.2
Air Passage Air Exit
Nose Row
Intermediate Row
Ball Bearing
Nose of Cone 2
Spear point
89
Nose of Cone 3 Cone Shell
Intermediate Row Nose Row Gage Row Roller Bearing
Gage Teeth
Gage Teeth Cone No.3
Cone No.1
Gage Insert Bush Bearing
Thrust Button
Figure 5.6 Nomenclature of tricone bits. Table 5.3 Acceptable tolerances in tricone bit diameter. Bit diameter mm
Higher size tolerance
Lower size tolerance
85.7–349.2 mm 355.6–444.5 mm Larger than 445 mm
0.79 mm 1.59 mm 2.38 mm
0 mm 0 mm 0 mm
In the early days, tricone rock bits of diameters as low as 100 mm were made but, since it is now possible to drill holes of diameter up to 150 mm by rotary percussion drilling at faster penetration rates, use of such small size tricone bits has remained restricted to soil investigation and some other purposes, where drilling is to be carried out with least disturbance to the rock mass. In any case, small diameter tricone bits do not perform well in hard formations, because small tricone bits have very small bearings. Further it is also not possible to give adequate feed force on the tricone bits through slender drill rods as they are susceptible to buckling. The cross section of a roller and corresponding leg is shown on the right-hand picture in Figure 5.6. To enable the roller to roll over the formation as the drill string is rotated, bearings are provided between the cones and respective legs. Usually the innermost bearing is a journal type. It involves no moving parts, and therefore can transfer very heavy radial loads. The middle bearing is ball type. It interlocks the cone with the leg. The outermost bearing is roller type. It allows the bit to withstand very heavy downward thrust while being rotated at low frictional resistance. Figure 5.6 shows three rows of teeth on the cone and three rows of bearings. These are applicable to bits of small diameter. Sometimes more than three rows of teeth as well as bearings are provided on large diameter bits. Components of a tricone bit are made separately and are assembled later. Each of the three legs of a tricone bit is made from a forging of an alloy steel. The forged leg is very precisely machined for the dimensions of grooves for roller, ball and
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journal bearings. Once the machining process has finished, the machined portion is heat treated to the desired hardness by carburizing. Figure 5.7 shows a tricone bit leg. The cones are made from their respective forging, then machined and heat treated. After this, if applicable, carbide inserts are shrink-fitted in the holes. Journal and rollers of the bearing are then inserted in the appropriate grooves of the cone. Once this is done, the leg is inserted in the cone in which the roller bearings are held in their final position. The balls of the ball bearings are inserted from outside through a hole provided for the purpose.
Figure 5.7 A tricone bit leg.
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Assembling of one leg is complete once the balls are appropriately inserted and the hole is plugged by a pin that is later welded to the leg. The three legs of a bit are then held together firmly and their sides are welded to form a tricone bit. After the welding and stress-relieving operations are over, the top portion of the bit is machined to form an API Regular pin connection. This connection enables coupling the bit to the stabilizers or bit subs which are the next bottommost components of a drill string. The names of many other portions of a tricone rock bit are shown in Figures 5.8 and 5.9.
5.3.2
Geometrical design aspects
Many factors relating to the geometry of the components of a tricone bit, such as cone size, cone offset, size and shape of teeth etc., play a very important role in its field performance. While designing a bit, all these must be given due consideration. 5.3.2.1
Fluid circulation
Aspects related to removal of cuttings from a blasthole by circulating compressed air have been dealt with in a separate chapter of this book. Apart from the need to bring the drill cuttings to ground surface, the compressed air is also required to dissipate the heat generated by friction between cones and the formation, as well as within the bearings located inside the cones of a tricone bit. If bits with bearings lubricated in the factory and meant to be used with water circulation are used in blasthole drilling, the compressed air proves insufficient in heat dissipation, and consequently the bearings fail soon after the lubricant melts and oozes out or burns and becomes ineffective. In 1949 this problem was solved by providing air passages like those shown in Figure 5.9. Another, more recent way, is to use a bit provided with a composite seal, made of elastomer plus metal, and a reservoir built into the bit leg to store a special lubricant. Such lifetime-lubricated bits can defy temperatures to even 400°C. 5.3.2.1.1
Air cooled tricone bits
Air-cooled bits are of two types viz. regular and jet circulation as shown in Figure 5.10. In a regular circulation tricone bit, three passages to the bearings of the three cones are provided for air circulation through the bearings and the main airflow is through a central hole. In a jet circulation tricone bit, three passages to the bearings of the three cones are also provided for air circulation through the bearings, but the main airflow is also through three jet nozzles provided on the three legs of the bit. As a general norm, it is desired that at least about 30% of the total compressed air supplied to the bit should flow through bit bearings so the bearings cool sufficiently. The size of the air passages to bearings is given appropriate dimensions to achieve this objective. In most of the regular circulation bits, this objective is achieved when compressed air pressure at the bit is about 280 to 300 kPa. In jet circulation bits, three independently replaceable nozzles are provided to achieve this. When nozzles of smaller
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Nozzle Retaining Pin
Bit Type Emboss
Bit Serial Number Emboss
Shank
Bit Size Emboss
Shank
Nozzle
Bit Assembly Number Emboss One of the Three Legs
Shirttail Hardfacing
Gage Teeth
Figure 5.8 Nomenclature of tricone bits.
Cone Gage Surface
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Water Separator
Snap Ring
Air Screen Bit Leg
Rubber Flapper Screen Tube Assembly
Main Air Passage Air Passage to Ball Bearing Air Passage to Pilot Bearing Pilot Pin Shirttail Inserts
Cone Gage Inserts
Leg Journal
Cutting Surface Inserts Cone
Roller Bearing Ball Bearing Bush Bearing
Figure 5.9 Air cooling of blasthole bits.
Regular Circulation
Jet Circulation
Air Flow through Central Hole
Air Passage to Bearings
Air Flow through Jet
Air Passage to Bearings
Jet Nozzle
Figure 5.10 Regular and jet circulation bits.
central bore are chosen, a lower volume of compressed air flows through the nozzles and consequently a higher volume of compressed air is forced to flow through the air passages of the bearings on account of higher pressure built up behind the nozzles. Blasthole drills often have provisions for injecting water into the flow of compressed air. This injected water flows to the drill bits. The purpose of such water circulation is
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to suppress the dust formed during formation fracture. This water, in the form of fog, also passes through the bearings of the bit and helps in cooling the bearings. Many blasthole drills are provided with filter type dust collection systems instead of dust suppression by water injection. In such blasthole drills, if an oil injection system is provided instead of water injection, the oil mixed with compressed air reaches the bearings of the drill bit and lubricates them. This type of bit lubrication proves far more effective and the life of the bit increases by as much as 20%. Air circulation tricone bits are usually provided with a back flow valve at the top, so sudden flow of water in the bit is prevented. 5.3.2.1.2
Sealed bearing bits
Originally these bits were developed for drilling geothermal wells where they had to work in high temperature surroundings. Now they are also used in blasthole drilling applications. The cross section of a leg of a sealed bearing bit looks like the one shown in Figure 5.11. The lubricant reservoir stores an adequate quantity of lubricant and supplies it to the bearings in the cone. The bellow type reservoir keeps the lubricant under pressure,
Sealed journal bearing
Diaphragm Reservoir Cap Grease Reservoir Silver Plates Beryllium Copper Thrust Face Radial Elastomer Seal Journal Hard Metal Gauge Gauge Protection Friction Pin Cone Tungsten Carbide Inserts
Figure 5.11 Cross section of a sealed bearing bit.
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so the lubricant flows to the bearings. The lubricant is prevented from escaping out of the cone by a composite seal of elastomer and metal. Sealed bearing bits have given considerably better performance in terms of drilled meterage and long life as compared to air circulation bits. 5.3.2.2
Size and shapes of teeth
A tricone bit, and the pattern of fractured formation made by one at the bottom of a hole, are shown in Figure 5.12. The indention made by the teeth can easily be observed in the pattern. Actually, the pattern shown is for a hole made in soft rock by mud flush drilling. In blasthole drilling, a bit is exerted with slightly higher feed force. The hole is flushed by compressed air. It removes larger cuttings from the hole bottom without any need for secondary crushing. Thus, in blasthole drilling, even deeper indentions than those shown in Figure 5.12 are formed at the hole bottom. Since in subsequent formation fracture the teeth have to touch at the crest in between two indentions, the number of teeth is often kept odd for the fracture to be more effective. Milled tooth bits have three distinct shapes for a tooth viz. Chisel, H and T. They all are shown in Figure 5.13. All these teeth shapes are formed by milling processes in the rock bit plant. When the angle between the two slanting surfaces of a tooth is small, tooth height is large and vice versa. For drilling in softer formations, the taller teeth give better performance because they can fragment larger cuttings of rock without getting blunt. For drilling in harder formations, the teeth are required to have lesser height since in such cases the tooth angle reduces and the teeth are less susceptible to breakage. The T and H type teeth are found only on the outermost, i.e. gage, rows where they give excellent resistance to higher wear. The teeth of milled tooth bits are usually given a hard facing on their sides for the purpose of extra wear resistance. In carbide inserted bits, the insert tips are made with different shapes during the sintering process. Some insert shapes are as shown in Figure 5.14. The apex angle between the sides of the surfaces near the tip differs considerably. Larger apex angle at the tip gives more strength and higher wear resistance to the insert; however, it reduces the penetration of the insert in the rock. Tricone bits meant for drilling in soft formations have inserts with low apex angle. Inserts at the cone gage and shirttail portion have a dome shape to give very high wear resistance because even in soft formations they merely rub against the formation. In extremely hard formations, carbide insert bits with synthetic diamond caps or diamond coating are chosen. 5.3.2.3
Cone size
The cones of the tricone bit are never true cones but a composite of two cones as shown in Figure 5.15. The main reason for choosing such a shape for the cone is that such a cone is bulged at the center and, therefore, it can accommodate much larger ball and roller bearings within it. This gives higher strength to the cone and the bearings, so much heavier bit loads can be exerted.
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Figure 5.12 Indention made by tricone bit.
When such a composite cone rolls along the formation surface, the teeth of the bit penetrate into the formation and cause indention. Neither the shapes of the teeth, nor the shapes of the indention caused by the teeth, have the shape of a perfect gear tooth. Therefore, slip is caused and side loads are exerted on the bit as well as on the formation. This causes large scale fracture of the formation.
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Rotary blasthole drilling bits
Top Surface
Side Surface
Side Surface
Chisel Type Tooth
H Type Gage Tooth Shirttail Hardfacing Gage Hardfacing Tooth Hardfacing
97
T Type Gage Tooth
Figure 5.13 Various shapes of tricone bit teeth.
Soft Formations
Soft to Medium Formations Medium to Hard Formations
Hard Formations
Figure 5.14 Various shapes of carbide inserts.
Deeper indention is desirable while drilling soft formations because it creates larger side loads and a larger portion of the formation gets fractured without causing any damage to the tooth. In the case of tricone bits to be used for drilling in hard formations, the outer cone is made bigger and the inner cone smaller, as shown in Figure 5.16. This enables accommodation of even larger bearings. The teeth of such bits are also made smaller so as to reduce the amount of slip when the cone rotates and the teeth penetrate. All this makes the bit suitable for drilling in hard formations where the strength of the cone has higher importance. In hard formations, the slip and consequent side force on the formation
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Envelope of Outer Cone
Cross Section of Actual Cone
Envelope of Inner Cone
Axis of Cone Rotation
Figure 5.15 Two cone composition of the cone of a tricone bit.
Envelope of Outer Cone
Cross Section of Actual Cone
Envelope of Inner Cone
Axis of Cone Rotation
Figure 5.16 Large outer and small inner cone envelopes of the cone of a tricone bit.
does not contribute much to the formation fracturing. Thus, a lower amount of slip is more desirable. A lower amount of slip also causes less wear on the tooth. From the above it can be easily surmised that a longer tooth will give more formation fracture but will be more susceptible to breakage and wear. If a bit meant for drilling soft formations is used for drilling hard formations, premature tooth breakage will occur. 5.3.2.4
Number of teeth
As the height of the teeth of a soft formation bit is large and that of hard formation bit is small, it is possible to accommodate more teeth on the cones of a hard formation bit. This can be observed in Figure 5.17, where cones of the same size have been
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shown. In harder formations, more teeth are necessary because the material removed by each tooth is relatively small, and to make full use of the bit strength in order to allow higher bit loads, more teeth are desirable. For the same reason, there are more inserts on a tungsten carbide inserted bit meant for drilling in hard formations. 5.3.2.5
Cone orientation
If a perfect cone is rolled over a smooth plane surface, it will roll through a perfect circle on that surface. The apex of the cone will lie at the center of such circle. In such rolling of the cone, there will be no slippage between the cone surface and the surface in the circle over which the cone rolls. Even a composite smooth cone (without any teeth) will not slip on a surface formed by two cones. In order to get more formation fracture and faster penetration rate in very soft formations such as coal, it is desirable to have larger amount of slip of the teeth on the cones. Such slip gives a much higher volume of fractured formation and consequently much faster penetration rate. There is always a limit imposed by the geometry on the height of the teeth. Thus, to achieve the higher amount of slip that is desirable in drilling soft formations, the cones of the bit are oriented with an offset as shown in Figure 5.18.
Many small height teeth can be accommodated in a bit meant for drilling in hard formation
Fewer tall height teeth can be accommodated in a bit meant for drilling in soft formation
Figure 5.17 More teeth on hard formation tricone bits.
Direction of Bit Rotation
Bit Rotation Axis Offset Angle
Axes of Cone Rotation
Figure 5.18 Offset of mounting of the cones.
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Offset can be defined as the perpendicular distance between the axis of rotation of the bit and the axis of rotation of the cone. When smooth cones with offset mounting are rotated on any surface profile they invariably slip. If the cones have teeth, the offset adds to the slip of the teeth and the formation is fractured with higher side forces. This results in a faster penetration rate. Even hard formation tricone bits also have some offset. Only the bits meant for drilling in extremely hard formations do not have any offset because normally such formations are very abrasive and providing offset in them creates a large magnitude of friction which results in very rapid tooth wear. 5.3.2.6
Top connections
Most of the developments of tricone bits were made long before the bits appeared on the blasthole drilling horizon. These developments were quickly built into the Standards set forth by the American Petroleum Institute. Air blast tricone bits are used not only by the blasthole drilling industry but also by the oilwell and waterwell drilling industries. It is, therefore, quite natural that tricone bit manufacturers have retained API Standard for air blast bits and have adapted API Regular type for the top-threaded connection of their bits. Tricone bits of different sizes have different API Regular pin connections as detailed in Table 5.4. The table also gives approximate weights of the bits. Some manufacturers do make blasthole bits with BECO connections which are more popular in the mining domain.
5.3.3
Metallurgy
Metallurgy plays a very important role in the manufacture of tricone rock bits. Important parts of a tricone bit, viz. the legs and cones, are made from forgings. After
Table 5.4 Top connections and weights of most commonly used for tricone blasthole bits. Nominal bit diameter
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mm
Inches
Connection
Weight of the bit in kg
143 152 159 171 187 200 229 251 279 311 349 381 406 445
5 5/8 6 6¼ 6¾ 7 3/8 7 7/8 9 9 7/8 11 12¼ 13¾ 15 16 17½
3.1/2″ API Regular Pin 3.1/2″ API Regular Pin 3.1/2″ API Regular Pin 3.1/2″ API Regular Pin 4.1/2″ API Regular Pin 4.1/2″ API Regular Pin 4.1/2″ API Regular Pin 6 5/8″ API Regular Pin 6 5/8″ API Regular Pin 6 5/8″ API Regular Pin 7 5/8″ API Regular Pin 8 5/8″ API Regular Pin 8 5/8″ API Regular Pin 8 5/8″ API Regular Pin
10 13.6 15.9 21.8 22.7 33 43 62 75 100 125 160 180 210
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machining, metallurgical process such as case hardening, stress relieving, shrink filling etc., have to be carried out in many stages. Even the tungsten carbide inserts are made by sintering – a metallurgical process. 5.3.3.1
Materials
The first step in the manufacture of a tricone rock bit is to select the right materials for various components of the bit. Since a tricone bit is subjected to very heavy stresses, the tolerance levels in the composition of the materials to be chosen for bit manufacture are very low. The materials usually used for milled tooth tricone bits are given in Table 5.5. The materials mentioned in the table are readily available, but in composition their tolerance limits are generally wider, and hence unacceptable for manufacture of a tricone bit component. Therefore, it is necessary to choose very specific materials from selected suppliers. Choosing appropriate inserts for tungsten carbide insert type bits is equally important. Detailed information on tungsten carbide is given in Appendix 7 at the end of this book. The following rudimentary knowledge in this regard is given for the sake of completeness. A tungsten carbide insert is formed by sintering a mixture of tungsten carbide and cobalt. Tungsten carbide is extremely hard and wear-resistant but at the same time very brittle. Cobalt, which serves as binder between tungsten carbide particles, is necessary to give the desired toughness. The proportion of tungsten carbide in an insert varies between 84% and 96%. A higher proportion of tungsten carbide is better suited for the manufacture of hard formations bits. Modern manufacturing techniques have made it possible to make even better tungsten carbide inserts. Two such techniques are as follows. Tungsten carbide (TC) inserts prior to the 1980s had the same composition throughout their volume. With the new technique it is possible to use a composition rich in tungsten for the outer side of the insert and a composition with larger proportion of cobalt in the central core. Sintering in this manner results in an insert with high wear resistance at the surface and the desired level of toughness throughout its Table 5.5 Materials used for manufacture of various components of tricone bits. Component name
Desired material properties
Material chosen
Material composition
Cones
Abrasion and Impact Resistance Weldability, High Impact Resistance, Surface Fatigue Resistance High Strength, Impact Resistance Wear Resistance Wear Resistance Wear Resistance Extreme Abrasion Resistance
ANSI 4817
C, Si, Mn, Ni, Mo, S, P
ANSI 8720
C, Si, Mn, Ni, Cr, Mo, S, P
ANSI S2
C, Si, Mn, Mo
ANSI 431 ANSI M2 Chrome Cobalt Tungsten Carbide
C, Si, Mn, Ni, Cr C, W, Cr, Mo,Va Co, Cr, C, W, Ni W, C, Co
Legs
Roller and Ball Bearings Journal Bushing Thrust Button Bearing Hardfacing Tooth Hardfacing
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volume. Often materials like silicon carbide, boron nitride etc. are also mixed in the outer layer in small proportions so the insert surface gets added wear resistance. Another manufacturing technique makes it possible to make tungsten carbide inserts with an outer cover of diamond. Because diamond is the hardest and most wear resistant material, diamond coated tungsten carbide inserts have extreme wear resistance. Modern bits usually use flat diamond compacts of 2 to 3 mm thickness on a base of tungsten carbide for gage wear resistance. When the formations to be drilled are extremely hard and abrasive all the inserts are diamond coated. During drilling operations the bearings of the tricone bits – particularly the bush bearings – generate a very high quantum of heat. This heat is released to atmosphere in two ways as follows: 1 2
Convection of heat through compressed air used for bit cooling. Conduction of heat from the bearings to the surfaces of cones and legs, from where it is again transferred by compressed air through convection.
While drilling in very hard formations with extremely heavy bit loads, the quantum of heat generated is so high that compressed air flowing through the bearings is not able to transfer the heat at the desired transfer rate. Three decades ago some manufacturers used silver alloy inlays at the bush bearing portion, as shown in Figure 5.19. Silver being a far better medium for the transfer of heat, the inlays quickly brought the heat to the outer surface of the cones. Current techniques provide small diameter through holes in the cones. These holes serve as passages for compressed air and dissipate heat efficiently. 5.3.3.2
Heat treatment
Heat treatment is the key process that gives the essential properties to a component of a tricone bit. Many intricate heat treatment processes are required to be adapted during the manufacture of a tricone bit component.
Silver Inlays
Figure 5.19 Silver alloy inlays in a cone.
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To give surface wear resistance to a cone, the surface is carburized. Carburization is a process where components are packed in a carbonaceous carburizing compound in a closed box and heated in a furnace up to about 925°C. At such temperatures the carburizing compounds liberate carbon monoxide (CO) gas. Carbon in this gas penetrates the surface of the alloy steel component through the process of diffusion, and dissolves. As a result, the surface of the component becomes very hard and wear resistant. The thickness of such a hard surface lies between 1.7 mm to 3.3 mm depending upon the component and the time period over which it is kept in the carbon- based atmosphere. The central portion, where carbon is not absorbed, remains tough. Carbon content of the surface material varies between 0.7% to 1.2%. Figure 5.20 shows a portion of a cone where the darker carburized surface can be easily differentiated. Sometimes instead of using carbonaceous compounds, CO gas is used directly. The hardness level achieved in this manner is 63–65 Rockwell C. Certain portions of the surface are not to be carburized as they have to be machined further. In such cases, a coating is applied to that portion before it is carburized. The coating does not allow absorption of carbon at the surface beneath it. This type of selective carburizing can also done by heating the desired portion of the part by using a high intensity oxyacetylene flame of proper size and shape. Figure 5.21 shows a bit leg where the dark carburized surface is seen only on one portion that was not coated while carburizing. Metallurgical processes described here need very specialized equipment in the manufacturing plant. Quality control is of prime importance in the manufacturing process since performance of the bits is watched by the end users very carefully. Guarantees of bit performance have now become an almost unavoidable feature of bit sale.
5.3.4
Summary of design features
The chart in Figure 5.22 gives the qualitative summary of design features of various bits. The levels of different parameters indicated by bar graphs are variable between the two extreme values for the parameter viz. 0 to 10. Naturally, all the levels are
Hard Carburized Surface Tough Non Carburized Core
Figure 5.20 Cone carburization.
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Carburized Surface
Non Carburized Surface
Figure 5.21 Selective carburization. Bit Type Basic Design
Soft
Medium
Hard
Extra Hard
Offset Journal Angle
Cutting Structure Design
Scraping Action Crushing Action Tooth Depth Tooth Spacing Included Tooth Angle
Strength
Bearing Strength Carburized Case Depth
Metallurgy Tooth Hardfacing Gage Hardfacing
Figure 5.22 Summary of design features.
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Table 5.6 Recommended bit loads and rotary speeds for tricone bits. IADC code
Bit load kN/mm
Rotary speed in RPM
Steel tooth bits 112⇒142 212⇒242 312⇒342
0.175⇒0.525 0.525⇒0.875 0.700⇒1.225
70–120 60–100 50–80
TC insert bits 412⇒442 512⇒542 612⇒642 712⇒742 812⇒842
0.175⇒0.875 0.525⇒1.137 0.700⇒1.225 0.700⇒1.400 1.050⇒1.575
50–150 50–150 50–120 50–90 40–80
relative and not absolute. Similarly the hardness of the formation indicated by ten columns is also relative and not to any specific scale.
5.3.5
Bit load and rotation of tricone bits
As is amply evident from the discussion about the cutting action under the teeth of tricone bits, very high side forces are also exerted on tricone bits during the formation fracturing process. Naturally, tricone bits have to be rotated at low speed and high torque. With increasing feed force the torque requirement rises. To avoid shocks being experienced by the teeth, the rotary speed is required to be reduced. Normal recommended bit loads and rotary speeds to be used in tricone bit drilling are as shown in Table 5.6. For larger diameter bits, lower rotary speeds and higher bit loads should be chosen. An empirical equation put forth by Praillet for calculating optimum weight on the tricone bit is as follows. Wo = σc ∗ D/2 where Wo = Optimum weight on the bit in kg σc = Compressive strength of rock in kg/cm2 D = Diameter of the tricone bit in cm
5.3.6
IADC code
The International Association of Drilling Contractors (IADC) has formulated a standard four digit code for classification of tricone bits. It is explained in Figure 5.23. Almost every tricone bit manufactured by any manufacturer can be classified by choosing appropriate numerals or letters in place of the digits. This code is very suitable for comparing tricone bits and for unambiguous report preparation.
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First Digit
Used for Bit Type and Formation
Second Digit
Used for Further Formation Suitability
Third Digit
Used for Bearing Type
Fourth Digit
Used for Additional Design Features
Figure 5.23 Explanation of Four Digits in IADC Code.
Table 5.7 Key to first digit of IADC code. Number
Description (bit type and formation)
1 2 3 4 5
Steel Tooth Bit for Soft Formations Steel Tooth Bit for Medium Formations Steel Tooth Bit for Hard Formations TC Insert Bit for Soft Formations TC Insert Bit for Medium Soft Formations TC Insert Bit for Medium Formations TC Insert Bit for Medium Hard Formations TC Insert Bit for Hard Formations
6 7 8
The first digit of the code is a numeral which broadly indicates the type of bit and the formation hardness for which the bit is suitable. One of eight numerals can be chosen for this digit. Table 5.7 gives the key for selecting the appropriate numeral for the first digit. The second digit of the code is a numeral that broadly indicates the formation hardness for which the bit is suitable. One of four numerals can be chosen for this digit. Table 5.8 gives the key for selecting the appropriate numeral for this basis. The third digit of the code is a numeral that indicates the type of bearing used in the bit. One of seven numerals can be chosen for this digit. Table 5.9 gives the key for selecting the appropriate numeral according to this criterion. The fourth digit of the code is a letter that indicates the additional features provided in the bit. One of nine letters can be chosen for this digit. Table 5.10 gives the key for selecting the appropriate letter according to this criterion. Due to continuous developments in manufacturing techniques of tricone bits, most of the manufacturers restrict to first three digits of IADC codes for their bits and use the fourth digit as per a table made by them for their bits which can contain more letters than those presented in Table 5.10.
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Table 5.8 Key to second digit of IADC code. Number
Description (further formation suitability)
1 2 3 4
Soft Formations Medium Formations Hard Formations Extra Hard Formations
Table 5.9 Key to third digit of IADC Code. Number
Description (bearing type)
1 2 3
Standard Open Bearing Standard Open Bearing for Air Flushing Standard Open Bearing with TC Hardfacing or TC Insert Protection at the Cone Heel Roller Sealed Bearing Roller Sealed Bearing with TC Inserts at the Cone Heel Journal Sealed Bearing Journal Sealed Bearing with TC Inserts at the Cone Heel
4 5 6 7
Table 5.10 Key to fourth digit of IADC code.
5.4
Letter
Description (additional design feature)
A R C S D E G Z J
Air Flush Application Reinforced Welds Single Central Jet Standard Steel Tooth Model Deviation Control Feature Extended Jet Nozzle Extra Gage Protection Other Shapes of TC Inserts Jet Deflection
DULL BIT ANALYSIS
Observations about the conditions of a bit are made frequently during drilling. Eventually, like everything else in this world, a tricone bit comes to the end of its useful life. Such bit can be scrapped or sent for salvage, if economically feasible. Before doing so, it is very essential to carefully observe the condition of the failed bit and analyze the reasons for failure. Reasons for failure established by such analysis give good guidance to the user for achieving more efficient operations and better drilling economy in future. If such data are provided to bit manufacturers it can be very helpful in their research towards improved products.
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5.4.1
Bit observation
In any observation ‘when’ and ‘what’ are the most important criteria. As far as ‘when’ is concerned, there are three stages of bit observations viz. 1 2 3
Observation During Drilling Operation Observation After Bit Withdrawal Observation At Final Bit Rejection
5.4.1.1
Observation during drilling operation
During blasthole drilling operations the driller does not see the bit because it is inside the hole. He cannot even see the cuttings ejected out of the blasthole as the area is covered by a large dust hood. However, the driller is able to see the rotating drill pipe and can judge the penetration rate from axial movement of the drill pipe. On most modern blasthole drills a drilling recorder is provided in the operators cab. In addition to the information on other drilling parameters, the recorder almost invariably gives the current penetration rate in numerical units. From such indirect observations an experienced operator can judge if the bit is performing as per expectations or whether there is something wrong with the drilling bit. 5.4.1.2
Observation after bit withdrawal
It is a good practice to inspect a tricone bit after drilling a certain number of blastholes. As the bit usually remains within the dust hood, the hood has to be lifted and a torch has to be used for inspection of the bit. In many sophisticated and large blasthole drills, arrangements are provided for mechanical raising of the dust hood. The dust hood must be lifted before the drill moves to the location of a new blasthole so that the hood itself does not push the cuttings back into the blasthole. These curtain lifting arrangements come in very handy while inspecting the drill bit, cuttings and other associated factors. The best time to inspect the bit is when the drill has moved to the new blasthole position and is being leveled by using its jacks. In this type of inspection, the bit can be inspected for general wear condition of the cones, loss of gage i.e. reduction of the bit diameter, free rotation of the cones and broken teeth. 5.4.1.3
Observation at final bit rejection
When a bit starts giving a penetration rate below the acceptable level, it is replaced by a new bit. Observation of the rejected bit can be done far more thoroughly by considering the following conditions: 1 2 3 4
Tooth or Insert Conditions Cone Shell Conditions Bearing Conditions Bit Body Conditions
For the purpose of keeping records, the abbreviations shown in Table 5.11 are often used. As most of the abbreviations are acronyms of the condition to be observed,
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Table 5.11 Key for recording observations of a rejected bit. Tooth or insert conditions BT Broken Teeth or Inserts CD Cone Dragged CR Cored RG Rounded Gage WG Worn Out of Gage BG Broken Gage Insert Cone shell conditions BA Broken Axially BC Broken on Circumference BS Broken Spear point IH Cone Interference – Heel IM Cone Interference – Middle IN Cone Interference – Nose CS Cracked Axially CC Cracked on Circumference EC Eroded Cone shell
Rotation or bearing conditions BF Bearing Failure BP Broken Bearing Pin BR Broken Rollers CL Cone Locked IT Inward Thrust Wear LC Lost Cone LR Lost Rollers NF Nose Bearing Failure NL Nose Bearing Loose OT Outward Thrust Bit body conditions BL Bent Legs DB Damaged Bit SB Shirttail Broken SW Shirttail Worn
they are easy to remember. The table also gives a very clear idea as to what is required to be observed under each of the above conditions. Besides the reasons for bit rejection, as given in Table 5.11, a remark regarding how well the bit has performed, must also be recorded. For this purpose letters G+, G, A+, A, A−, P and P− are used to indicate very good, good, above average, average, below average, poor and very poor.
5.5
BIT RECORDS
It is very important to keep records of bit performance. For this purpose, a sample form duly filled is shown in Table 5.12. In the form, anything that has to appear on the original blank form, is printed on gray background. Whatever appears on white background is the data to be filled by the user. Sample data is filled up in these cells for guidance. All the data mentioned in the form, including the name of the company, drill models bit manufacturers etc., is hypothetical. Since the data have been chosen arbitrarily, two values of otherwise dependent factors may not even have any correlation. Therefore, the data should not be used for drawing any conclusions. Instructions for filling the form are as follows.
5.5.1
General data
All the details to be filled in are self explanatory. They can be obtained from geographical and meteorological records of the mine site.
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Table 5.12 Sample bit record form. Indian Copper Ltd.
Head Office - 1732, Uttung Towers, 9th Road, Andheri, Mumbai 400053. Mines - Anypore, Bihar; Malanjkhand, Madhya Pradesh.
Tricone Bit Record
Date :- April 14, 1997
General Data Mine Location
Malanjkhand
Block Name
17H
Altitude in m
1000
Approximate Temp. in °C
24
Waste Type
Granite
Ore Name
Chalcopyrite
UCS of Waste in MPa
220
UCS of Ore in MPa
100
DRI of Waste
38
DRI of Ore
68
Special Details of the Block
The block has very low percentage of intermixed ore and therefore all the blasted and excavated material will be treated as waste
Formation Data
Blasthole Drill And Accessories Details Make and Model
Bucyrus Intl. 39R
Drill Recorder and GPS
Drill Recorder Yes, GPa None
Maximum Bit Weight (kN)
540
Max. Rotary Speed (RPM)
160
Air Discharge (m3/min)
85.5
Air Pressure (kPa)
448
Drill Pipe Diameter (mm)
273
Single Pass Length (m)
13.7
Drill Bit Diameter (mm)
349
Stabilizer Type
Roller Stabilizer
Bit Performance Details Bit Make Bit Model
Serial No.
Meters Drilled
Assay. No.
Ore
Waste
Total
Hours in Use
Bit Weight kN
ROP in m/h
RPM
Perform. Class
Hughes
HH55
AA123
722K
-
2615
2615
38.5
67.92
360 - 380
50 - 60
G+
Smith
Q5JL
BQ117
573H
1320
1257
2577
48.25
53.41
350 - 370
60 - 70
A+
Reed
MCM
AJ338
119P
837
1570
2407
53
45.42
360 - 380
45 - 60
A
Security
M8M
ZZ101
357S
889
1805
2694
55.5
48.54
350 - 390
50 - 60
G
Bit Discard Details Serial Number
AA123
BQ117
AJ338
Discard Reasons
BF, RG
NF
SW, LR
ZZ101 BT
Cone Number
1
2
3
1
2
3
1
2
3
1
2
3
Bearing Condition
7
6
7
7
7
8
7
7
7
6
6
5
Spear point Condition
5
0
0
7
0
0
2B
C
1M
7
6
7
Inner Teeth Condition
5
5
5
6
6
6
1B1M
3B
C
5
6
6
6B1M
1B
8B2M
5
5
4
1B
1M
2B
1B
2B
2M
Gage Condition NOTES -
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5.5.2
111
Formation data
UCS means Unconfined Compressive Strength. DRI means Drilling Rate Index. Both these values are determined in a rock testing laboratory. In most of the mining operations the type of rock in which blastholes are drilled does not change very frequently. Hence the tests for UCS and DRI are carried out only when a substantial change is experienced in penetration rate, bit wear etc. In the item ‘Specific Details of the Block’, as many relevant details of the rock mass as possible are to be given.
5.5.3
Blasthole drill and accessories details
All the details to be filled in, are self-explanatory. These can be copied from the manufacturers’ details of the blasthole drill and accessories.
5.5.4
Bit performance details
Bit details, viz. bit type, serial no., assembly no., etc. are embossed on the pin as shown in Figure 5.8 from where they should be copied. ROP means the net rate of penetration which equals meters drilled divided by hours spent on actual drilling operation. Time for moving the drill to the blasthole location, leveling the drill, adding or removing the drill pipe etc. should not be included in the actual drilling operation hours. Bit weight should be mentioned in terms of kN i.e. if the pulldown ranged between 550 kN and 580 kN the figure to be entered in the column should be 550–580. RPM means rotary speed in revolutions per minute. The range should be mentioned by choosing the fastest and slowest speed. For entering proper values in the bit weight and RPM columns, data may be obtained from the drilling recorder provided in the blasthole drill. Abbreviations meant for indicating the performance class have been mentioned in earlier sections.
5.5.5
Bit discard details
Discard reasons should be mentioned by using the abbreviations given in Table 5.11. Bearing, spear point and inner teeth conditions should be mentioned for each of the three cones of the bit by estimating the amount of wear in terms of levels 0 to 8 (Table 5.12). Here 0 means virtually no wear and 8 means full wear. If some inserts of the bit have broken or are missing, the number of such inserts followed by B (broken) and M (missing) should be mentioned. Photographs of rejected bits may be taken in cases where the failure is abnormal and they should be stored systematically by modern methods of electronic data storage. For any additional details separate blank pages should be used. If a warranty claim for very poor performance of the bit is to be raised, the bit along with copies of the bit report form, corresponding drill recorder chart and the bit photograph must be sent to the manufacturer. The manufacturer is not likely to consider replacement without such information.
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Depending upon the manner in which bits have failed, it may be possible to repair and reuse some tricone bits. Many manufacturers do such types of refurbishing. Lastly, a discarded tricone rock bit may still have tungsten carbide inserts, silver alloy etc. and can fetch some salvage value. Therefore, it should be preserved for a meaningful salvage. 5.6
BIT FAILURE ANALYSIS AND REMEDY
Table 5.13 gives many types of bit failures along with their photographs and discusses the reasons of failure as well as some suggested remedial actions for achieving better performance. Table 5.13 Bit failure analysis and remedial measures. The teeth of this bit are worn out only at the surface that touched the bottom of the blasthole. Teeth on the inner portion of the cone are intact. Spiral marks can be easily observed at the worn out surface. This means that the bit failed prematurely because the rollers, somehow, got locked and drilling operation was carried out by friction scraping of the formation at the bottom of the hole much like that of a diamond bit. Tricone bits are primarily designed for crushing action with little scraping associated with it due to offset. Such failure can be easily noticed by an attentive operator because the penetration rate drops suddenly and the torque required for rotating the drill string increases suddenly. Wear on the gage teeth of this bit is much more than that on the inner teeth. Such failure pattern can be encountered when medium soft but very abrasive formations, such as sandstone, are drilled with a bit meant for drilling non abrasive formations. Obviously the selection of this bit must have been incorrect. A bit with better gage protection such as H or T gage with gage hardfacing or a milled tooth bit with TC inserts on the gage will give better performance in the formation. Excessive wear of the teeth can be seen on all the three cones. Further the wear is uniform from gage to the center of the cone as well as all around the cone. If the bit has failed prematurely the reason of failure must have been excessive weight on the bit. Excessive weight can result in quick wear of the tip of the teeth and once the carburized portion at the tip of the teeth is worn out the inner portion wears very easily as it is not hardened. Operating drill with lesser feed force may give better performance. Such a wear can be caused when the bit is rotated in hard and abrasive formations at a high speed without adequate feed force. Rotary speed in such conditions causes wear on the bit in a manner similar to grinding action. Performance of the bit will improve by reducing the rotary bit or alternatively by choosing a harder formation bit which has more number of teeth.
(Continued)
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Table 5.13 (Continued) In this bit very heavy wear on the shirttail and the gage inserts of left hand side cone is clearly visible. Such failure can be caused when the bit is coupled to a bent drill pipe. Rotary motion given to such a pipe always cause very heavy side thrust through the bit. In particular, one or two of the cones have to bear the load to a much greater extent than the others. Such failure can also occur if the carbide inserts of the bit brake loose and remain in a position just above the cone due to insufficient flushing. Remedial action will be in replacing the damaged drill pipe. In this bit the wear on all the three cones of the bit is very similar. The inserts on the gage of all the three cones have broken but comparatively much less wear can be seen on the inner cones. Such wear pattern can occur when a bit meant for drilling in the soft formation is used for drilling very hard formations. In such conditions the force to which the gage inserts are subjected is very high. Remedial action will be to choose a proper hard formation bit.
In this picture a vertical crack can be seen on the pin of the bit. Besides some damage to the threads of the pin shown by another bunch of arrows can also be seen in the middle of the pin. Such thread damage and cracks are formed when the bit is cross threaded to the drill pipe or the stabilizer and is used for drilling. Remedial measure will be to remove the drill pipe or the stabilizer with damaged threads of the box. Normally the box of the drill pipe or stabilizer. Bearing failure of the bit can be easily observed in upper photograph. The failure is so predominant that the rollers of the bearing have been lost. The bearing failure can be seen only on one of the three cones. For further investigations when the three legs of the bit were separated, it could be seen that the thrust button below the cone of which the bearing failed, had bent as shown in the lower photograph.This must have resulted in very little or no air circulation through the bearing. Consequently the bearings and the surrounding surface got hot and the carburized case as well as the bearings must have lost hardness.The journals of the bearings of other two cones were intact. A centrally located crack, all along the seam of the two legs, can be observed on the bit. Such a failure can occur if the bit was accidentally dropped to the bottom of the hole.The impact experienced by the bit was so high that the tensile stresses developed along the seam were higher than the strength of the weld. Alternatively such failure can occur if the stress relieving operations during the process of manufacturing the bit were not done or were not properly carried out. A centrally located crack, all along the seam of the two legs, can be observed on the bit. Such a failure can occur if the bit was accidentally dropped to the bottom of the hole. The impact experienced by the bit was so high that the tensile stresses developed along the seam were higher than the strength of the weld. Alternatively such failure can occur if the stress relieving operations were not done or were not properly carried out during the process of manufacturing the bit. (Continued)
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Table 5.13 (Continued) The cone of this bit has broken circumferentially. The nose and inner portion of the cone has been lost. The bearings under the cone have been exposed and some have been lost. When bearing of a bit fails the cones may interfere with each other and only one of the three cones gets subjected to full weight on the bit. In such circumstances this type of failure occurs. The cone of the bit has cracked near the shirt tail of the bit. This type of failure can take place when the bit is dropped in the blasthole with a heavy drill string weight on it. Such failure can also be caused by loading the bit with excessive bit weight in extremely hard formation. One of the three legs of the bit is completely broken and lost. This type of failure takes place when a bit is dropped or when it hits a ledge in the blasthole or when it hits. If the leg has a crack and that has remain undetected in the quality control checks the failure of this type can take place very easily. One cone of the bit is completely lost. The other cone is also broken circumferentially. This failure can be considered as advanced stage of cone breakage because a cone can be completely lost only when one cone breaks circumferentially. The root cause of this type of failure is the bearing failure which in turn is caused by when one of the three cones gets subjected to the full weight exerted on the bit.
5.7
DISCARDING DRILL BITS
Bits wear out when used for drilling. In some cases the loss of gage is intolerable or in other cases the penetration rate drops down to an unacceptable level. A driller must know what is meant by ‘intolerable’ or unacceptable. Drilling with a bit that has lost its gage will give lower blasthole volume. The quantity of explosive accommodated in the blasthole will thus be lower, and insufficient for achieving the desired level of fragmentation. Generally the tolerance level of a quantity of explosive in a blasthole is about 8%. Therefore, a drill bit should be discarded when the cross sectional area of the hole reduces to 92% of the original. This happens when the diameter of the bit reduces to about 0.96 times the original nominal diameter. It is very uneconomical to continue drilling with a bit that is giving low penetration rate because the blasthole drill that costs several hundred times more than the cost of a drill bit, is being utilized wastefully for merely saving the cost of a bit. Solutions to these situations are in sharpening the bit by grinding to give a faster penetration rate, or in a very few cases welding hardmetal by special electrodes. Blade type drag bits can easily be reconditioned by blade grinding. Hardmetal side welding to restore gage is a bit more difficult but can be tried. As such, blade bits can be used almost to complete destruction.
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Percentage of Penetration Rate of the New Bit
The teeth of a claw type drag bit can be changed when penetration rate is unacceptable. It is possible to grind these tips and reuse them, but in most cases this will result in a serious loss of gage so replacing the peripheral inserts is very essential. Steel tooth tricone bits are not subjected to grinding because in grinding the hardened surface is seriously weakened and cannot be restored. If a steel tooth bit is reused after grinding, it may give satisfactory penetration rate for a very short duration but very soon will become worth discarding. Thus, the reuse will not be worth the cost incurred in grinding and keeping the drill functionally idle while changing the bit. TC (tungsten carbide) insert tricone bits can be subjected to grinding if the wear on the inserts is very high and the bit is otherwise, i.e. from the viewpoint of bearing condition etc. in good condition. Figure 5.24 shows the manner in which the penetration rate of a tricone bit reduces with the increase in bit wear. The actual curve for a bit lies within the shaded area and the curvature is very similar to the upper or lower limiting curve. Averages of the data indicate that in the case of tricone bits, a completely worn out bit gives only 20% penetration rate as compared to the new bit. Whether to continue drilling till the bit is completely worn out, is more of an economical decision rather than a technical one. From the economic angle this topic has been discussed in the chapter on costing. It is not out of place to mention here that many bit manufacturers conduct short training courses about use of their drill bits in the field. Such training courses can also be arranged at the worksite office. Attempts should be made to arrange such a training course so that many drillers can take advantage of it.
100 80
60 40
20 0
0
20
40
60
80
100
Percentage of Bit Wear Figure 5.24 Reduction in penetration rate with bit wear.
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Chapter 6
Rotary blasthole drilling accessories
6.1
INTRODUCTION
Besides a drill bit, the other components in a rotary blasthole drill string are a stabilizer, drill pipes, a shock absorber and crossover subs. Apart from their specific individual functions, all these components have three common functions viz: 1 2 3
Transmitting the feed force exerted by the feed mechanism of the blasthole drill to the drill bit. Transmitting the rotary motion generated at the rotary head of the blasthole drill to the drill bit. Allowing a continuous flow of compressed air to the next lower component in the drill string without significant pressure loss.
During the transmission of feed force and torque, the components are subjected to very heavy vibrations and must be very strong in their fatigue life. Blasthole drilling operations also require some more items that are not part of drill string but are frequently needed for some specific purpose. Specific information on these accessories is given in the following elaboration.
6.2
DRILL PIPES
The drill pipe is a very important component of a drill string. In all the rotary blasthole drills, the drill pipes are stored in a pipe changer fitted in the mast and are mechanically added to the drill string as the drilling advances, or removed and stored back in the pipe changer when the drill string is being withdrawn after completion of the blasthole. Drill pipes used in rotary blasthole drilling are flush type – meaning that their outside diameter is the same through out their length. They have a box connection (i.e. female threads) at one end and a pin connection (i.e. male threads) at the other end. They have wrench flats near both their ends to facilitate easy coupling or uncoupling by the wrenches provided in the blasthole drill. Drill pipes are of two types – fabricated or integral. Both are illustrated in Figure 6.1.
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Upper Tool Joint Fabricated Drill Pipe
Lower Tool Joint
Lower Connection
Wrench Flat
Weld
Seamless Tube
Weld
Wrench Flat
Upper Connection Integral DrillPipe Upper Tool Joint
Figure 6.1 Fabricated and integral drill pipes.
6.2.1
Integral drill pipes
Integral drill pipes are made from solid alloy steel bars and as such they are in one piece. They have a central bore of uniform diameter through which compressed air flows to the bit. As the wall thickness of integral drill pipes is large, they are very heavy and sturdy. Due to large wall thickness they are stiff, and their susceptibility to buckling under heavy feed force is very low. Integral drill pipes are made from alloy steel, such as ANSI 4145, and are surface hardened by heat treatment. Compared to fabricated drill pipes, integral drill pipes are expensive but have much longer life. Single piece integral drill pipes are available in lengths up to about 12 m.
6.2.2
Fabricated drill pipes
Fabricated drill pipes have three pieces, viz. upper tool joint, seamless tube and lower tool joint, welded together. The upper and lower tool joints are made from alloy steels such as ANSI 4140 or 4145. The seamless tube is made from standard carbon steels such as A-106B or J55. Some manufacturers use an alloy steel tube of the same material as that of the tool joint so as to enable uniform heat treatment to be given to the complete drill pipe. Due to lesser weight and manufacturing cost, the price of fabricated drill pipes is relatively low. Fabricated drill pipes have a large diameter bore. Therefore air pressure loss in these pipes is slightly less compared to integral drill pipes. For large diameter drill pipes, tubes with larger wall thickness are chosen. Such thick tubes enable the drill pipes to withstand very heavy compressive and shear stresses generated by heavy feed force and torque exerted by the blasthole drill. Details of standard fabricated and integral drill pipes are given in Appendix 11 at the end of this book. The length of tubes used for fabricating drill pipes is normally limited to about 12 m. Longer drill pipes, of length ranging between 13.71 to 21.3 m length are also required for some blasthole drills. In such case many manufacturers use special welding techniques to weld two tubes end to end, and then add the tool joints to the long
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119
tube. This alternative is better than using two shorter drill pipes permanently coupled together. Tool joints, i.e. box or pin type end connections of the drill pipes, are either API Regular or BECO type. API Regular tool joints originate from the oilwell drilling industry. For a long time they were successfully used for drill pipes in the long rotary drill string used for oilwell drilling. In the beginning of rotary blasthole drilling, they were adopted for the short drill strings of rotary blasthole drilling also. In the late 1950s Bucyrus Erie Company developed BECO tool joint with a coarser BECO thread form. Profiles of the thread forms used for API and BECO tool joints are shown in Figure 6.2 and their dimensional details are given in Table 6.1 and 6.2 respectively. Profile and dimensional details of the API Regular tool joints are in Figure 6.3 and Table 6.3. Profile and dimensional details of the BECO tool joints are in Figure 6.4 and Table 6.4. API threads were made from the viewpoint of deep drilling with fluid circulation. They were fine threads, so the cross section of the spiral tunnel formed between apex API Threads
P Box
Srn rcn H
h
r
Position of Spiral Tunnel Fcs
Fcn 60°
r 30°
rcs
fcn Pin Axis of Thread
90°
BECO Threads Box
T
P H
Position of Spiral Tunnel
30° 30° r
Pin 90°
Figure 6.2 Thread forms used for API and BECO tool joints.
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Table 6.1 Dimensional details of thread forms used for API regular tool joints. Thread form
Taper %
H in mm
hn = hs in mm
Srn = Srs in mm
fcn = fcs in mm
Fcn = Fcs in mm
rcn = rcs in mm
r in mm
V.0.040 V.0.050
25 25
4.38 5.47
2.99 3.74
0.51 0.63
0.88 1.09
1.02 1.27
0.51 0.63
0.38 0.38
Table 6.2 Dimensional details of thread form used for BECO tool joints. Thread form
Taper %
H in mm
P in mm
T ± 0.0508 in mm
r in mm
BECO
25
7.366
12.7
2.7686
1.1938
A
F 5/ 8” 30°
G
D
C
H – Threads Per Inch
B
E G
K – Inclination of the Taper
Figure 6.3 Profile of API regular tool joint. Table 6.3 Dimensional details of the API regular tool joint and thread forms used (All dimensions in mm). API pin size
A
B
C
D
E
F
G
H
K
Thread form
2–3/8 2–7/8 3–1/2 4–1/2 5–1/2 6–5/8 7–5/8 8–5/8
3 3–1/2 3–3/4 4–1/4 4–3/4 5 5–1/4 5–3/8
3–1/8 3–3/4 4–1/4 5–1/2 6–3/4 7–3/4 8–7/8 10
2–5/8 3 3–1/2 4–5/8 5–33/64 6 7 7–61/64
1–7/8 2–1/8 2–9/16 3–9/16 4–21/64 5–5/32 5–11/16 6–19/64
2–11/16 3–1/16 3–9/16 4–11/16 5–37/64 6–1/16 7–1/16 8–1/64
3–3/8 3–7/8 4–1/8 4–5/8 5–1/8 5–3/8 5–5/8 5–3/4
1 1–1/4 1–1/2 2–1/4 2–3/4 3–1/2 4 4–3/4
5 5 5 5 4 4 4 4
3 3 3 3 3 2 3 3
V-.040 V-.040 V-.040 V-.040 V-.050 V-.050 V-.050 V-.050
of male threads and trough of female threads was very small and did not allow any significant leak of the circulating fluid that was flowing under very high pressure. As against this, BECO threads were made for blasthole drilling, where the pressure of compressed air was low. Thus, despite the larger cross section, the leakage in the spiral tunnel was sufficiently low.
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A
121
F 5/8” 30°
G
D
H – Threads Per Inch
B E
C
G
K – Inclination of the Taper
Figure 6.4 Profile of BECO tool joint.
Table 6.4 Dimensional details of the BECO tool joint and thread forms used (Dimensions are in mm unless specified otherwise). BECO tool size
A
B
C
D
E
F
G
H K Thread form
3″ BECO 3½″ BECO 4″ BECO 4½″ BECO 5¼″ BECO 6″ BECO 8″ BECO 10″ BECO 11″ BECO (11½″)
80 83 90 93 109 112 115 166 179
4½″ 5″ 5½″ 6¼″ 7½″ 8 5/8″ 10¾″ 13 3/8″ 15″
85.70 98.43 111.00 123.80 142.80 162.00 212.70 263.50 289.00
51.00 60.33 73.00 84.00 98.50 117.50 167.00 205.00 255.50
92.00 104.80 117.50 130.20 149.00 168.30 219.00 270.00 295.00
111.00 114.00 120.00 124.00 140.00 140.00 146.00 194.00 206.50
38.00 38.00 57.00 57.00 76.00 102.00 120.00 120.00 120.00
2 2 2 2 2 2 2 2 2
3 3 3 3 3 3 3 3 3
BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard
Comparatively very coarse BECO threads enable the tool joint to have the following advantages: 1 2 3 4 5
Ease in frequent coupling and uncoupling of the drill pipes. Rigidity in the tool joint to reduce vibration. Leakproof when they are lubricated with thick grease that fills the spiral tunnel. Excellent ability in withstanding rough use. Much higher maximum torque rating as can be seen from Table 6.5.
Since all the above aspects are of greater importance in rotary blasthole drilling, BECO type connections are becoming more popular in rotary blasthole drilling practice. On many occasions purchasers are unaware of the difference between the drill pipes used in the oil or waterwell drilling industry and those used in the blasthole drilling industry. Purchase of inexpensive drill pipes meant for the waterwell drilling industry and their use for rotary blasthole drilling work can prove disastrous. Hence to avoid such instances, the differences are elaborated in Table 6.6.
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Table 6.5 Max. torque ratings of BECO and API reg. tool joints. BECO tool joint
API regular tool joint Maximum rated torque
Maximum rated torque
Drill pipe diameter in
mm
Tool loint name
lbft
Nm
Tool joint name
lbft
Nm
3.500 4.000 4.500 5.000 5.500 6.000 6.250 6.500 6.625 7.000 7.625 8.625 9.250 9.625 10.750 11.750 12.750 13.375 15.000
88.9 101.6 114.3 127 139.7 152.4 158.75 165.1 168.275 177.8 193.675 219.075 234.95 244.475 273.05 298.45 323.85 339.72 381
None None None 3½ BECO 3½ BECO 3½ BECO 4½ BECO 4½ BECO 4½ BECO 5¼ BECO 5¼ BECO 6 BECO 6 BECO 7 BECO 8 BECO 8 BECO 8 BECO 10 BECO 12 BECO
Nap Nap Nap 18000 18000 18000 24300 24300 24300 36300 36300 55600 55600 88000 120000 120000 120000 120000 NS
Nap Nap Nap 24405 24405 24405 32946 32946 32946 49216 49216 75383 75383 119312 162698 162698 162698 222100 NS
2 3/8 API R 2 7/8 API R 3½ API R 3½ API R 4½ API R 4½ API R 4½ API R 4½ API R 5½ API R 5½ API R 5½ API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 7 5/8 API R 8 5/8 API R
NS 7000 9500 9500 16000 16000 16000 16000 32000 32000 32000 44000 44000 44000 44000 44000 44000 70000 98000
NS 9491 12880 12880 21693 21693 21693 21693 43386 43386 43386 59656 59656 59656 59656 59656 59656 94907 132870
6.2.3
Choice of a drill pipe
For efficient blasthole drilling at fastest penetration rates, choosing the most appropriate drill pipe is of extreme importance. The choice is mainly related to the diameter of blasthole. The diameter of a blasthole is decided on the basis of many factors related to production requirements, blasting method, parameters pertaining to loading and hauling equipment used in the mine, etc. They are elaborated elsewhere in this book. Factors to be considered for choosing the appropriate drill pipe are: 1 2 3 4 5 6
Drill Pipe Weight Drill Pipe Dimensional Parameters Drill Pipe Surface Treatment Size and Shape of Drill Cuttings Bailing Velocity in the Blasthole Drill Pipe Wall Thickness
A logical method to be adopted for choosing the most appropriate drill pipe is presented in Chapter 10 of this book.
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Table 6.6 Comparison of drill pipes used for blasthole drilling with those used for oil well/water well drilling. Aspect
Oil well or water well drilling
Blasthole drilling
Dimensional precision
Need not be so exact as the pipes are not handled by a mechanical pipe changer but are stacked loosely near the mast. Length is medium i.e. often of the order of 9 m. Often 3 drill pipes are stacked without being separated.
Has to be very precise as they are handled by a mechanical pipe changer. Length is long. Mostly more than 9 m. Often as much as 18 to 21 m. They are stored singularly in a pipe changer. Since the pipes are subjected to high compressive forces during drilling hence they are susceptible to buckling. For this reason the thickness of drill pipe has to be large so the moment of inertia increases. The weight of the drill pipe is very important as they are stored in a pipe rack and affect the stability of drill. Surface hardening treatment is very essential because the pipes are continuously subjected to sand blasting from the large rock cuttings that travel upwards through the annulus at a very high uphole velocity.
Length
Thickness
Are relatively thin as they are mainly subjected to high tensile stresses while drilling or very low compressive stresses while being stacked. There is no question of buckling of these drill pipes.
Weight
The weight of drill pipe does not matter as the drill pipes are stored on ground.
Wear Properties and surface hardening
The velocity of flushing water is low. The size of particles is also small. Hence these pipes are not subjected to very high external abrasion. Surface hardening treatment of the drill pipes is not very essential when used for water well drilling. In oil well drilling it may be necessary because when holes are deviated intentionally the pipes rub against the walls of the hole. Sealing properties are very important as the pressure inside the pipe bore and that in the annulus differ very greatly. Small pitch threads or special threads that have a very small cross section of the spiral tunnel at the apexes of the threads, are used for tool joints.
Threads
6.3
Sealing properties are not very important as the pressure inside the pipe bore and that in the annulus is nearly the same. BECO threads are preferred for tool joints.
STABILIZERS
Even though the depth of the blasthole is shallow, the characteristics of the formations lying on its alignment are not truly uniform. They change in terms of voids, cracks, compressive strength etc. If drilling is continued in a blasthole having formation conditions as shown in Figure 6.5, the direction of progress of the blasthole tends to deviate from the intended direction. Reasons for such deviation are: 1 2
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Male and female threads of the tool joint of the drill pipe are not perfect. Due to the bend caused, the feed force exerted on the drill bits is not truly axial. Drill pipe is relatively slender.
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Drill Pipe Upward Air Flow of Compressed Air Flow Used For Flushing Blasthole Wall
Drill Bit Soft Rock
Hard Rock Deviated Path of the Blasthole Intended Path of the Blasthole
Figure 6.5 Blasthole deviation.
3
4 5
A large gap between outer surface of the drill pipe and the walls of the hole may allow the drill pipe to buckle under heavy feed force exerted by the feed mechanism of the drill. Vibrations are caused by imperfect material balance in the drill string. The length that effectively resists hole deviation is only the short length of about 100 to 200 mm of the drill bit shoulder.
Besides deviation, the hole may also not be truly cylindrical as it assumes a spiral shape. Disadvantages of having deviated blastholes are: 1
2
3
The periphery i.e. the shirttail of the tricone bit undergoes non-uniform wear. The gap between the bit leg and the cone, meant for circulation of air through the bearings, may increase and small cuttings may enter into the gap. This causes bearings to wear. Withdrawal of the drill string and the tricone bit from such a drill hole requires more retract force. This results in heavy wear on the shirttail i.e. outer periphery of the drill bit leg. When the crooked or randomly deviated blastholes are filled with explosive, the distribution of the explosive in the rock mass is not uniform. This causes ineffective and non-uniform blast that results in non-uniform fragmentation and secondary blasting may have to be carried out frequently.
Stabilizers are used in the rotary blasthole drill string to lessen blasthole deviation and crookedness.
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A stabilizer is fixed just above the drill bit. Since the outside diameter of the stabilizer is only slightly less than that of a drill bit, the assembly of bit and stabilizer becomes more rigid. With such a close fit stabilizer, the length of drill string that resists deviation becomes much longer. This forces the progress of the blasthole much more closely in the intended direction than when the drill bit is used without a stabilizer. Since the outer periphery of the stabilizer is made with embedded tungsten carbide inserts, during rotation of the drill string these inserts abrade the walls of the hole and reduce hole crookedness. The stabilizer body is usually manufactured from ANSI 4145H Series Alloy or non–magnetic steel. Rotary blasthole drilling stabilizers are of four types: 1 2 3 4
Replaceable Sleeve Stabilizers Welded Blade Stabilizers Integral Blade Stabilizers Roller Stabilizers Details of these are given below.
6.3.1
Replaceable sleeve stabilizers
This type of stabilizer comprises a mandrel and a replaceable sleeve. The sleeve is threaded from inside and is coupled to the mandrel. The mandrel has a box connection to enable its coupling to a tricone bit as shown in Figure 6.6.
Top Connection
Mandrel
Sleeve Body in Elevation Tungsten Carbide Inserts on Sleeve Body Sleeve Body Cross Section
Thread Connection Between Mandrel and Sleeve
Sleeve
Tricone Drill Bit
Figure 6.6 Replaceable sleeve stabilizer stabilizers.
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The sleeve has four ribs as shown in the cross section of the sleeve in Figure 6.3. These ribs have large diameter tungsten carbide inserts. These tungsten carbide inserts come in contact with the wall of the blasthole and prevent other portions of the drill string from getting abraded. The sleeve can be replaced after it has worn out. Replaceable sleeve stabilizers have fallen out of favor with users for the following reasons: 1 2 3
Due to greater tooling and materials required for manufacture, their prices are higher. Their resistance to deviation is very limited. Assembling the two parts in the field is a time-consuming job.
6.3.2 Welded blade stabilizers Welded blade stabilizers are easier to manufacture and hence are relatively inexpensive. They have straight or spiral ribs as shown in Figure 6.7. These ribs have studded tungsten carbide inserts or carbide hardfacing on their outer periphery. Technologically, welded blade stabilizers are better than the replaceable sleeve stabilizers because they prevent deflection of a drill string over a longer length. It is also possible to rework and reuse them with inexpensive welding and grinding facilities.
6.3.3
Integral blade stabilizers
These are very similar in their shape and their look to the welded blade stabilizers, except that their ribs are integral with the main body and not welded later. Straight as well as spiral integral blade stabilizers are shown in Figure 6.8.
Figure 6.7 Welded blade stabilizers.
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Figure 6.8 Integral blade stabilizers.
Figure 6.9 Air flow in roller stabilizers.
6.3.4
Roller stabilizers
Roller stabilizers have three or more rotating elements called rollers that have studded tungsten carbide inserts. The rollers are cooled by compressed air that is made to flow through the central bore of the pin on which the rollers are mounted as illustrated in Figure 6.9. The moisture-laden air flowing upward through the annular space also cools the rollers. Technologically, roller stabilizers are better than the other types as they require lesser torque than the others.
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Use of rollers stabilizers is particularly beneficial in angle hole drilling, where other types of stabilizers give very high rotary resistance caused by drill string weight. If roller stabilizers are used in their place the resistance experienced is much less, as the rollers can rotate around their bearings. Stabilizers with three rollers are commonly used, but in extremely hard and abrasive formations stabilizers with even six rollers have been used. The advantages of roller stabilizers stated above become more prominent as the diameter is larger, hence they are available in large diameters. Roller stabilizers are more expensive than the other types. Details of some standard roller stabilizers are given in Appendix 12 at the end of this book. If the diameter of the blasthole is very large, say more than 381 mm, and if vertical holes are being drilled with drill pipes with large wall thickness, a stabilizer may not be required at all because the likelihood of hole deviation without the stabilizers is very low.
6.4
CROSSOVER SUBS
Crossover subs, also called ‘Threadsaver’ subs, are shown in Figure 6.10. The main reason behind the use of these subs is that when they are coupled at both ends of the drill pipe and used in drilling, the frequent coupling and uncoupling takes place at the threaded connection of the subs and not the drill pipes. This results in long life to the drill pipes that are far more expensive items. Another feature of these subs is that they can have two different types of threads at their two ends and so enable inclusion of tools with different threaded connection in the drill string. In blasthole drilling, a crossover sub is often located between the tricone roller bit that has API Regular Pin connection and the drill pipe or stabilizer that is often provided with the popular BECO threads.
6.5
SHOCK ABSORBERS
When a drill bit is fracturing the formation at the bottom of a blasthole, it is continuously subjected to erratic movements in the vertical as well as radial directions. Shock waves are generated in the drill string due to such movements, and also due to heavy feed force and high torque. Such shock waves are particularly predominant in rotary drilling with tricone bits. In blasthole drilling the rotary and feed
Figure 6.10 Crossover subs.
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force mechanisms are somewhat firmly attached to the blasthole drill. Hence the shock waves are transmitted from the rotary head to the other parts of the blasthole drill. Basically, shocks experienced by the drill string are of four types as shown in 1 2 3 4
In the axial direction of the drill string, In the direction of the blasthole drill length In the direction of the blasthole drill width In the direction of the drill string rotation
If the drilling process is considered minutely, then the reasons for generation of these shocks can be thought of as follows: A
B
C D
When the teeth of a drill bit crush the formation, the cuttings get displaced, and the teeth also move very slightly but suddenly in the direction of pulldown force and in the direction of rotation. This results in sudden reduction and rebuilding of compressive force, and sudden decrease and increase of rotary torque transmitted through the drill string. This creates vertical and circular shock waves that travel from the bit to the drill head through the drill string components. However perfectly manufactured they may be, all the drill string components have some acceptable tolerances in their axis, threads, materials etc. Thus, each and every item of the drill string deviates from a perfect axis either in direction or offset or both. Further, material distribution around this axis also changes in terms of inertia. Hence the drill string generates radial vibrations in a plane perpendicular to the drill string axis. By nature, all rocks are formed of small particles bound together. They are not ideally homogeneous. This creates shock waves of all the four types mentioned above. Often, large size rock fragments formed in the process of drilling do not get lifted with the available air drag. They greatly resist rotary motion momentarily when stuck in the annulus and release the resistance suddenly when they are further fragmented. This also creates a shock wave in the radial and circular directions of the drill string axis. Axis of Drillstring
Lateral Vibrations Rotary Vibrations Longitudinal Vibrations
Direction of Length of the Blasthole Drill Direction of Width of the Blasthole Drill Axial Vibrations
Figure 6.11 Modes of vibration in a drill string.
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All the four types of shocks reduce the fatigue life of the blasthole drill to a very great extent. Shocks in the axial direction of the drill string are particularly damaging. Shock absorbers are the devices which reduce the transmission of shocks to the rotary head and other components of blasthole drills. Advantages of installing a shock absorber in the drill string can be summarized as follows: 1 2 3 4 5 6
Increased life of the drill head components due to reduced shock waves. Slight increase in penetration rate. Reduced maintenance. Increased life of the drill pipes and bit due to jerkless feed force and rotary motion. Possibility of exerting more bit weight and rotational speed to give increase in the penetration rate. Reduced noise level due to reduction of metal-to-metal contact between vibrating elements in the blasthole drill components. There are two types of shock absorbers as follows:
1 2
External Shock Absorber In-the-Hole Shock Absorbers Details of these shock absorbers are as under.
6.5.1
External shock absorber
External shock absorbers are fitted in the drill string just below the rotary head, as shown in Figure 2.6. Two basic designs of shock absorbers have evolved. Both use synthetic rubber for shock absorption. Figure 6.12 illustrates the one made by Mining Technologies International (MTI). The complete assembly, shown on the left, consists of elements like the one shown on right. In this type of shock absorber, the axial shocks are absorbed by the shear stresses and strains developed in the rubber elements that are sandwiched between two drive lugs as shown in the figure. For absorbing the shocks in a plane perpendicular to the drill string axis the rubber element moves in compressive stress and strain. The construction of the MTI shock absorber is such that, if one or even two rubber elements fail, the other elements bear the load and the shock absorber continues to transmit the feed force and rotary motion to the drill pipes. The elements can then be replaced after the blasthole is completed. The effectiveness of the MTI shock absorber can be seen from Figure 6.13 that shows a vibrograph of the torque absorption. Another shock absorber, shown in Figure 6.14, is from Duraquest. It features a rubber element sandwiched between two plates. These plates have lugs that pass through the rubber element and holds all the components together. The rubber element absorbs all the four types of shocks with the help of
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Single Element
Vertical Shocks
Drive Lug
Horizontal Shocks
Drive Lug Rubber Element Shock Absorbing Elements
Vertical Shocks
Figure 6.12 MTI shock absorber.
Vibrations Recorded Before Installing a Shock Absorber
Vibrations Recorded After Installing a Shock Absorber
Figure 6.13 Effectiveness of MTI shock absorber in reducing shocks in rotary motion. Lugs
Complete Assembly
Top Plate Lugs
Rubber Element
Bottom Plate
Figure 6.14 Duraquest shock absorber.
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the lugs. The effectiveness of Duraquest shock absorbers is very evident from the vibrographs shown in Figure 6.15, which reveal that these shock absorbers can reduce vibrations to an extent of 79% in the axial direction, 10% in the radial direction parallel to the length of drill and 39% in the radial direction parallel to the width of the drill. Details of shock absorbers manufactured by MTI and Duraquest are given in Appendix 13 at the end of this book.
6.5.2
In the hole shock absorbers
These shock absorbers are attached at the bottom of the drill string just above the bit. They thus go in the hole, and hence the name. One shock absorber of this type is shown in Figure 6.16. These shock absorbers have a mechanical spring between two elements to absorb the shocks. The rotary motion and torque are transmitted through the splines of the two elements that intermesh with each other. In-the-hole shock absorbers are meant primarily for the oilwell drilling industry and are designed for absorbing shocks in the axial direction only. They are also used in horizontal directional drilling but very rarely in blasthole drilling. The blasthole drilling industry almost exclusively uses external shock absorbers for the following reasons: 1 2
3
4
External shock absorbers can absorb all types of shocks, whereas an in-the-hole shock absorber absorbs only the axial shocks. In blasthole drilling the drill head is very firmly fixed to the blasthole drill. Further, the drill string is very short and in heavy compression. Hence shocks transmitted to the blasthole drill components are of very high intensity. In oilwell drilling, the drill string is very long and much of it is in tension. Further, in oilwell drilling practice the drill string is hanging on wire ropes, hence the question of transmission of shocks to the drill components (except the rotary table) does not arise at all. So, in-the-hole shock absorbers are more suitable in oilwell drilling but not in blasthole drilling. An in-the-hole shock absorber goes inside the hole. In blasthole drilling it is continuously met with harsh sandblasting treatment from the drill cuttings flowing with the compressed air. Therefore an internal shock absorber wears very fast. The external shock absorber does not meet with any sandblasting. In blasthole drilling, failure of external shock absorber and subsequent remedial measures may be time-wasting but are not disastrous like the failure of an inthe-hole shock absorber. Shock absorbers have many fixtures to match with the fixtures of the rotary drill head. They should be selected for maximum rated bit weight of the drill because maximum bit weight is occasionally exerted in the process of drilling.
Many drill manufacturers provide shock absorbers of their own make on their drills. Most of these are very similar to those made by Atlas Copco. Eventually upon failure they can be replaced by shock absorbers made by specialist companies.
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Vibrographs on the right side were taken under identical drilling conditions but with and without Duraquest Smoothdrive”. Pulldown was 85000 lb. rotary speed was 90 rpm and average penetration rate was 0.80 ft/min. The upper vibrograph represents the vibrations at the drill head without Duraquest Smoothdrive, whereas the lower graph represents the vibrations at the drill head with Duraquest Smoothdrive. Comparison of the two vibrographs reveals following.
Aspect
With Smoothdrive
X Axis Average Acceleration
9g
5.5g
X Axis Peak Acceleration
12g
9g
Y Axis Average Acceleration
5g
4.5g
Y Axis Peak Acceleration
9g
11g
Z Axis Average Acceleration
12g
2.5g
Z Axis Peak Acceleration
22g
11g
Reduction of Vibrations in X Axis
(9g-5.5g)/9g = 39%
Reduction of Vibrations in Y Axis
(5g-4.5g)/5g = 10%
Reduction of Vibrations in Z Axis
(12g-2.5g)/12g = 79%
Figure 6.15 Effectiveness of duraquest shock absorber in reducing shocks in XYZ directions.
Rotary blasthole drilling accessories
Without Smoothdrive
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Figure 6.16 In the hole shock absorber.
6.6
MISCELLANEOUS ACCESSORIES
Apart from the items that go into the drill string, many other items are used in rotary blasthole drilling. The following information about such items is given to bring completeness to this book.
6.6.1 Thread protectors Thread protectors are made up of nylon or similar stiff material or even cast steel. Many varieties exist. Their function is to protect the threads of the drill pipes or other threaded drill string components during transportation and handling. Male and female type thread protectors are shown in Figure 6.17.
6.6.2
Lifting bails and hoisting plugs
The main function of lifting bails is to enable easy handling of drill pipes and other heavy threaded drill string items from their horizontal position on the ground. To ensure that the lifted items are not unduly stressed they have a box or pin joint as shown. Lifting bails are made of cast steel. Hoisting plugs are meant to lift the drill string from the hole. They have bearings that enable rotation of the drill string components without rotating the wire rope used for lifting. This facilitates easy coupling and uncoupling of the drill string components. Hoisting plugs are seldom used in rotary blasthole drilling. Lifting bails as well as hoisting plugs for male and female threads are illustrated in Figure 6.18.
6.6.3
Blasthole plugs
Blastholes, particularly those with a large diameter, are very dangerous if left open. A serious injury or even fatal accident can result when some person or animal inadvertently steps on the cutting surrounding the blasthole that slides down to the hole. In very large blastholes someone going inside the hole to bottom cannot be unimaginable. Such incidents have happened in large diameter holes drilled for waterwells and left unattended at night. Besides, a blasthole may get filled up partially when the loosely heaped cuttings on the side slide into them. To prevent such incidents the best way is to plug the blastholes at the top. For this purpose, plastic plugs as shown in Figure 6.19, are used. They are made of plastic and have various sizes to match the diameters of blastholes.
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Figure 6.17 Thread protectors.
Figure 6.18 Lifting bails and hoisting plugs.
Figure 6.19 Blasthole plugs.
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6.6.4
Rotary deck bushing
The main purpose of rotary deck bushing is to provide horizontal stabilization to the drill string at the deck level of the blasthole drill. Rotary deck bushings are usually provided on the drill platform floor by the drill manufacturers. Their design and construction can, however, be a little less sophisticated in the sense that they may merely be non-rotating bushings. Such a simple friction bearing requires higher torque for drill string rotation and extra power, and they undergo rapid wear-out. They are also required to be lubricated from time to time. Besides, some blasthole drill manufacturers offer this as an optional item with their drills. Many manufacturers of rotary drilling accessories have developed more sophisticated rotary deck bushings that consist of an outer casing which fits on to the drill platform and the inner sleeve through which the drill pipe is passed down. Two or more rows of bearings allow the inner sleeve to rotate with drill pipe. The cross section of such rotary deck bushing is shown in Figure 6.20. When a drill pipe rotates and touches the inner sleeve due to radial vibratory movement, the inner sleeve also rotates as required. This reduces the torque requirement and wear. As less torque is spent in overcoming the friction, the drill bit gets more torque and penetration rate increases. Due to the soft inner sleeves of rotary deck bushings, the wear of the drill pipe reduces. These sleeves are relatively inexpensive and are replaceable.
6.6.5
Drill stem wrench
Early rotary blasthole drills used to have fork type spanners for uncoupling the drill pipes. They engaged at the wrench flats provided on the drill pipes. Due to several reasons these devices proved unsatisfactory. Most modern blasthole drills have one spanner to hold the drill pipe lying below the tool joint at its wrench flat and one wrench that grips the drill pipe lying above the tool joint at most of the cylindrical periphery. This wrench is called the drill stem wrench.
Top Flange to Suit the Drill
Outer Housing Adapter Bushing Inner Rotating Sleeve
Ball Bearings
Figure 6.20 Rotary deck bushing.
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Drill stem wrenches operate on hydraulic cylinders and are available in different levels of sophistication. They have jaws for gripping the drill pipes or accessories. One such wrench is shown in Figure 6.21. In the first part of the hydraulic cylinder stroke the wrench grips the drill pipe with jaws and in the second part of the stroke it loosens the drill pipe. Once the drill pipe is loosened the rest of the uncoupling can be done by rotating the drill head.
6.6.6
Bit breaker
Drill pipes usually have slots and a uniform diameter, hence while coupling or uncoupling they can be restrained from rotation by using a spanner or drill stem wrench. When a drill bit is required to be detached from the drill string, the spanner or drill stem wrench is of no use as the bit does not have slots or uniform diameter. For uncoupling the drill bit, a bit breaker as shown in Figure 6.22 is used. The bit breaker fits into the top frame of the dust hood where it is restrained from rotating. The drill bit in turn fits into the bit breaker and is also restrained from rotating. Thus, the bit can be uncoupled by counter-rotating the crossover sub or the stabilizer with the drill stem wrench.
6.6.7
Blasthole inclinometer
In early days, blasthole drills used to be equipped with simple pendulum type inclinometers. With these it was possible to align the mast for drilling inclined blastholes. The least count of such an inclinometer was about 1°.
Figure 6.21 Drill stem wrench.
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Figure 6.22 Bit Braker.
Later more sophisticated blasthole inclinometers were developed by some manufacturers. The precise angle measuring mechanism in such an inclinometer could measure the mast inclination angle up to an accuracy of 0.5° and show it by means of a digital display that could be read easily by the operator while sitting in the cab. One such inclinometer to be appropriately fitted on the mast is shown in Figure 6.23. An even more sophisticated inclinometer is shown in Figure 6.24. It comprises a vertical and two horizontal angle sensors. The horizontal sensors, fixed on the main frame of the blasthole drill, measure the angle of inclination of the main frame in the direction of drill length and drill width. The vertical sensor fixed on the mast shows the angle of the mast from the vertical direction. The digital display, central unit and control unit of such a system is kept in the operator’s cab. Such systems have even better accuracy – of the order of 0.2°. It is sometimes necessary to measure alignment and inclination of a blasthole after it has been drilled and the drill has moved away. For such measurements a hand-held inclinometer as shown in Figure 6.25 can be used. It can also be used for checking the correctness of mast alignment. In such inclinometers the magnetic compass, fitted on the top tilt arm of the instrument, can read angles with respect to north within an accuracy of 2°. For measuring blasthole inclination a torch, provided with the inclinometer and equipped with necessary fixtures, is lowered to the blasthole bottom. The torch light is then seen through the cylinder on the inclinometer. Pressing a button on the holding arm at the time when the torch light is fully visible, gives the angle of inclination with an accuracy of 0.1°.
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Figure 6.23 Inclinometer to be fitted on the mast.
Reading Unit Vertical Sensor Central Unit
Control Unit Horizontal Sensor
Power Supply Cable
Figure 6.24 Schematic of inclinometer system.
Even more advanced inclinometers, that record angle of inclination continuously while they are being lowered in the blasthole, are available.
6.6.8
Blasthole camera
Occasions when a blasthole camera needs to be used are rare, but it is essential for logging a blasthole. As it is an expensive item, it is usually hired. Most of the cameras are video type. While being lowered they store an image of the inner side of the blasthole continuously and show it on the screen.
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Figure 6.25 Hand held inclinometer.
These cameras have a cylindrical probe which consists of a light-emitting device and a camera that takes images through all 360° rotation. For taking images of the blasthole walls from the same distance, they are equipped with fixtures that align them centrally in the blasthole. Some cameras also have facilities of taking images on the front side so they can be of great help in locating a valuable item fallen into the blasthole. The image captured by the camera it is sent through a fiber optic cable to the main equipment on the surface of the ground. In this equipment, facilities for viewing the image through a television terminal as well as recording all the images on video cassette or a compact disk are available. One such camera is shown in Figure 6.26. In most cameras the images taken by the camera are immediately converted into electric signals and these signals are sent to the ground equipment. The signals themselves can be analog or digital. Analog or digital recording of these images can be done by using video cassette recorders or computer and hard disk through specialized computer software. Some advanced cameras have arrangements for being lowered into the blasthole in a controlled manner. This enables them to sense the depth, inclination, direction of inclination etc. and after analyzing the signals in a computer, a continuous log of the complete blasthole can be generated. The main control unit and the probe of such a camera, and photo log generated by it, are shown in Figure 6.27. The blasthole photo log shown in the figure is actually unwrapped in the sense that it is as if the inner surface of the blasthole is cut along a line parallel to the blasthole axis and then flattened on to a plane vertical surface. The probe containing the camera and the light-emitting diodes is shown in the left picture in Figure 6.27 while the right picture shows a photo log of a blasthole.
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Control Unit Casing and Guide
Reel
Probe
Camera
Figure 6.26 Blasthole camera system with fiber optic cable.
Figure 6.27 Digital blasthole camera and blasthole photo log.
6.6.9
Feed force measuring kit
This kit is needed for calibrating the feed force mechanism of a blasthole drill after some major repairs, or installations of additional items, or alterations in the drill. The kit is shown in Figure 6.28. The pressure unit of the kit is like a hydraulic jack and contains hydraulic oil in the enclosed space between the lower container and upper ram. For actual measurement of the feed force an adapter is fitted at the bottom of the first drill pipe instead of the drill bit. The pressure unit is placed on a firm surface below the adapter in such a way that the spherical surface of the adapter sits properly in the hemispherical groove on the top of the pressure unit. When feed force is exerted from the feed mechanism, pressure is developed in the hydraulic fluid. As the ram diameter is precisely known, the gage attached to the pressure unit gives direct reading of the force exerted. Unless the feed force mechanism on the blasthole drill is properly calibrated, the feed force measured by this kit, and indicated by the feed force reading in the blasthole, differ as shown in Figure 6.29. The chart is based on actual measurements on a well-known blasthole drill.
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Figure 6.28 Feed force measuring kit. 300
Bti Loading in kN
250 200 150 100
999.74
913.56
861.84
775.66
689.48
603.29
517.11
344.74
430.92
258.75
127.37
0.00
0
86.18
50
Pressure in kPa Indicated on the Pressure Gage in the Drill
Figure 6.29 Comparison of feed force, externally measured and indicated in a blasthole drill.
The reasons for differences can be due to one or more of the factors mentioned below. 1
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In any feed force mechanism the weight of the drill string and rotary head starts acting on the bit even before the feed force mechanism comes into action.
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3
143
A manufacturer may have calibrated the bit loading on the basis of a certain weight of drill head and drill pipes. Even if the weight of the rotary head is constant, the weight of the drill pipe changes when drill pipes are added. A manufacturer may use the weight of drill pipes with certain diameter and certain wall thickness for the calibration but the drill pipes used in actual drilling are different than those used for calibration.
It is very difficult to precisely calibrate the bit load indication system in the blasthole drills that do not have any microprocessor arrangements. When the drill is equipped with a microprocessor-based system all the reasons for inaccurate calibration can be wiped out and nearly perfect calibration can be achieved.
6.6.10
Blasthole dewatering pump
Conventional suction and discharge pumps have two limitations in dewatering blastholes as follows: 1 2
Atmospheric pressure limits their suction capabilities to 3 to 8 m. Such pumps do not throw out drill cuttings easily.
A compressed air-operated positive pressure lift pump, such as the one shown in Figure 6.30, can prove ideal for dewatering blastholes. The pump does not have any mechanical moving parts but is made up of a bellows and hose pipes. When the bellows are in a deflected state, the pump including the bellows is lowered into the blasthole in such a way that the hole below the bellows is immersed as deep in the water as possible while the bellows itself is above the water level. Compressed Air Ejection Hose for Water and Cuttings
Hose for Supplying Compressed Air to the Bellow and Blasthole Springs Meant For Deflating the Bellow
Bellow Compressed Air Pushing the Water Downward to the Bottom of the Central Hose Blasthole Wall
Water in Blasthole Under Compressed Air Pressure
Central Hose Through Which Water and Drill Cuttings are taken out
Figure 6.30 Schematic of blasthole dewatering pump.
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When the compressed air supply to the pump is started the bellows inflates and grips the walls of the hole. Thus, the blasthole below the bellows becomes airtight. As soon as this happens the compressed air is diverted to the lower side of the blasthole and compressed air pressure starts acting on the water surface in the airtight blasthole. Naturally, the water starts flowing up through the central hose till such time as the top level of the blasthole drops below the lowermost level of the central hose. After the blasthole is completely dewatered the compressed air supply is stopped. Due to the springs acting on the bellows the bellows deflates and can be withdrawn from the hole for dewatering the next blasthole or can be lowered further for dewatering the same blasthole to greater depth. This type of positive pressure lift pump can be used for dewatering very deep blastholes by using compressed air with adequate pressure. Since there are no obstacles in the central hose, the pump can lift cuttings along with the flowing water. Such a pump, however, is size-specific and can be used for holes of certain diameter range only. Advantages of positive pressure lift pumps are that they act very fast, lift cuttings of large size with the flowing water and dewater blastholes far more completely than other types of pumps. Sometimes diaphragm-type suction discharge pumps or submersible pumps are also used for dewatering blastholes.
6.6.11
Recovery tools
In very rare instances the drill pipe can break while rotary drilling is going on. In such an event a piece of drill pipe along with the drill bit remains inside the blasthole. Instances where a complete cone or part thereof has broken and remained in the blasthole, are also known. Whether the part of the drill string or the piece of broken bit is to be taken out and drilling is to be continued in the same blasthole, or whether the blasthole is to be discarded and another one drilled in the near vicinity, is a decision that changes from case to case and can be based on a little financial and technical common sense. In cases where the drill bit is almost at the end of its life and only a small part of drill pipe has remained in the blasthole, it may be worthwhile to abandon the hole, drill another blasthole at a small distance and use it for charging and blasting instead of the abandoned blasthole. If a blasthole is abandoned and a new one is chosen for charging, the subsequent blast can give rise to two possibilities as follows: 1 2
The possibility of very ineffective fragmentation in the vicinity of the abandoned blasthole and need for subsequent redrilling. The possibility of the steel pieces left in the abandoned blasthole flying to distances and damaging objects in a manner similar to a flyrock.
In many cases it is desirable to recover the broken part of the drill string. For this purpose many different types of tool, often referred to as fishing tools, are used. These tools are attached to the smaller size drill rods for ease in handling
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and are lowered into the blasthole to go inside the drill pipe bore or to cover it from outside. When pushed down and rotated, their sharp tap-like threads actually make threads in the drill pipe and the drill pipe gets attached to the fishing tool. Thereupon the rod string can be lifted so the drill pipe can be retrieved with the bit. One of the recovery tools is a recovery tap, meant to go within the central bore of the drill pipe. It is shown in left side picture of Figure 6.31. The other, meant to cover the drill pipe from outside, is called a die collar overshot. Fishing tools are case-hardened to have extremely high hardness to enable them to make threads in the drill pipes. Therefore they are sharp as well as brittle. They must be handled very carefully so as to avoid any impact with other objects.
6.6.12
Blasthole sampler
In the last three decades there has been virtually an explosion in the improvements in techniques of rotary blasthole drilling. Blasthole drills can now be precisely leveled, accurately trammed to the position of the next blasthole, and with automatic capture of all the drilling parameters such as load and torque exerted on the drill bit, penetration rate and vibration level experienced while drilling, flushing air flow etc. How this is achieved is the subject of one of the forthcoming chapters. However, such automation does not indicate the exact mineral composition of the material drilled by the drill bit. The mineral composition can only be known by visually observing the sample and later analyzing the sample in the laboratory. Therefore in many cases it becomes essential to obtain a sample of the material drilled. Collecting the samples manually is not possible because for such collection it is necessary to take pause drilling operations, open the hood and expose the collector to dust. To avoid all these hazardous and time-consuming operations a blasthole drill sampler has been developed. It is an attachment that can be fixed on a rotary blasthole drill, as shown in Figure 6.32.
Figure 6.31 Recovery taps (Left) and Die collar overshots (Right).
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Figure 6.32 Blasthole sampler.
In a recent patent filed in 2003 in the USA for the device, Harrison, the inventor of the device, has given its brief description as: “A sample collection system for collecting a drill cuttings sample from a hole being drilled during its drilling includes a stem collector surrounding the drill pipe. Pressurized air injected into the hole through the drill pipe forms an air-entrained drill cuttings stream which travels up the drill hole during drilling and which is directed by the stem collector into a conduit which directs stream to a sampling device. Sampling device samples drill cuttings stream and creates a stream of sample cuttings which is directed to a diffuser which separates the cuttings sample from the air stream. The cuttings sample passes into a sample collection chute from which it or portions thereof may be transferred into a sample container or bag. Stem collector seals against the ground surrounding the hole being drilled to direct substantially all drill cuttings to sampling device, thus substantially eliminating loss of fines prior to sampling and thus helps form a substantially representative sample. The sampling system can be mounted under the drill deck of a drilling rig.”
In prototype trials of the blasthole sampling device, the following things were noticed: 1
2 3
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The blasthole sampler may do the job of sample collection but it cannot be guaranteed that the sample is truly representative of the formation at a certain depth in the blasthole. This is because large cuttings from the bottom of the blasthole do erode the walls of the blasthole at upper parts of the blasthole and the eroded material mixes with the sample at the particular depth. In certain formations where voids exist in the ground, the flushing air is partly or totally lost. In such instance no sample is collected. In deep blastholes drilling has to be paused for short intervals.
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147
Most importantly, the blasthole sampler interferes with the blasthole flushing process and reduces the penetration rate that might be achievable without the blasthole sampler. Therefore, production department personnel do not like the idea of such a sampler being fitted on a production tool i.e. blasthole drill.
In view of all the above a blasthole sampler is not likely to be a popular attachment on a rotary blasthole drill.
6.6.13
Laser measuring instruments
In simple words these gadgets are a measuring instrument with a dedicated computer built into them. In this modern era, this instrument has put several other measuring instruments like old styled dumpy level, theodolite etc. in obsolescence. As far as blasthole drilling is concerned, a special version of a laser measuring instrument, called a burden finder has been made by MDL Systems Ltd. This handheld instrument looks like the one shown in Figure 6.33. The laser rays emitted from the instrument measure the distance of any target on which they strike. To actually measure the burden the surveyor has to stand on the bench floor in such a way that the vertical plane, in which he stands and takes all the measurements, is perpendicular to the row of blastholes. At first the surveyor aims at a point C that lies on the true crest line and stores the reading of distance OC and angle VOC. In the next step he aims at point T on the toe line and stores the readings of distance OT and angle VOT. All these points are shown in Figure 6.34. With these, angle COT is known by the instrument and it trigonometrically calculates the distance CT and angle OCT. Similarly, with the known bench height i.e. the distance CH, it calculates toe distance and face slope angle. Now referring to Figure 6.35, the distance between the true crest line (point C) and the blasthole row, which actually equals to the intended burden, is known and the inclination of the blasthole is also known. When all the above information is stored in the computer that is built into the instrument, the coordinates of the vital points are defined and the process of measuring actual burden can be started. When the laser is aimed at any point on the bench face as
Figure 6.33 Burden finder (Laser measuring instrument).
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C Crest
Face Slo
V
pe Len gth
Face Slope Angle
O
T
H
Toe Toe Distance
Figure 6.34 Measurement for basic data input.
Intended Burden
C
Actual burden
Angle of Inclination
Figure 6.35 Measurement for basic data input.
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shown in Figure 6.35 it is immediately defined in terms of its coordinates with reference to the operator. This enables the calculation of actual burden. Burden Finder does the calculations and shows the results on the screen provided on the side of the instrument. Some instruments have a communication port that enables printing the data on paper or transferring the data to a computer. Sometimes it is necessary to map the whole face of the bench. In such cases use of a handheld Burden Finder is not advisable as it is meant to take reading only in 2D. For such measurements a 3D instrument named Quarryman Pro is available. It is mounted on a tripod as shown in Figure 6.36. As Quarryman Pro is mounted on a tripod, it can turn in accurate angles in both vertical and horizontal directions. It is motorized and is turned in the horizontal and vertical axes by the motors in the instruments. With this facility and the built in computer, the Quarryman Pro can take measurements of 250 points per second
Figure 6.36 Quarryman Pro.
Figure 6.37 Profiling of a bench face with Quarryman Pro.
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on the face of the bench. The distance between the scan lines can be defined in the instrument. With large distance between scan lines, a large bench face can be mapped with somewhat less accuracy. For some part of the bench face where a high resolution scan is needed, the distance between scan lines can be kept very low. This is clear from Figure 6.37.
6.7
MISCELLANEOUS SAFETY ITEMS
Several miscellaneous items are required by the driller, the assistant, or the blaster who carries out blasting operations subsequently. Such items are expected to be provided for them by their employer i.e. the owner of the drill. Some of these items and their functions are as follows. Some more items are described in chapter 22 that pertain to blasting consumables and accessories
6.7.1
Close fitting shirt and long pants
If the clothes worn by the operators while on the drill are loose, there is a chance that they may get entangled with a lever and operate it inadvertently. Such clothes can easily get gripped within moving parts and pull the operator to cause him injury. To avoid these possibility close fitting shirts and long pants must be used by the operators.
6.7.2
Safety glasses
To avoid tiny flying objects, such as very small rock pieces, dust formed in grinding, tiny oil drops etc., hitting the eye, safety glasses to cover both the eyes from all sides are essential.
6.7.3
Safety toed shoes
These are needed by the operators to protect their feet and particularly toes from a hand tool falling on the feet.
6.7.4
Ear plugs or ear muffs
One of these two items is needed by the operators to reduce the nuisance of noise. Ear muffs cover the ear whereas the earplugs go inside the ear and close them. These items reduce noise intensity by as much as 22 db. They are shown in Figure 6.38.
6.7.5
Safety vest
These vests not only offer protection to the operator’s chest, back and abdomen but because of the orange fluorescent color the operator becomes easily visible in low light.
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Figure 6.38 Ear plugs and ear muffs.
6.7.6
Respirator
A respirator is occasionally needed by the operator to filter and allow only clean air to his nose for breathing.
6.7.7
Rain coat
These are needed when the operator or helper have to go out of the cabin to protect from rain, snow stormy wind etc.
6.7.8
Face shield
On a few occasions the driller or the assistant has to do some welding. For such activity the safety glasses are not sufficient and a face shield is required for safety of the eyes as well as the face.
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Chapter 7
Rotary blasthole drills
7.1
INTRODUCTION
The era of churn drills in the mining and quarrying industry started declining with the introduction of the rotary blasthole drill in the late 1940s. Soon after the introduction of the model 50R top drive rotary drill by Bucyrus Erie Company in 1952, most of the mines shifted to rotary drilling and the era of churn drills ended by about 1960. In the early days, rotary blasthole drills were used in surface mines and in large size quarries for drilling soft and medium-hard formations. In the late 1950s and early 1960s many equipment manufacturers started producing bigger surface mining equipment such as shovels, draglines, dumpers etc. To match with such equipment, larger blasthole drills were developed by Bucyrus Erie, Marion, Joy and Gardner Denver. Blastholes of diameters up to 381 mm could be satisfactorily drilled by such drills even in hard formations. Today gigantic blasthole drills, e.g. Bucyrus 59R or 49R, P&H 120A, Atlas Copco PV351 etc., are rated for blastholes of diameter 406 mm or larger. Their tall mast enables drilling a hole to a depth of about 20 m with the first drill pipe itself. Price of such blasthole drills can reach US$ 6,000,000. This chapter is devoted to explaining various layouts and components used on rotary blasthole drills of various types and sizes.
7.2 ASSEMBLIES IN ROTARY BLASTHOLE DRILLS A rotary blasthole drill is made by putting together many assemblies. These assemblies are very costly, have many small parts and can attain very long life. They are carefully chosen by the design engineers together with the manufacturers of the blasthole drill. Most of these assemblies can be repaired by changing a few parts. Assemblies mounted on a rotary blasthole drill have specific purposes as listed in Table 7.1. Accessories, elaborated in chapter 6, are not affixed to the blasthole drill. They are comparatively inexpensive, have no or few parts, and a relatively short life. Accessories are chosen by the user. In most cases, they cannot be repaired satisfactorily.
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Table 7.1 Common assemblies on rotary blasthole drills and purpose behind their inclusion. Assembly
Purpose
Undercarriage
Undercarriage enables the movement of the machine from one hole to the other or in some cases from one worksite to the other.
Main Frame
Excepting the undercarriage all the other assemblies of a rotary blasthole drill are mounted on the top of a very sturdy frame called main frame.
Leveling Jacks
Leveling jacks, attached to the main frame, are meant for leveling the machine after it moves to the location of the hole to be drilled.
Prime Mover
A prime mover is the main source of power. All the driven components in the machine are driven by use of this power source so that they generate desired movements of the components. Diesel engines are used as prime movers in many drills. In some very large rotary blasthole drills electric power supplied to the mine is directly supplied to the drill from where it is distributed to various components.
Air Compressor
Air compressor is meant for compressing the atmospheric air to a preset pressure and circulating it to the bottom of the hole through the drillstring components for flushing the blasthole and removing cuttings formed in the process of formation fracture.
Operator Cab Operator’s cab, located at the rear of a rotary blasthole drill, has all the necessary devices for observing and controlling all the necessary operations required to be performed during the process of drilling blastholes.
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Driver Cab
In truck mounted rotary blasthole drills the operator’s cab is located at the rear end. It is, therefore, not suitable for moving the drill on the road over long distance at higher speeds. In such case a separate driver’s cab on the front – just like that of a heavy truck – becomes essential.
Mast
Mast accommodates the rotary head assembly and the feed force mechanism that enables linear movement of the drill head with necessary feed force. By suitably aligning the mast in vertical or angled position, drill string can be moved and blasthole can be drilled in the intended blasthole alignment.
Auxiliary Winch
Most of the rotary blasthole drills are equipped with a winch. Its wire rope enables handling the heavy accessories from the top of the mast.
Rotary Head
Rotary head rotates the drill string at desired rotary speeds by exerting the required torque.
Pipe Rack
Pipe Rack – also known by many different names – is built into the mast. It enables storing and adding or removing the drill pipes to the drill string mechanically. This gives speed and accuracy in the operations.
Hydraulic System
Only a hydraulic system is capable of generating linear motion with very high force required for such operations as mast raising/lowering, leveling the machine by hydraulic jacks etc. Every blasthole drill, therefore, consists of hydraulic systems. Besides, in many rotary blasthole drills, hydraulic system also supplies high pressure hydraulic oil to many hydraulic motors to generate rotary motion.
Dust Control Equipment
During drilling operation, huge quantity of very fine dust is generated at the bottom of the hole. It is ejected out of the hole with the circulating compressed air. To prevent such dust from mixing with the atmosphere and polluting it, a dust hood is provided on the drill. However, due to many reasons the dust hood proves insufficient in preventing mixing of the dust with atmosphere. This makes it essential to incorporate dust control equipment in a blasthole drill. It is of two types viz. Water injection or Dry dust control.
Machinery House
Machinery house is an enclosure that shelters many components and assemblies within it so as to protect them from heat, cold, rain, dust and flying rocks. It also creates pleasant surrounding to the personnel to safely repair or replace the components.
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7.2.1 Assembly placement Three major groups of assemblies, viz. lower, upper, and mast, are generally recognized in a blasthole drill and are shown in Figure 7.1. Lower refers to the lowermost part of the blasthole drill. It comprises of ground support components, viz. a frame to hold the components together, two crawler tracks and the components that impart movement to the blasthole drill through the frame. Upper refers to a main frame that is placed on the top of the lower frame and contains many important assemblies viz. leveling jacks, prime mover, air compressor, operator’s cab, catwalks, hydraulic system components, dust control equipment, machinery house (if provided) etc. Figure 7.2 shows the configuration of the components on the upper. The mast is the tall, tower-like structure of a blasthole drill that is supported by the main frame. A mast consists of a rotary head, feed force mechanism, pipe changer, wrenches for coupling and uncoupling the drill pipes, auxiliary winch etc. Details of the layout of a blasthole drill and placement of assemblies are given later in this chapter.
Mast
Upper
Lower
Figure 7.1 Assembly groups in a rotary blasthole drill.
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Operator’s Cab
Engine
Air Compressor
Hydraulic Oil Tank
Dust Collector
Engine Exhaust Radiator Catwalks
Railings of Catwalks Hydraulic Oil Cooler
Figure 7.2 Configuration of assemblies on upper frame.
7.3 TYPES OF ROTARY BLASTHOLE DRILLS Rotary blasthole drills are differentiated into various types on the basis of diverse criteria. If the suitability of a rotary blasthole drill for drilling a particular class of blastholes is used as a criterion, it can be classified as conventional or special purpose. If the type of undercarriage of the rotary blasthole drill that enables its movement on the ground is taken as a criterion, it can be distinguished as either crawler-mounted or carrier-mounted. When the power source and distribution in a rotary blasthole drill is the basis of a class, diesel-hydraulic, diesel-electric, electric-hydraulic or electric-electric are the four types. Of the two words in the class names, the first refers to source and the second to the mode of power distribution. A generalized classification of rotary blasthole drills that takes into account all the four classification criteria is shown in Figure 7.3. The capability of a rotary blasthole drill in exerting feed force (FF) on the drill bit through its rotary head and the drill string components is often taken as a criterion. On this basis a blasthole drill can be placed in one of the following types: Small (100 kN < FF < 225 kN), Medium (226 kN < FF < 350 kN), Large (351 kN < FF < 500 kN), and Extra Large (FF > 501 kN)
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Blasthole Drills
Special Purpose
Conventional
These drills are diesel hydraulic type and of small size
Crawler Mounted
Diesel Hydraulic These drills can be small, medium, large or even extra large
Diesel Electric Nowadays these drills are almost extinct
Carrier Mounted
Electric Hydraulic
ElectricElectric
These drills can be small, medium, large or even extra large
These drills are usually extra large
Figure 7.3 Generalized classification of rotary blasthole drills.
7.3.1
Special purpose rotary blasthole drills
In very soft formations, e.g. a thick coal seam in a surface coal mine, a potash bed etc., blastholes of diameters ranging between 152 to 229 mm are required to be drilled at close spacing so the blast will give very good fragmentation and the blasted material can be easily loaded by wheel loaders and transported in coal haulers or dumpers. A few manufacturers have designed blasthole drills specially for drilling such blastholes. Such rotary blasthole drills are often referred to as ‘Single Pass Drills’ because they have only a single long Kelly instead of a drill pipe. A tricone or a drag bit is attached to the bottom of the Kelly through a sub. The Kelly is rotated by means of a rotary table placed on the main frame of the drill. Load on the bit is exerted through the Kelly which is pushed down through the feed force mechanism built into the mast. The use of such blasthole drills proves economical because the conventional blasthole drills, meant for attaining an equal hole depth in a single pass, are too expensive. A typical single-pass blasthole drill, the Atlas Copco model DMLSP, is illustrated in Figure 7.4. Different models of single pass drills are available to suit different hole depths and diameters. It can be easily observed that the single pass drill shown in Figure 7.4 looks far smaller but taller than the conventional drill capable of drilling similar hole diameters shown in Figure 7.1.
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Rotary drilling and blasting in large surface mines
Figure 7.4 Special single pass blasthole drill.
The mast of a special single pass rotary blasthole drill is very tall to accommodate the long kelly. The cross section of the mast is very small for the following reasons: 1 2
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The drill does not have any pipe changer and the mast is not required to support it. The rotary torque on the drill bit is exerted by the rotary table fixed on the main frame instead of a moving rotary head in the mast. Thus, the mast need not be designed for heavy twisting reaction to the torque exerted by the moving rotary heads as in conventional drills.
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3 4
5
159
As the Kelly drills the blasthole in a single pass, no coupling and uncoupling wrench is involved. As the drill is basically meant for drilling in soft formations, limited feed force is to be exerted. A 250 mm blasthole can be drilled with about 225 kN feed force. If the same size blasthole is to be drilled in very hard formations the rotary blasthole drill must be capable of exerting a feed force of about 350 kN. Feed force is exerted by hydraulic motors and chains which weigh far less as compared with the hydraulic cylinder and chain or rope mechanism in conventional drills.
The mast designed by taking into consideration all the above factors is lightweight, and the drill can be made stable without having any need for heavy main frame members, ballast, or a wide and long crawler base for the purpose of stability of the drill. Further, as a special single-pass rotary blasthole drill is not meant for drilling holes over a wide range of diameters, there is no choice of many alternative engines or compressors. The designer can therefore place the assemblies on the main frame very closely. Some time ago, it was thought that blasting of horizontal blastholes at the bench toe in addition to normal vertical blastholes would result in better fragmentation, smoother bench surface, etc. To cope with such requirement of drilling horizontal blastholes, a special twin mast drill, as shown in Figure 7.5, was developed by Joy Manufacturing Company during
Figure 7.5 Twin mast rotary blasthole drill for drilling horizontal blastholes.
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Rotary drilling and blasting in large surface mines
early 1970s. This drill could simultaneously drill two blastholes of size up to 270 mm in hard formation. It was very heavy as it had to withstand high horizontal reaction. Such drills, however, never became popular because: 1 2
Small horizontal blastholes of 100 to 127 mm diameter were found to be better for the purpose than the 250 to 270 mm blastholes drilled by using the twin mast drill. Horizontal holes of diameter to 127 mm could be drilled at a much faster speed by small sized, comparatively inexpensive rotary percussion drills rather than the bulky and heavy twin mast rotary blasthole drill.
Some designers also thought of a rotary blasthole drill with two masts on one frame so two blastholes can be drilled simultaneously. This concept, however, proved impractical for large rotary blasthole drills. One rotary blasthole drill with two masts meant for drilling two small diameter holes simultaneously is manufactured in South Africa by Smith Capital Equipment.
7.3.2
Conventional rotary blasthole drills
Conventional rotary blasthole drills are designed to be useful for drilling holes over a wide range of drilling parameters viz. hole diameter, hole depth, hole inclinations, formation hardness etc. Almost all the manufacturers have a range of models of such drills. They can be equipped with many options available to suit specific drilling conditions. 7.3.2.1
Crawler mounted rotary blasthole drills
Crawler-mounted blasthole drills are very popular. More than 98% of the rotary blasthole drills purchased all over the world are crawler type. These drills are very sturdy and rugged. In most of the cases they are used for round-the-clock operation throughout the year. Primarily they are used in surface mining operations. 7.3.2.2
Carrier mounted rotary blasthole drills
In the 1970s and 80s a few carrier-mounted blasthole drills were chosen by some mines for surface drilling operations. For many reasons they proved unsatisfactory. Today hardly anybody ventures in using carrier-mounted blasthole drills for such mining operations. Carrier-mounted rotary drills are very suitable for drilling waterwells that are usually located at long distances from each other. Therefore, on-the-road mobility of the drill is of prime importance. Although termed as rotary, these drills are more often used for rotary percussion drilling with down-the-hole hammers. Many manufacturers publish brochures of carrier-mounted rotary blasthole drills but they are essentially waterwell drills that can also be used for drilling blastholes. In any case, the size of these so-called blasthole drills is small. They can be used for blastholes of diameters up to about 170 mm. An extra large carrier-mounted rotary blasthole drill, Atlas Copco Model T4BH, that is sometimes used for blasthole drilling, is shown in Figure 7.6.
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161
Figure 7.6 Carrier mounted rotary blasthole drill.
7.3.2.3
Comparison of crawler vs. carrier
Why almost all the users select a crawler-mounted rotary blasthole drill in preference to carrier mounted blasthole drill can be understood from the detailed comparisons of the two types made in Table 7.2.
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Table 7.2 Comparison between crawler mounted and truck mounted rotary blasthole drills. Aspect
Crawler mounted
Road Worthiness Virtually nil due to large size, huge weight and slow speed of the crawler undercarriage.
Carrier mounted Designed to travel on most of the paved roads by taking into account proper weight distribution on the axles and other essential features.
Ability of Long Travel
Very poor due to very low travel speed Very good due to high travel speed that is often lower than 4 km/h that ranges between 50 to 70 km/h
Power Source
Diesel or electric power sources are easily possible
Can only be diesel driven. Electric power can not be used.
Width
Wide, usually between 3 m for small drills to 8.5 m for extra large drills
Narrow, often less than 2.4 m or even in extreme cases less than 3 m as the drills are required to travel on regular paved roads.
Weight Range
Ranges between 25000 kg for small Limited to about 60000 kg even in drills to 184000 kg for extra large drills. case of largest drills. The drill must adhere to 20000 kg per axle as per road worthiness for usual roads.
Rough Terrain Ability
Excellent due to crawler undercarriage that gives low ground pressure of the order of 100 kN/m2 and widely spaced long tracks.
Very poor due to wheeled undercarriage that gives high ground pressure of the order of 600 kN/m2 and narrow width.
Feed Force Ability Can exert very high feed force due to their heavy weights. The usual range is between 110 kN for small to 600 kN for extra large drills
Can not exert high feed force due to limited weights. The usual range is between 80 kN for small to 300 kN for extra large drills.
Mast Height
Masts can not be tall. Heights are usually limited to 12 m.
Masts can be very tall. Heights can reach to height in excess of 27 m in case of extra large drills.
Longest Drill Pipe Long drill pipes of about 18 m length Drill pipes are short. Limited to about are possible in many medium size drills. 7.5 m in almost all the drills. Can be even 21 m in case of extra large drills. Dust Suppression Both dust collector and water injectors Water injection is easily possible. are easily possible. Dust collectors are difficult to mount. Stability
Excellent because heavy weight in the lower and upper. Can travel with raised mast on the rough surface at the top of mine bench.
Maintenance Cost Relatively low maintenance cost due to crawler under carriage. Most of the machines can be equipped with automatic lubrication system.
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Poor because of limited weight and limited width. Traveling with raised mast on the rough surface of a mine bench is dangerous. Relatively high maintenance cost due to tired under carriage. Machines usually can not be equipped with auto lubrication system
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Rotary blasthole drills
7.4
163
LAYOUTS OF ROTARY BLASTHOLE DRILLS
The layout of assemblies and components of a rotary blasthole drill differs from manufacturer to manufacturer as well as model to model. The objectives behind an ideal layout are given below. 1 2 3
4
There should be adequate space for repair or maintenance of assembly/component. Removal or replacement of the assembly or component should be possible without removing other assemblies or components. The length of pipes, hoses and cables used for carrying the hydraulic oils, compressed air or electric current should be minimum so the replacement cost becomes low and pressure or voltage losses are minimized. The position of the combined center of gravity of all the assemblies/components should be such that the blasthole drill will remain sufficiently stable in different positions of the mast that prevail while carrying out drilling operations, or while the blasthole drill moves from one location of hole to another.
Layouts of typical diesel engine-driven carrier-mounted and crawler-mounted blasthole drills are shown Figure 7.7. Figure 7.8 shows layout of a typical all-electric rotary blasthole drill.
7.5
DETAILS OF ASSEMBLIES
There is no specific standard that stipulates what assemblies or devices should be provided as standard and what should be optional in a blasthole drill. In some rotary blasthole drills some of the assemblies mentioned in the table are not provided as standard but are available as options at extra cost. Here is some additional elaboration.
7.5.1 Assemblies in the lower group This group is also called an undercarriage. It includes the items required to support the upper and the mast and enable movement of the blasthole drill. As said earlier, crawler and carrier are the two types of undercarriage used in rotary blasthole drills. 7.5.1.1
Carrier undercarriage
Carrier undercarriages are not very common in rotary blasthole drills. They are described briefly. The prime objective behind choosing a carrier undercarriage to a drill is to make it roadworthy and capable of traveling at fast speed on paved roads. A crawler undercarriage is never considered as roadworthy as the point loads between track and the surface of a paved road destroy the surface of the paved road. Further, for many reasons, they cannot be easily made to travel at fast speed. A carrier undercarriage, since it has to travel on paved roads, must comply to all the regulations stipulated for being roadworthy. These regulations change from country to country. Some common regulations are given in Table 7.3.
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Radiator
Air Receiver
Hydraulic Pump Drive
Operator’s Cab
Operator’s Cab Cab Air Cond., Heating and Pressurizing Unit
Leveling Jack Cylinder
Valve Bank
Hydraulic Tank Diesel Hydraulic Engine Pumps
Air Compressor
T Tank
Leveling Jack Cylinder
Air Cleaners
Hydraulic Oil Cooler
Diesel Engine Hydraulic Oil Tank Dust Collector Front Hydraulic Leveling Jack Engine Oil Filter Mast Raising Lube PLC Compressor Oil Cooler Tank Control Diverter Cooler Leveling Jack Cylinder Cylinders Engine Drill Deck Leveling Jack Cylinder Muffler
Air compressor
The Layouts in this Figure are not to the Same Scale
Driver’s Cab
Front Hydraulic Leveling Jack Fuel Tank
164 Rotary drilling and blasting in large surface mines
Figure 7.7 Layout of a carrier mounted (top) and crawler mounted ( bottom) rotary blasthole drill.
Carrier undercarriages used for rotary blasthole drills look like a common three or four axle truck. The normal heavy duty trucks, made by heavy vehicle manufacturers, are generally not suitable as a carrier to a blasthole drill. This is because the frame of these
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Figure 7.8 Layout of an all electric crawler mounted rotary blasthole drill.
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Bit Viewing Hatch – Hydraulically Operated
Hydraulic Breakout Wrench
Dry Dust Control
Leveling Jack – Hydraulically Operated
Service Platform
1
2
3
4
5
Item Description
Transformers Dry Type Compressor Oil Cooler
8 9
Heated and Insulated Cooler Compartment
Hydraulic Oil Cooler and Guard
7
10
Water Injection System for Dust Control
6
Item Description
Main Air Compressor Compressor Drive Motor Hydraulic Pumps and Gearcase Pipe Racks Operator’s Cab Hydraulic Oil Reservoir Electrohydraulic Valves
13 14 15 16 17 18 19
25
24
23
22
21
Automatic Lubrication System
Hoist /Pulldown Drive Control Cabinet
Rotary Drive Control Cabinet
Programmable Controller
Low Voltage Motor Control Center
Air Compressor Motor Control
Boarding Stairs
12
20
Cable Reel
11
Item Description
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Rotary drilling and blasting in large surface mines
Table 7.3 Restrictions imposed on carrier undercarriages for being worthy of traveling on paved roads. Aspect
Restriction on carrier undercarriage
Road Support
Road support must essentially be rubber tired wheel type. Steel crawlers are not allowed as they damage the road surface. Maximum permitted axle weight is usually of the order of 20000 kg. For many reasons the front axles of the carrier are restricted to an axle loading of about 12000 kg. With these restrictions a rotary blasthole drill carrier is required to have minimum of three or four axles. Usually the maximum travel speed of the vehicle is desired to be in excess of 80 km/h. Lower speeds may be allowed but are undesirable from the viewpoint of travel time. Commonly height of the vehicle in excess of 4.5 m is not allowed The width of the vehicle should to be 2.5 m. Width of a maximum of 3 m is allowed as special case and if vehicles with 3 m width are to travel on roads special arrangements may be necessary. Maximum permitted length for a vehicle is of the order of 10 m. Longer vehicles may be permitted travel but only with special travel arrangements. All the regulations regarding front lights, turn signals, side marker lights and special marker lights The drivers cab has to be in such a position that he can see the road ahead in the best possible manner to drive the vehicle efficiently. In a rotary blasthole drill drivers cab is invariably separate because the operators cab is located at the rear end from where the road view is almost totally invisible due to several assemblies mounted on the deck. Drivers controls such as steering, brakes, accelerator, gear selector etc., along with all the indicators have to be in the drivers cab. Drilling controls are required to be in the operators cab located on the rear.
Axle Weights
Travel Speed
Height Width
Length Lighting Drivers Cab
Controls
trucks is not very rigid and too weak to withstand the stresses developed during drilling operations. Hence drill manufacturers themselves manufacture the carrier or get it manufactured as per their design. A few more points about carrier undercarriages are presented in a subsequent chapter on design criteria of blasthole drills. 7.5.1.2
Crawler undercarriage
A crawler undercarriage comprises two crawler tracks. A photograph of one such track is shown at the top of Figure 7.9. Each of these tracks contains many parts shown schematically in the lower drawing in Figure 7.9. These crawler tracks are joined together by means of a rigid cross frame in the case of a slewing undercarriage, or a frame having a pivot type and a fixed type cross beam to make the undercarriage pivot type. 7.5.1.2.1
Slewing crawler undercarriage
A typical slewing undercarriage is shown in Figure 7.10 where the rigid cross frame is clearly visible.
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Drive Sprocket
Track Pad Chain
Lower Load Bearing Rollers
Idler Roller
Track Beam
167
Idler Sprocket
Slack Adjuster Mechanism
Figure 7.9 Side frame of a crawler track.
Figure 7.10 A slewing crawler type undercarriage.
In the center of the cross frame is a large diameter circular roller bearing. The upper frame is mounted on this bearing. With such an arrangement the upper can be rotated in a 360° angle with respect to the lower crawler tracks. In certain cases such maneuverability enables the drilling of two blastholes without moving the drill. Besides this, a slewing crawler undercarriage offers good advantage in terms of placement of the drill on a narrow surface mine bench that has few rows and many blastholes in each of the rows.
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This can be explained by means of Figure 7.11. The left side illustration shows placement of the drill having a slewing undercarriage. The arrow shown in the figure is the direction of drill travel and also represents the movement of the center of gravity of the drill. This arrow lies at a distance D from the edge of the bench. It is also possible to make a drill with pivot type undercarriage to travel in this manner as shown by position B in the right side illustration in the figure. However, in this case the center of gravity of the drill, which is at a distance d from the bench edge, happens to be very near the edge. This greatly increases the danger to the drill on account of the slope stability. Hence, a drill with pivot type undercarriage has to be placed like the one shown by position C. The advantage of having a slewing crawler undercarriage, however, is not very great as it is only for drilling in the first row. For drilling blastholes in the second row, the drill with pivot type undercarriage can also be placed with crawler tracks as in position B, because in such case the distance D happens to be sufficiently great. Despite the above advantage, almost all the rotary blasthole drills being manufactured now are with pivot type undercarriages. These undercarriages have many advantages over the slewing undercarriages, as below.
Drill With Slewing Undercarriage
Drill With Pivot Beam Undercarriage
Elevation View
Elevation View
Position A
Position C
Plan View
Plan View
D
Blasthole Positions Drill Travel Direction
Position B
d Drill Travel Direction
Figure 7.11 Advantage in placement of a drill equipped with a slewing crawler undercarriage.
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1 2 3
4 5
169
A slewing undercarriage must be able to revolve within the space between hydraulic jacks. This imposes limitations on the crawler track size and blasthole drill stability. Design of the blasthole drill with slewing undercarriage is more difficult as the designer has to take into account different slewing positions while evaluating drill stability. A special hydraulic swivel has to be used between the junction of upper and lower to facilitate the flow of hydraulic oil to the propel motors in the undercarriage in all the slewing positions. Such swivel greatly obstructs the space available on the upper frame. Maintenance efforts required in the drill with slewing undercarriage are more than those required in case of pivot type undercarriage. In slewing undercarriages the crawler tracks are rigid and do not ensure uniform ground contact pressures as in the pivot type undercarriage.
In fact some manufacturers used to have slewing undercarriage for their rotary blasthole drills but later shifted to pivot type undercarriages. 7.5.1.2.2
Pivoted crawler undercarriage
In a pivot mounting, the main frame of a blasthole drill is supported on two crawlers through one fixed cross beam and one pivot type cross beam as shown in Figure 7.12. Such arrangement enables the crawler track to remain in contact with the ground even if the ground is not level because the crawler tracks are free to rotate around the fixed beam within the limits imposed by the pivot beam that is supported on the two crawlers through pins as shown in the figure. 7.5.1.2.3
Crawler track components
As the need for crawler-mounted equipment has grown, a few manufacturers have started making crawler tracks that can be used for a variety of equipment. They are available for supporting gross weights up to even 120000 kg. These ready-made crawler tracks are used for most blasthole drills. When, in case of extra large blasthole drills, the weight of the drill exceeds the load limits, manufacturers choose even sturdier tracks specially made for the purpose. Most of these manufacturers also make
Track Beam
Pin Support
Pivot Beam Pivot Pin
Rigid Beam Crawler Tracks
Figure 7.12 A pivot crawler type undercarriage.
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Rotary drilling and blasting in large surface mines
large crawler equipment e.g. shovels, and have the necessary manufacturing and heat treatment facilities for making suitable components. All the components in a crawler track are made from special alloy steels that are deep case hardened. An undercarriage may be slewing type or pivot type, but the crawler tracks comprise the same components as schematically shown in Figure 7.9. Features and functions of these components are as follows. TRACK BEAM
A track beam is usually made up of a welded box section. For the purpose of additional weight and rigidity the box can be filled with a special type of concrete that contains scrap steel pieces treated to avoid corrosion. A track beam supports the rigid frame from the side as shown in Figure 7.10 or fixed and pivot type cross beams from the underside as shown in Figure 7.12. IDLER AND LOAD ROLLERS
The upper side of the track beam supports a few idler rollers. These idler rollers in turn support the crawler pads from underneath. On the lower side of the track beam many load rollers are fixed. These rollers lift the beam on the crawler pads. They also ensure smooth and straight line movement of the crawler pads. SPROCKET
On one side of the track beam a sprocket is mounted. The sprocket is also called a drive tumbler. A sprocket has many teeth. These teeth interlock with the track pads. When the sprocket is rotated the track pads are forced to move. As these track pads are offered resistance from the ground, in effect the blasthole drill mounted on the undercarriage moves. In modern blasthole drills the rotary motion to the sprockets is imparted by means of a high torque hydraulic motor through a gear train like the one shown in Figure 7.13 or through an epicyclic gear box that is more compact. IDLER TUMBLER
On the other end of the track beam an idler tumbler is mounted. This idler tumbler also supports the track beam like the load rollers and sprocket. In addition, the tumbler is also equipped with a spring mechanism that keeps the tumbler under a pushing force so the crawler track does not lose its desired tightness. One type of spring arrangement is shown in Figure 7.14. TRACK PADS
A chain of track pads linked together forms a crawler track. When the crawler track is made from shovel type track pads as shown in Figure 7.12, the track pads are directly linked together by means of pins. The large bearing area between the pads and pins results in a very sturdy crawler track.
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171
Hydraulic Motor
Gear Train
Track Pad
Figure 7.13 Hydraulic motor and gear train for track drive.
Springs
Idler Tumbler
Figure 7.14 Spring arrangement for tightening the crawler track.
Track manufacturers usually make grouser type track pads made from cold rolled steel that is cut and heat treated. These track pads are bolted onto the track links that are joined together by means of springs to make a crawler track. Three types of grouser pads are available viz. single grouser, double grouser and triple grouser, as shown in Figure 7.15. Single grouser pads have excellent grip with the ground due to larger grouser height but lower bending strength. They are meant
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Rotary drilling and blasting in large surface mines
One Large Grouser
Holes for Bolting the Track Pad on Track Link
Two Medium Grousers
Three Small Grousers
Figure 7.15 Three types of track pads.
for dozing applications in softer ground. For rotary blasthole drills the high bending strength of the grousers is more important than the grip as the drills rarely work in soft grounds. Obviously, most of the rotary blasthole drills have triple grouser track pads. The number of pads in each of the track is so chosen that the area of contact of all the track pads on both the crawler tracks exerts pressure less than 50 to 100 kPa on the ground. For reduced maintenance, most of the ball or roller bearings used in crawler tracks are lifetime lubricated and sealed. Travel speed of most of the crawler blasthole drills ranges within 1 to 4 km/h. Higher travel speeds enable quicker placement of blasthole drills as well as bench-to bench-travel movement. However, speeds more than about 3 km/h do not offer much advantage as the travel distances are limited to about 6 to 15 m on average. Making the drill travel at high speed is very dangerous from the viewpoint of the stability of the drill. Larger blasthole drills with tall masts have lower travel speeds. Maximum grade ability of a crawler-mounted blasthole drill is about 30% with lowered mast.
7.5.2 Assemblies in the upper group All the assemblies in the upper group are mounted on a frame called the main frame. Even the mast that contains all the assemblies in the mast group, is also mounted on the main frame.
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173
In the case of a diesel engine-driven blasthole drill, the important assemblies in the upper group are: 1 2 3 4 5 6 7 8 9 10 11 12
Main Frame Leveling Jacks Prime Mover Compressor Hydraulic System Automatic Lubrication System Radiators and Oil Coolers Machinery House Dust Control System Operator’s Cab Driver’s Cab Fuel Tanks
If the blasthole drill is electrically operated the diesel engine is replaced by an electric AC motor that is suited to run directly on the high voltage electric supply in the mine. For all electric drills, some functions are better performed by AC electric supply and others need DC electric supply. Low power motors need lower supply voltage. Thus, a transformer and a converter are needed. These are also mounted on the drill. All electrically operated blasthole drills usually have transformers to reduce the high voltage of mains to an appropriate voltage for various motors, and also a converter to convert some electricity from AC to DC. They also have a provision for a cable reel mounted on front side of the drill. Details of the above assemblies/components are given hereunder individually. 7.5.2.1
Main frame
During blasthole drilling operations the weight of all the components in the three groups is borne by the main frame and transmitted to the ground through hydraulic jacks. The main frame is also subjected to heavy vibrations generated in drilling. Therefore, the frame not only has to be strong but very rigid as deflections in the frame give rise to high bit wear and blasthole deviations that are undesirable for drilling and subsequent blasting. A typical main frame is made by using two extra thick I/C sections welded together by means of cross beams. For many large drills I/C sections of appropriate size are not readily available. They have to be made by welding three pieces thereof. To ensure adequate support to the many assemblies that are to be placed outside the frame, many cantilever beams are also welded to the frame. Vertical members that support the mast, hydraulic jacks, machinery house, operators cabin etc. are also directly welded to the main frame. In most cases the steel chosen for manufacture of the main frame is ASTM A36 or equivalent. Often main frame is provided with catwalks, also called walkways, that are required for easy access to the assemblies/components mounted on the main frame. Anti-skid plates or thick wire gages are used for the floor of catwalks and mainframe.
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Rotary drilling and blasting in large surface mines
The main frame in turn is fixed onto the crawler undercarriage and looks as shown in Figure 7.16. 7.5.2.2
Leveling jacks
Rotary blasthole drills exert very heavy feed force on the drill string. A crawler track is not ideally suited to resisting the heavy reactions generated in drilling. Further, a blasthole drill resting on uneven ground through the crawler track rarely gives a firm horizontal position to the main frame. As all the mast alignments are designed for a perfectly horizontal position of the main frame, it is essential to ensure that while drilling, the main frame is in a firm horizontal position. Leveling jacks facilitate horizontal positioning of the main frame. A blasthole drill is equipped with two jacks on the rear side i.e. the mast side and either one or two jacks on the front side. In electrically-operated drills two jacks on the front side are often essential to accommodate the cable reel. Leveling jacks are operated hydraulically. For guiding the operator in accurate leveling, two tube type, and one circular, bubble levels are provided in the operators cab. The operator controls the flow of hydraulic oil to the jacks and extends/retracts them as required to level the drill. Hydraulic valves used for the hydraulic jacks are an automatically locking type so even if the tube supplying hydraulic oil to the jack leaks, the hydraulic oil confined within the cylinder of the jack does not flow out and the jack remains in the same extended position. Usually the length of hydraulic jacks is so designed that they enable leveling the drill even if the ground is sloping at an angle of 10%. To prevent the penetration of the jack rod into the ground a thick circular steel pad is provided at the bottom of the jack pad. The joint between jack pad and jack rod is knuckle-type so the pads can easily align with the ground surface without inducing stresses on the pad or rod. The diameter of the jack pads is so designed that the contact pressure between jack pad and ground is between 200 to 350 kPa.
Operator’s Cab Mast Support
Transverse Beam
Crawler Undercarriage Cross Beam Support for Leveling Jack Figure 7.16 Main frame of a blasthole drill mounted on crawler undercarriage.
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7.5.2.3
175
Prime mover
Depending upon its type and size, the power requirement of a rotary blasthole drill ranges between 200 kW to 1300 kW. Diesel electric drills are rarely made nowadays. For diesel-hydraulic, electrichydraulic and electric-electric crawler drills, typical schematic power source and distribution diagrams are shown in Figures 7.17, 7.18 and 7.19.
Air Compressor
Batteries
Diesel Engine
Starter
Alternator
Controls
Miscellaneous Operations Fans, Lighting and Air Conditioning in Cab and Machinery House.
Hyd. Pump 1
Hyd. Pump 2
Drill String Rotation, Propel Left Track, Leveling Jacks
Drill String Feed Propel Right Track
Hyd. Pump 3
Work Lighting Auxiliary Operations Mast Raise/Lower, Mast Brace, Carousel Rotation, Hyd. Spanner in/out, Breakout Wrench in/out, Breakout Wrench Lock, Breakout Wrench Rotation, Hydraulic Oil Cooling, Water Injection Pump, Oil Injection Pump, Auxiliary Winch, Radiator Cooling Fan, Other Auxiliary Operations
Figure 7.17 Power source and distribution schematic for diesel – hydraulic blasthole drill. High Voltage Supply SD Transformer Batteries
Electric Motor
Starter
Controls
Miscellaneous Operations Fans, Lighting and Air Conditioning in Cab and Machinery House.
Air Compressor
Hyd. Pump 1
Hyd. Pump 2
Drill String Rotation, Propel Left Track, Leveling Jacks
Drill String Feed Propel Right Track
Hyd. Pump 3
Work Lighting Auxiliary Operations Mast Raise/Lower, Mast Brace, Carousel Rotation, Hyd. Spanner in/out, Breakout Wrench in/out, Breakout Wrench Lock, Breakout Wrench Rotation, Hydraulic Oil Cooling, Water Injection Pump, Oil Injection Pump, Auxiliary Winch, Radiator Cooling Fan, Other Auxiliary Operations, Cable reel Winding
Figure 7.18 Power source and distribution schematic for electric – hydraulic blasthole drill.
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High Voltage Supply SD Transformer
Starter
AC/DC Converter
High Voltage Motor (AC) Rotation Motor (DC)
Hoist and Pulldown Motor (DC) Propel Left Track
Low Voltage Motors (AC)
Air Compressor
Hyd. Pump 1 Propel Right Track
Hyd. Pump 2 Leveling Jacks
Water Injection Pump, Oil Injection Pump, Radiator Cooling Fan, Cab and Machinery House Pressurization, Cable Reel Winding Converter
Batteries
Controls
Fans, Lighting and Air Conditioning in Cab and Machinery House, Work Lighting Auxiliary Operations Mast Raise/Lower, Mast Brace, Pipe Rack Movements, Hyd. Spanner in/out, Breakout Wrench in/out, Breakout Wrench Lock, Breakout Wrench Rotation, Auxiliary Winch, Other Auxiliary Operations
Figure 7.19 Power source and distribution schematic for electric – electric blasthole drill.
How much power is required for propelling and the different drilling operations is a topic discussed in a later chapter on design criteria for blasthole drills. In carrier-mounted rotary blasthole drills, the diesel engine used for driving the carrier is used for actuating all the devices used in drilling. Power from the carrier engine is taken through a power take off, also called the transfer case. A typical power train diagram for a carrier-mounted rotary blasthole drill is shown in Figure 7.20. In many such drills the power available from the carrier engine is sufficient for actuating all other assemblies except the compressor which is driven by a separate engine. 7.5.2.4
Compressor
Compressors currently used on rotary blasthole drills are oil-flooded rotary screw type or in a few cases rotary vane type. Whichever the type may be, a compressor consumes more power than any other assembly or component on a rotary blasthole drill except perhaps the rotary head. A compressor is directly driven by the diesel engine or a high voltage high power electric motor. As the two components are very heavy they are transversely mounted at the center of the main frame for stability. Other components associated with an oil-cooled compressor are air oil separator and oil cooler. They are separately mounted on the main frame. More details in respect of the working of the compressors are given in chapter 8 in this book.
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Mast Rotary Head Pipe Changer Rotary Head Drive Assembly Twin Shaft Transmission Clutch Assembly Triple Gear Pump Air Compressor Transfer Case i.e. Power Take Off Main Transmission Driver’s Cab
Operator’s Canopy Rear Steering Wheels
Front Steering Wheels
Figure 7.20 Power train in a carrier mounted blasthole drill.
7.5.2.5
Hydraulic system
Every blasthole drill has a hydraulic system. Major items in the hydraulic system are hydraulic reservoirs, filters, pumps, tubing, valves, motors, cylinders etc. Pumps and motors used in the hydraulic system are often variable displacement type. Hydraulic pumps are directly driven from the other end of the diesel engine or the main electric motor. In many cases special pump drives, like those shown in Figure 7.21 are used rather than double or triple pumps mounted on the same central shaft, so they can be serviced separately. Wherever possible metal tubing is used as it give less pressure loss and seldom requires replacement. Most of the valves are electro-hydraulic type. This eliminates the need for linkages that are costly and difficult to repair or maintain. A typical circuit diagram of hydraulic system of a electric-electric blasthole drill is given in Figure 7.22. Blasthole drills with hydraulic power distribution requires many more components. 7.5.2.6
Bit lubricating system
Tricone bits used for rotary blasthole drilling are of two types viz. air cooled or sealed bearing. In air cooled bits, even if the bearings are cooled by compressed air flowing through the bits, the requirement for lubrication of the bearings remains unfulfilled. For this reason the provision of a bit lubrication system on a blasthole drill becomes essential. A typical bit lubrication system consists of a tank and a timer-controlled adjustable discharge pump for injecting lubricant oil into the compressed air piping through a nozzle.
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For Mounting Four Hydraulic Motors
For Mounting Two Hydraulic Motors
Clutch Assembly
Figure 7.21 Hydraulic pump drives.
Figure 7.22 Hydraulic circuit diagram for a typical blasthole drill.
When the pump develops very high pressure in the delivery line, the nozzle discharges the oil in the air stream in the form of very fine particles. This oil mist travels along with the compressed air to the tricone bit bearings. Many parameters such as hole diameter, compressor discharge, type and viscosity of lubricant etc. are required to be considered while designing a good bit lubrication system. 7.5.2.7
Automatic lubrication system
Despite use of lifetime-lubricated sealed bearings in most of the locations, many components of rotary blasthole drills, such as the gear assembly in the rotary head, slack adjusting assembly, sprockets in the crawler tracks etc., need to be externally lubricated.
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It is a very tedious job to reach near the lubrication points and lubricate all these parts by using a grease gun. Many points, particularly those on the mast and components therein are accessible only with great difficulty. Further, during such manual lubrication the drilling operation is required to be stopped for the purpose of safety. To overcome this difficulty many blasthole drills have an automatic lubricating system as shown schematically in Figure 7.23. The following is the general idea of the working of such a system. Primarily the system comprises a central pump and control station located on the main frame of the drill. This station has a lubricating oil reservoir from where the oil is fed to a lube line by a feed pump. The control unit ensures that the pressure of lubricant in the line is kept at a certain predetermined level so the lubricating oil remains without any voids or bubbles. The presence of bubbles in the oil creates an air lock and
Motorized Pump High Pressure (3.5 MPa) High Volume (4 L/hr) 24 VDC Battery Operated Inbuilt Filter, Pressure Gage, Relief Valve Excess Pressure Switch Electrical Vent Valve Injector Adjustable Output Visual or Electrical Indicator Independent and Serviceable
Progressive Distributor Output Adjustable in Multiples Contamination insensitivity Special Relief Valve to Identify Blocked Line Choice of Output Connection Electric Cable Lubricant Piping
Figure 7.23 Central pump and control station of the lubricating system.
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stops the flow of oil in the line. A timer in the station operates the pump and increases the pressure in the manifold to certain higher levels only at predetermined time intervals. The tubing attached to the manifold carries oil up to the point where lubrication is required to the parts. Near every point of lubrication, a spring loaded injector is placed. This injector opens its nozzle only when the pressure in the manifold rises to preset higher levels. Lubricant released by the injector in this manner is taken up to the point of lubrication by means of a small tube and released. The pressure setting of the injectors can be adjusted as shown in Figure 7.24 so that the right quantity of lubricant is fed to the point from the manifold by means of small tubes as shown in Figure 7.25. A system monitor lights located in the operator’s cab indicates the status of the system by sensing the pressure in the manifold. If sufficient pressure is not being developed, the operator is warned through an audible alarm as well as a blinking red light signal. In some blasthole drills two or more pumping and control stations are provided, so the time intervals between lubrications can be kept flexible. The number of points which an automatic lubrication system lubricates automatically is dependent upon the design of the rotary blasthole drill. In many large blasthole drills the number of points is in excess of 100 as listed in Table 7.4.
Lubricant Tube from Injector
Lubricant Tube to Injector
Injectors
Figure 7.24 Lubricant injectors.
Figure 7.25 Lubricant tubes on a track beam.
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Table 7.4 Points of lubrication on a rotary blasthole drill. Assembly of the drill
Lub. Points
Details
Crawler Undercarriage
35
Drive tumbler, Fixed axle bushings, Lower rollers, Slack adjuster, Equalizer beam, Propel chains
Hydraulic Jacks
16
Upper and lower guides
Propel Machinery
4
Mast and Platform
33
Tool wrench cylinder guides and bearing blocks, Sprocket pins, Wire ropes, Pulldown chains, Hoist cylinder pins, Curtain rope pulleys, Angle drilling pipe positioner, Idler sprockets, Shipper shaft bearings, Rack and pinions, Mast guides
Rotary Head
11
Main Machinery
20
Air seal, Adjusting collar, Gear case, Seal, Rotary coupling, Pipe joint Propel gear case, Propel shaft, Propel sprockets, Propel brakes, Hoist/Pulldown gear case, Hoist sprocket, Planetary unit, Sprockets, Auxiliary winch drum, Auxiliary brake lever shafts
7.5.2.8
Intermediate propel shaft inner and outer bearings
Radiators and oil coolers
Almost all the diesel engines used on blasthole drills are the water-cooled type. Water circulated for engine cooling needs a large radiator to ensure adequate cooling of the engine. Besides this, the compressor oil also needs a cooler. To ensure that the radiators and coolers transfer heat to atmospheric air, they are placed on the edge of the main frame. 7.5.2.9
Machinery house
The importance of having a machinery house on a rotary blasthole drill for protection of the assemblies and components on the drill, is emphasized in chapter 10 which deals with operation of the drill at high altitudes and in severe weather conditions. The interior of a machinery house looks like the cut view shown in Figure 7.26. The roof the machinery house is either easily removable or can be tilted to near vertical position to allow lifting of the heavy assemblies or components by a crane. The walls, doors and even rooftop of the machinery house can be made up of a double wall where two sheets – an inner and outer – are provided. The air between the two sheets acts as a thermal insulator. In many cases the machinery house is pressurized so the outside hot or cool air does not enter the machinery house. The machinery house can also be equipped with air conditioner or heater or both. 7.5.2.10
Dust control system
Rotary blasthole drills flush the hole by compressed air. Along with large chips, very tiny microscopic particles are also formed during the process of rock crushing under a drill bit. The chip- and dust-laden air emerges out of the blasthole during blasthole flushing. If such dust-laden air is allowed to mix with the atmosphere, the atmosphere gets polluted. Inhaling such polluted air gives discomfort to everyone. The harmfulness
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1 - Machinery House 2 - Rotary Motor 3 - Air Compressor 4 - Full Height Door 5 - Electric Console 6 - Machinery House Filter 7 - Mast 8 - Rotary Head 9 - Operator's Cab
Figure 7.26 Cut view of machinery house and operator’s cab.
of polluted atmospheric air depends upon the composition and concentration of particles mixed with the air. Medical research has proved that crystalline varieties of silicon dioxide or silicates in the form of asbestos are dangerous and cause silicosis or lung cancer respectively. Both these diseases are considered to be incurable, but some oxygen therapy practitioners have claimed that oxygen therapies can reverse them. Other ingredients such as fines of feldspar, olivine, amorphous varieties of quartz glass etc. are relatively harmless. To minimize pollution, dust control arrangements are essential on blasthole drills, particularly the rotary type as they create a large quantity of fine particles. Dust control arrangements are of two types viz. wet dust control and dry dust control. 7.5.2.10.1 Wet dust control This is also called water or foam injection. This simple method of dust control is quite effective as can be seen from Figure 7.27. The wet dust collection mechanism consists of a water pump and a water tank mounted on the main frame of the blasthole drill. The water or foam from the tank is sprayed into the compressed air line by the water pump. A mist of water or foam formed in the compressed air line travels to the bottom of the hole and escapes out into the annular space through bit nozzles and rises up. The fine cuttings in the compressed
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Without Water Injection
With Water Injection
Figure 7.27 Effectiveness of wet dust control.
air get coagulated with the moisture or foam. When such aggregates of fine particles come to the top of the hole, being heavier they fall to the ground within the dust hood enclosure. Thus, clean air escapes to the atmosphere. Mixing foam with the water proves very effective because the higher surface tension of the foam mixture coagulates the particles with higher binding forces. Bailing velocity decreases by use of foam but can achieve effective blasthole flushing. The quantity of water to be injected by a water injection system should be such that it forms a mist and remains in gaseous form in the compressed air. The quantity of water to be injected can be determined on the basis of Table 7.5. If foam is to be used the quantity of water must be reduced depending upon the properties of the foaming agent. In wet dust collection the dust hood must have adequate openings within it so the clean air escapes to atmosphere without building up pressure within the dust hood itself. The advantage of water or foam injection is that it is inexpensive to construct, maintain and operate. The disadvantages of water injection system are: 1
2
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Since the compressed air is mixed and the mist travels to the bearings of the drill bit, the life of bit bearings and, therefore the bit, decreases considerably – in many cases as much as 15% in terms of meterage. In conditions where atmospheric temperatures are sub-zero, water in the water tank and particularly the water injection pump and tubes can easily freeze and crack the system. Use of antifreeze liquids can be harmful to the bit and stabilizer bearings and other components of the drill string.
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Table 7.5 Quantity of water to be injected for dust suppurations. Atmospheric temperature in °C
Quantity of water to be injected in cc per m3 of compressed air
45 40 35 30 25 20 15 10 5
40–70 30–60 25–40 18–30 14–22 10–17 8–13 6–9 4–7
Figure 7.28 A dust hood at mast bottom.
Some countries do not accept water injection dust control since the dust deposited on the surface of the ground becomes dry in very short time and gets easily mixed up with the atmospheric air thereafter. 7.5.2.10.2
Dry dust control
In dry dust control systems, when compressed air laden with cuttings, dust and dry clay moves up into the dust hood, most of the sand and larger size particles settle down within the dust hood itself due to decrease in air velocity. But still a large quantity of dust and clay-laden air remains in the dust hood. Dust hoods provided on dry dust collection systems have to be airtight. Further they must keep in contact with the ground even when the mast is tilted for angle hole drilling. One such dust hood is shown in Figure 7.28. The alignment, size and shapes of dust hoods are often controlled by mechanical or hydraulic means. The dust and clay-laden air in the dust hood is further cleaned by a dust collector. Dry dust collector units provided on rotary blasthole drills are usually the filter type. These units suck the dust and clay-laden air from the dust hood and bring it into the filtration unit. Dust collector operation can be explained by means of Figure 7.29.
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Blower Motor
Vibrators Reverse Blower
Blower Fan Reverse Blower
Dust Hood Middle Chamber Filter Elements Hopper
Dust Collection Bag
Figure 7.29 Dust collector schematics.
Drill Pipe
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The blower fan at the top of the dust collector creates a vacuum in upper chamber. This results in a flow of dust-laden air through the conduits to the middle and upper chamber of the dust collector. As the dust-laden air passes into the upper chamber from the middle, it gets filtered by the filter elements in the middle chamber. Thus, the air that passes to the upper chamber and released to the atmosphere is very clean. Dust collected in the middle chamber side of the filter element is made to fall down in the dust bag by a vibrator that vibrates the filter elements at predefined time intervals. A reverse blower also blows compressed air on the filter elements in the reverse direction at some predefined time intervals. This cleans the filter elements by forcing the dust to fall into the dust bag. When the dust bag becomes full of dust, it is replaced with another empty bag. The best way to dispose of the dust in the dust bag is to mix it with some cement and make a concrete block. Some sophisticated dust collectors have electronic circuits that automatically sense the air pressure differential between the upper chamber and middle chamber and operate the vibrator or reverse blower if the differential exceeds a predefined value. It may be a good idea to have another filtering unit to continue the filtering operation even when the dust in the first unit is being cleaned. The filter element fabric is such that it arrests almost all the particles of size larger than 5 microns. Particles of size smaller than this are considered harmless and, therefore, can be ejected with air blowing through the blower fan into atmosphere. The air blowing capacity of the blower fan has to be far in excess of the capacity of the air compressor on the blasthole drill. Only this can ensure that no pressure is built up in the dust hood and that the blasthole flushing remains very effective. Dry dust collectors are more expensive but are more effective in pollution control as compared to wet dust control. They require more maintenance in changing the filter fabric elements. 7.5.2.11
Operator’s cab
In a blasthole drill the operator’s cab is located on the rear end, from where the mast, drill pipe, rotary deck bushing, drill stem wrench, pipe changer and many other items that take part in drilling operations are easily visible. For this purpose one side of the cab is equipped with large glass windows right from the floor to the roof. For an all-round view the other walls of the cab have windows in the upper half of the wall. In most cases the roof also has a glass window so the operator can see the rotary head and upper portions of the mast. The glass used is shatter resistant type, tinted glass. Roof window glass is extra thick and is provided with a steel guard. A typical view of the mast through the roof window is shown in Figure 7.30. The cab is constructed from thick steel sheets and is of FOPS (Falling Object Preventive Structure) type. For noise and vibration reduction, the cab is insulated from the main frame by synthetic rubber mounts. Facilities provided in an operator’s cab include operator’s adjustable seat, helper’s seat, windshield wipers and washers, cab heater and air conditioner, cab pressurization unit etc. Nowadays, in most of the drills electro-hydraulic valves are used. Thus, the controls are electrically operated. Apart from on-off and rotary switches, joystick controls
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Figure 7.30 Mast view through roof window.
Figure 7.31 Operator’s console of a blasthole drill.
are also used. In computer-controlled drills a large color LCD display in front of the operator shows all the important drilling parameters as shown in Figure 7.31. The console in front of the operator contains many indicators and controls. Indicators of two types viz. gauges and lights are provided in the operator’s cab of the drill. A rotary blasthole drill usually has some or all of the indicators mentioned in Table 7.6. Some drills also have some additional indicators depending upon the assemblies incorporated in it. Controls are also of two types viz. levers or switch buttons. Controls normally provided in a rotary blasthole drill are mentioned in Table 7.7. Some drills also have some additional controls.
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Table 7.6 Indicators usually provided on rotary blasthole drill. Indicator list
Indicator list
Drill Level Indicator Rotary Speed Gauge Water Injection Tank Level Gauge Machinery House Heater or Air conditioner On Light Engine Water Temperature Gauge Voltmeter or Ammeter Engine Hour Meter Fuel Level Gauge
Compressor Temperature Gauge Pulldown Force Gauge Rotary Torque Gauge Water Injection System On Light Cab Heater or Air conditioner On Light Engine Oil Pressure Gauge Engine Speed Gauge
Table 7.7 Controls usually provided on rotary blasthole drill. Control list
Control list
Engine Start Button Ether Injection Control Button Leveling Jack Levers Pulldown Force Control Lever Breakout Wrench Control Levers Pipe Rack Control Lever Mast Locking Lever Water Injection Start Button Dust Curtain Lifting Control Levers
Tool Handling Winch Lever Engine Stop Button Compressed Air Shut Off Lever Mast Hoisting Lever Propel Control Levers Rotary Speed Control Lever Pipe Rack Indexing Control Pipe Rack Locking Lever
7.5.2.12
Driver’s cab
When the blasthole drill is carrier-mounted, it is meant to travel on highways. In such case, the drill itself and the driver’s cab must have all the facilities to conform to road regulations. Every manufacturer takes care of providing all the required facilities to comply with such regulations. Commonly essential facilities are: 1
2
3
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Travel lights consisting of side marker lights, red light on the front top and rear top of the horizontal mast, revolving dome light, two position headlights, direction signal lights, parking lights etc. Facilities required in driver’s are driver’s and rider’s seats, seat belts, rear view mirrors, light with switch, steering wheel, accelerator, gear shift lever, clutch, air brakes, parking brakes etc. The front panel requires air pressure gage, voltmeter, fuel level gage, engine oil pressure gage, engine coolant temperature gage, low air pressure indicator, speedometer, tachometer with totalizer, audiovisual warning, high beam indicator light, parking brakes on indicator light, key type starting, engine start button, engine stop button, hand throttle control, transfer case lock-up control with indicator light etc.
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Optional items available in the driver’s cab may include air conditioner, heater, AM/FM radio etc. 7.5.2.13
Fuel tanks
Every diesel-operated rotary blasthole drill is equipped with a fuel tank, usually located below the upper floor on the main deck. In most of the cases the capacity of the tank is such that fuel in the completely filled tank is sufficient for continuous engine operation of 24 hr. An option to enable filling the fuel in the tank from a fuel tanker very rapidly is usually available. A few drills have the option of having an additional fuel tank so as to reduce the frequency of fuel fill. 7.5.2.14
Transformer and cable reel
All electric drills have electric motors working on different voltages and different types of current i.e. AC and DC. The main electric supply available in a mine is AC type and has voltage range between 4160 to 6600 V. The DC motors used in the drill typically work on 475 V, whereas the other AC motors work on 440 to 460 V. Low voltage AC power used for air conditioning and lighting is typically 120 or 240 V. Therefore all electric drills are equipped with one or two transformers. Oil-cooled transformers of the past have now been replaced by dry transformers. A cable reel, mounted on the front end of a blasthole drill, as shown in Figure 7.32, is usually an optional item, but is very essential in electric drills. A cable reel through an electric or hydraulic motor automatically ensures tidily and tightly wound position of the power cable that supplies power from the mine power supply to the blasthole drill even as the drill moves from one blasthole to another.
7.5.3 Assemblies in the mast group Assemblies in the mast group are usually mounted on the mast itself and include: 1 Mast 2 Mast Raising Cylinders and Mast Braces
Figure 7.32 Cable reel.
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Mast Ladder Pipe Changer Angle Hole Attachments Rotary Head Feed Mechanism Auxiliary Winch Tool Handling Jib Centralizer Details of these are given hereunder individually.
7.5.3.1
Mast
The mast of a rotary blasthole drill is an item of predominant importance. It is subjected to many reactions in drilling which result in complex stresses in the mast members. When modern concepts of design of rotary blasthole drills were first introduced, the most radical change was the mast and moving rotary head accommodated in the mast. The mast in itself is a column structure fabricated from square or rectangular steel tubing. Unlike the circular tubing that requires contour cutting, square or rectangular tubing needs plane cuts and the welds can be easier as well as more reliable. The cross section of the mast is rectangular in most cases but some drills have trapezoidal or triangular masts. Three or four main members of the mast are joined together by means of lattice members to form a open-faced hollow structure as shown in Figure 7.33. The rotary head is visible and accessible from the open face and slides in an up and down direction within the members. Main Members Rotary Head Containing Gear Reduction Mechanism
Lattice Members
Hydraulic or Electric Motor Position of single pipe racks
Rotary Swivel and Spindle
Hydraulic or Electric Motor
Main Members
Figure 7.33 Cross section through a mast.
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191
Mast raising cylinders and mast braces
The mast is raised and lowered by means of two hydraulic cylinders. For long distance travel such as from one bench to the other, the mast of the blasthole drill is kept in the lowered position so that the center of gravity of the drill attains its lowermost position and travel is safer. While moving from one blasthole to the other, the mast is kept in the raised position to save time. Some blasthole drills have a very tall mast in order to accommodate long drill pipes to facilitate drilling deep holes in a single pass. In such structures the stresses and vibrations induced in the mast by the drill head are amplified. Thus, a tall mast needs additional support that is provided by two mast braces. Mast braces are often hydraulically extended and mechanically locked so the longer fixed length gives necessary rigidity to the mast. This reduces the unsupported mast length and results in low vibrations which otherwise would have been higher by relying only on the hydraulic mast raising cylinders. For alignment of the mast in prefixed angles, mast braces have additional fixing points. Mast-raising cylinders and mast braces can be easily seen in Figure 4.12. 7.5.3.3
Mast ladder
Occasionally it becomes necessary to climb up to the top of the mast. For this purpose a mast is equipped with a ladder. For safety purposes the ladder is surrounded by circular rings or a steel grill. With this grill climbing becomes very easy and safe. 7.5.3.4
Pipe changer
A pipe changer is also known by many names viz. pipe rack, rod changer or carousel. In blasthole drilling, the addition of a drill pipe to the drill string and its removal after drilling a blasthole to the desired depth is avoided by selecting a drill that can drill to full depth with one drill pipe. However, for drilling deep holes, extra drill pipes are to be added. For this reason, almost every blasthole drill has an arrangement for storing additional drill pipes in the mast and mechanically joining them to the drill string. This mechanism is called a pipe changer. Pipe changers differ from drill to drill but mostly they are either carousel type and single pipe type. 7.5.3.4.1
Carousel pipe changers
Small, medium, and some of the large, blasthole drills are equipped with a carousel pipe changer. It is often fixed to the main member of the mast on the opposite side of the operator’s cab or within the mast behind the drill head like the one shown in Figure 7.34. The number of drill pipes stored in a carousel pipe changer depends upon the pipe diameters but in most cases it is between 3 to 6. In a carousel pipe changer the drill pipes are stored in a rack that consists of a central spindle with one bottom and two top plates. As shown in Figure 7.35 the bottom plate has sockets on the upper side to accommodate the lower ends of the drill pipe. The upper ends of the drill pipe are contained within a slot in the lower of the two top
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Figure 7.34 Pipe changer within a mast.
Upper of the Top Plates
Drill Pipe Free to Move Out from the Carousel
Lower of the Top Plates Drill Pipe Locked in the Pipe Rack Slot
Socket Bottom Plate
Figure 7.35 Pipe locking in the pipe changer.
plates. When the pipe is in lowered position it is restricted in its upper movement by the upper of the top plate. It cannot come out of the slot because the slot opening is of smaller size than the drill pipe diameter. Thus, the drill pipe in its lower position is locked. The upper of the top plate has an opening from the top. The drill pipe can be
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lifted through this opening by the drill head. In the lifted position the drill pipe comes out of the socket on the bottom plate and the small diameter groove on the drill pipe aligns to the level of the slot opening as shown in Figure 7.35. Thus, in this position the drill pipe is unlocked and is free to come out of the carousel. The steps taken in adding a drill pipe to the drill string are as follows: 1 When drilling with one drill pipe is finished and a drill pipe needs to be added, a spanner at the bottom of the mast is forced out and holds the upper end of the drill pipe through its flat. This is shown in Figure 7.36. 2 The rotary head rotates in the reverse direction so the drill pipe is uncoupled from it. In most drills such uncoupling is started by a hydraulically operated tool wrench that grips the saver sub at the lower end of the drill head and slightly rotates it. After this loosening the drill head uncouples the saver sub. 3 The rotary head moves to the top position in the mast along with the saver sub. 4 The complete carousel assembly is moved from its normal position to the center of the mast by means of a hydraulic cylinder in such a way that the opening of the upper of the top plates is just below the saver sub. 5 An indexing mechanism, also operated hydraulically, rotates the lower of the top plates and the bottom plate in such a way that one drill pipe is in alignment with the saver sub and the drill pipe in the blasthole is held at the mast bottom by the spanner.
Drill Head
Spanner Saver Sub
Drill Pipe
Figure 7.36 Bottom spanner.
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6 The rotary head moves down and rotates the saver sub in the right direction so that the drill pipe in the carousel is coupled to the saver sub. 7 The rotary head is then moved up slightly so the drill pipe in the carousel moves up and attains a position in which it can come out of the socket at the lower end and the narrow groove of the drill pipe comes in the level of the slot opening. 8 When the pipe is in such position the carousel is moved back to the original position on the left of the mast. Thus, the newly added drill pipe remains in the drill string alignment and the pipe changer is away from it. In case of inclined blastholes it may be necessary to hold the newly added drill pipe in proper alignment by yet another arm. 9. The rotary head with newly coupled drill pipe is then moved down. When the drill head rotates in the right direction the newly added drill pipe gets coupled with the drill pipe in the blasthole held in position by the spanner. 10. The spanner is then retracted back. This once again allows free rotation of the drill string and drilling is continued further. The above steps are repeated for adding one more drill pipe whenever required. This mechanism appears quite simple but to have it working precisely, it has to be designed very carefully because even small deflections can cause problems. Such problems increase considerably in coupling drill pipes in an inclined direction while drilling angle holes. The advantage of the carousel type pipe changer is that it can allow the use of many drill pipes. 7.5.3.4.2
Single pipe changers
Single pipe changers are used in some large and most of the extra large blasthole drills. This is because single pipe changers are sturdier without being very heavy, and are better suited to handle very long drill pipes of large diameter. In modern drills there can be as many as four single pipe changers as shown in Figure 7.37. Single pipe changers are totally independent of each other and are operated separately. After completing drilling with the first drill pipe, it is held in position by a spanner as explained earlier for a carousel changer. The rotary head detaches from the pipe and moves up in the same manner as explained in step 1 and 2 for a carousel pipe changer. The drill pipe in a single pipe changer is held in locked position by means of a socket shown in the in Figure 7.37 at the bottom and a gate at the top. Besides this, other means of locking the drill pipe are also used for the purpose of safety. By extending the hydraulic cylinder at the bottom of the pipe rack, as shown Figure 7.38, the drill pipe gets positioned in line with the drill pipe held by the spanner at the bottom of the mast. The gate is then opened to allow coupling of the drill pipe to the rotary head as described in step 5 and 6. Once the drill pipe is coupled to the rotary head, a centralizer swings into position and holds the drill pipe at some distance from the lower end of the mast. The centralizer is particularly essential while coupling drill pipes in angle hole drilling. At this juncture, since in this position the drill pipe is safely and securely held, other locking arrangements are released by retracting the hydraulic cylinder. Thus the
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Hydraulic Wrench
195
Four Single Pipe Changers
Figure 7.37 Four single pipe changers and hydraulic wrench.
Upper Arm
Gate
Latch
Drill Pipe
Main Section of the Mast
Hydraulic Cylinder for Lowering the Drill Pipe
Lower Arm
Figure 7.38 Operation of single pipe changer.
single pipe changer moves back to its original position. As the pipe is free to slide and rotate in the centralizer, it is rotated by the drill head and coupled to the drill pipe held by the jaw at the bottom of the mast. Drilling can be continued after withdrawing the jaw and releasing the pipe from the centralizer.
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7.5.3.5
Bit changer
Pipe changers have been used on rotary blasthole drills for the past sixty years. In fact it was one of the first automation features of the blasthole drills. Recently Bucyrus International have come up with a new bit carousel and forked pipe wrench, as shown in Figure 7.39 on their 49 series blasthole drill. The bit changer accommodates up to four tricone bits. In the bit changing operation the forked pipe wrench engages the drill pipe flats just above the bit. Then the wrench rotates while simultaneously transmitting impacts on the pipe. In the process the tool joint loosens. The drill bit is then maneuvered into the bit carousel and by reversing the sequence a new bit is attached to the drill pipe. All these operations are automatic and are carried out by the driller while sitting in the cab. Even the assistant need not go to the drilling platform. All this means significant safety measures are introduced into the bit changing operations. More importantly the time required for changing the drill bit reduces to about 10 minutes from about 60 minutes that are required for the same operation without the tricone bit carousel. 7.5.3.6
Angle hole attachments
For reasons elaborated elsewhere in this book, blastholes are often drilled at an angle with the vertical. In most instances this angle is within limits of 0 to 30° off vertical. When the mast of a blasthole drill is raised from the horizontal to the vertical position by using mast raising cylinders it does attain such angles, but mast raising cylinders are meant for raising the mast and not for firmly holding it at an angle with the vertical. These cylinders are rather short and keep the very long length of the mast unsupported. This causes very heavy vibrations in the mast and if resonance occurs it can fail. Thus, to align the mast at the desired angle and keep the unsupported length to a bare minimum, some additional arrangements are essential. These are termed angle hole attachments.
Figure 7.39 Bit carousel (upper) and forked wrench (lower).
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Most of the modern blasthole drills have angle hole attachments that enable mast inclinations at angles up to 30° at increments of 5°. Some have a maximum inclination angle of only 20°. In some drills the angle can be infinitely variable within their range from vertical to maximum inclination. The only rotary blasthole drill that can give reverse inclination is the Bucyrus 39R. It can give reverse inclination of 15° in addition to the normal inclination of 30° in an infinitely variable manner, as illustrated in Figure 7.40. There are three types of angle hole attachments. The first type contains a yoke provided on the main frame at the bottom of the mast as shown in Figure 7.41. The mast is raised by means of mast raising cylinders up to the desired angle and then it is firmly fixed on the yoke by means of pins or nut bolts. This alternative is used in blasthole drills where the unsupported distance of the mast above the mast raising cylinder support is not great, or in other words the mast is short.
Figure 7.40 Negative inclination angle of bucyrus 39R.
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Figure 7.41 Bottom yoke arrangement for angle locking.
The second type, usually adapted for large blasthole drills that have tall and heavy masts, consists of telescoping braces with locking fixtures at desired distances as can be seen in Figure 7.42. The mast is raised in the desired angle by using the mast raising cylinders and then the telescoping braces are locked at the holes provided within them. The third type is very similar to the second type. It consists of telescoping braces that extend by means of hydraulic cylinders. The main hydraulic mast raising cylinders tilt the mast up to the vertical and then the hydraulic cylinders provided on the braces lower the mast to the desired inclination. This type gives infinitely variable mast angles within the two limits. Attractive and advantageous as it may appear, this arrangement does not offer a great advantage over the telescoping braces because once the blasthole drill mast is set for a particular inclination, it is not required to be changed frequently since all the blastholes are to be drilled at the same inclination. When a blasthole drill frame is positioned horizontally the angle of inclination is automatically achieved. 7.5.3.7
Rotary head
The rotary head of a blasthole drill rotates the drill string and exerts feed force on the drill bit through the drill string. In rotary drills the rotary head is fabricated in the form of a box from steel plates. Power to the rotary head is given by hydraulic or electric motors. Trains of gears drive the main spindle at highly reduced speed and increase torque. The drill string is attached at the lower end of the spindle. The upper end of the spindle is attached to the air hose through a built-in swivel. An inner view of a rotary head in Figure 7.43 clearly shows all the gears used for transmitting power. In large and extra large blasthole drills the rotary speeds of the spindle on rotary blasthole drills are infinitely variable from 0 to about 125 rpm. In small blasthole drills the rotary speed range is often 0 to about 200 rpm.
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Figure 7.42 Angle hole drilling attachment through mast braces.
Figure 7.43 Inner view of a rotary head.
Some drills have dual speed hydraulic motors so the slow speed high torque or high speed low torque options can be chosen as required. Hydraulic motors enable finer control of the drill string rotation. In extra large electric drills the motors are DC type. Electric motors have high efficiency, low operating cost and low maintenance requirements compared to the hydraulic motors. A shock absorber, fixed at the lower end of the spindle, has now become almost a standard item. Shock absorbers have been described in detail in the chapter on drilling accessories.
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7.5.3.8
Feed mechanism
The feed mechanism is meant for moving the rotary head in an up and down direction and to exert necessary feed force on the drill head. As the drill head moves down, the drill string exerts force on the bit. There are different types of feed mechanisms: 1 2 3 4 5 6
Double Drum and Wire Rope Hydraulic Cylinder and Wire Rope Hydraulic Cylinder and Chain Fixed Sprockets and Chain Rack and Pinion with Chain Chainless Rack and Pinion Details of these mechanisms are given below.
7.5.3.8.1
Double drum and wire rope
This mechanism was used on blasthole drills made in the early 1950s but has now been discarded, as the other mechanisms are far superior in functi0nal efficiency. The following description is for the record. The mechanism consisted of one drum, called a shipper shaft drum, on the drill head. The other drum was that of a powered bull reel placed on the main frame. Two pieces of rope were wound on both the drums in opposite directions. When the bull reel rotated, one of the ropes was wound on it and the other unwound from it. These ropes gave up and down movement to the shipper shaft drum on the drill head and consequently the drill head itself. Possible free fall of the drill head was prevented by the brake on the shipper shaft drum. This mechanism was used for long feed travels. 7.5.3.8.2
Hydraulic cylinder and wire rope
Schematic illustration of this mechanism is in Figure 7.44. In this mechanism, when the yoke mounted on the rod of the hydraulic cylinder moves through a distance X the rotary head moves through a distance 2X in the opposite direction. In this process, the speed at which the ram moves is doubled at the drill head, but the force produced by the hydraulic cylinder is reduced to half at the drill head. As constant-pressure variable displacement hydraulic pumps are used to supply hydraulic oil to the cylinder, the feed rate gets automatically adjusted to the penetration rate. The merit of this mechanism is that it allows precise control over the feed rate by finely adjusting the hydraulic oil outflow from the variable displacement pump. The drawbacks of this mechanism are: 1 2
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Large sheaves are needed, to match the desired ratio of sheave to wire rope diameter. Thus, the cross section of the mast becomes large. Since it is not easily possible to manufacture hydraulic cylinders of length in excess of about 6.5 m, the maximum travel of the rotary head is limited to about 13 m.
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Upper Sheave or Sprocket
201
Upper Rope or chain Fixing Block Wire Rope or chain
Front Member of the Mast Through Which the Drill Head Moves Up or Down
Yoke
Rod of Feed Cylinder Drill Head Mast Shock Absorber
Drill Pipe Lower Sheave or Sprocket
Hydraulic Feed Cylinder Lower Rope or chain Fixing Block
Figure 7.44 Hyd. cylinder and wire rope feed mechanism.
3 4
The wire rope is required to be tightened and lubricated very often. The yoke needs a guide for its movement.
Occasionally, in this mechanism the hydraulic cylinder is fixed to the yoke rather than the mast, and the piston rod of the cylinder is fixed in the mast i.e. exactly opposite placement of the cylinder to the one shown in Figure 7.44. This gives higher retract force and lesser pulldown. 7.5.3.8.3
Hydraulic cylinder and chain
This mechanism is identical to the hydraulic cylinder and wire rope mechanism described above and shown in Figure 7.44, except that the wire rope is replaced by a chain. It has all the limitations cited earlier but the use of a chain allows use of small diameter idler sprockets in place of large diameter sheaves. This, in turn, allows the mast cross section to be reduced. The maintenance required in tightening and lubrication of chain is also less as compared to that of wire rope. This mechanism is adopted on many of the current small and medium blasthole drills. 7.5.3.8.4
Fixed sprockets and chain
The fixed sprocket and chain mechanism comprises two pairs of components. One pair is positioned on the left of the mast and other on the right of the mast. Each pair
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consists of two upper and two lower sprockets, two idler sprockets, one drive sprockets and a chain as shown in Figure 7.45. Two ends of each of the chains are attached to the drill head – one from the bottom and one from the top. When the shaft common to both the drive sprockets is rotated by an electric or hydraulic motor, through a reduction gear box if necessary, the sprockets pull the chain up or down and the rotary head also moves up or down. In most of the cases a single strand chain is used for sturdiness and ease in maintenance, but in a very few cases a multi-strand chain may have to be used. The assembly comprising the motor, reduction gear box, drive shaft and driving sprockets is kept near the lower end of the mast for ease of maintenance and lower center of gravity. This mechanism can be used for very tall masts meant for drilling with long drill pipes. 7.5.3.8.5
Rack and pinion with chain
A schematic illustration of rack and pinion with chain mechanism is shown in Figure 7.46. In this mechanism the drill head slides through the two front main members of the mast. Each of these two main members has a gear rack from top to bottom. The pinions on the drill head intermesh with this rack. The pinions are driven through a gear box by chains and sprockets located on the drill head. The chains are moved up
Upper Rear Sprocket
Upper Front Sprocket
Mast Front Member of the Mast Through Which the Drill Head Moves Up or Down
Drill Head
Chain
Idler Sprocket Drive Sprocket
Shock Absorber
Idler Sprocket
Drill Pipe Lower Rear Sprocket
Lower Front Sprocket
Figure 7.45 Fixed sprocket and chain feed mechanism.
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Upper Rear Sprocket
203
Upper Front Sprocket
Mast Front Member of the Mast Through Which the Drill Head Moves Up or Down
Rack Teeth Driven Sprocket
Idler Sprocket Pinion
Idler Sprocket
Drill Head
Drive Sprocket
Idler Sprocket
Idler Sprocket
Shock Absorber Drill Pipe Lower Rear Sprocket
Lower Front Sprocket
Figure 7.46 Rack and pinion with chain.
and down by drive sprockets positioned on the mast which are driven by an electric or hydraulic motor. As the pinions rotate the drill head moves up or down. The racks are made from forgings that are cut and deep hardened. The pinions with a little less hardness may have to be changed, but the rack gears are extremely durable and last almost the lifetime of the drill. This mechanism can be used for very tall masts meant for drilling with long drill pipes. Nowadays this mechanism is rarely used as it requires many more components, particularly the chain, that require considerable maintenance. 7.5.3.8.6
Chainless rack and pinion
Chainless rack and pinion mechanism is shown in Figure 7.47. In this mechanism two gear racks are welded onto two main members of the mast. Each rack extends from the bottom to the top of the mast. The rotary head slides up and down between these two main members. Four pinions are built into the drill head as shown. These pinions are intermeshed with the rack gears. The pinions are driven by a hydraulic/electric motor through a reduction gear box. When the pinions rotate the drill head moves up or down. The rack gears are made from special forgings that are deep hardened after gear cutting.
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Upper Left Pinion
Upper Right Pinion Drill Head
Lower Left Pinion
Lower Right Pinion
Left Rack
Right Rack
Shock Absorber
Figure 7.47 Chainless rack and pinion feed mechanism.
The chainless rack and pinion feed mechanism is reckoned to be the best and very long lasting. The pinions with a little less hardness may have to be changed after a long time, but the rack gears are extremely durable and last almost the lifetime of the drill. This mechanism can be used for drills with a tall mast that accommodates long drill pipes. It is found mostly on the extra large blasthole drills. Maintenance efforts required for this mechanism are considerably low as compared to the other mechanisms. 7.5.3.9
Auxiliary winch
Almost all the accessories used in rotary blasthole drilling are so heavy that they cannot be manually lifted, shifted and handled. Therefore, almost every rotary blasthole drill is provided with a wire rope and an auxiliary winch. The winch is powered by hydraulic or electric motors through a planetary reduction gear box for compactness. The winch can be placed on the main frame or near the lower end of the mast as shown in Figure 7.48. Depending upon the size of the drill, and thus the weights of the accessories used with it, the single line pull of the auxiliary winches range from 15 to 50 kN. The diameter and width of the winch drum depends upon the wire rope diameter and length, which in turn are dependent upon the maximum single line pull of the winch and the height of the mast. The wire rope is taken up to the top of the mast through a few idler sheaves. It comes down from a set of sheaves placed at the top of the mast. 7.5.3.10
Tool handling jib
To enable movement of front sheave at the top of the mast to right or left, a tool handling jib, similar to the one shown in Figure 7.49, is sometimes provided at the top of
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Figure 7.48 Auxiliary winch.
the mast. The movement of the jib is controlled by means of two hydraulic cylinders, one of which enables the jib to telescope in and out and the other facilitates lateral shifting of the jib. 7.5.3.11
Centralizer
This item, shown in Figure 7.50, consists of an arm attached to the mast at its middle height. A long drill pipe, when it is being attached to a drill string, tends to bend in the middle because of its slenderness. This is particularly so when the mast is at an angle for drilling inclined blastholes. The curvature of the new drill pipe makes it difficult to attach it to the drill string and the drill head. Cross-thread connection and thread damage can occur easily. The centralizer keeps the alignment of the new drill pipe very close to the alignment of the drill string and reduces the possibility of cross-connection and thread damage.
7.5.4
Special purpose items
Apart from the main items described earlier in this chapter, many more items are built into a blasthole drill for fulfilling some particular needs.
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Figure 7.49 Tool handling jib.
In the absence of standardization in nomenclature, same items are also known by different names. The following are details of some items available, either as standard or optional, on blasthole drills.The order in which they appear is random. 7.5.4.1
Language name plate
Many instructions/warnings are printed or stuck through printed stickers at various places on a blasthole drill to fulfill safety regulations. These instructions are normally in the English language and in Roman Script. Blasthole drills are used in many countries where the English language is not well understood. If this language name plate option is chosen the instructions and warnings are printed in the language and script chosen by the purchaser. 7.5.4.2
Lighting
Almost all the blasthole drills are operated round the clock. On certain locations the light available during daytime hours is also insufficient for properly viewing drilling
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Figure 7.50 Centralizer supporting long drill pipe.
operations and various portions of a blasthole drill. Lighting arrangements are very essential in such circumstances. Most manufacturers provide a lighting system as a standard feature of their blasthole drill. Depending upon the size of the drill, mercury vapor, quartz-iodine or other types of lamps are provided at different parts of the blasthole drill. The number of lamps ranges from 7 to 10 and wattage of each lamp is about 75 to 100 W. Figure 7.51 shows night operation of a rotary blasthole drill from which one can get an idea about the placement of lamps. 7.5.4.3
Fast fuel fill system
This system is often known by other proprietary names. On a large blasthole drill diesel is consumed very rapidly. Even a large capacity fuel tank has to be refilled frequently without stopping the engine and drilling operation. If the tanks is filled by the usual method of rapidly pumping the fluid from a pump on
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Figure 7.51 Night operation of a blasthole drill.
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the tanker through a hose pipe, there is a danger of bubble formation in the tank. It can cause severe troubles in engine running. One component of a fast fuel fill system is mounted on the fuel tank itself. The other part is the fuel filling bayonet. The fuel starts flowing to the fuel tank automatically only after these two components are coupled together properly. The component on the fuel tank has an automatic level sensing device. When the tank is filled to the appropriate level the fuel flow to the tank automatically stops and the indicator light on the fuel filling bayonet gets illuminated. When the bayonet is disconnected from the fuel tank inlet there is no drop spillage of the fuel. Fast fuel fill systems are designed to ensure that fuel will be filled in the tank at rapid flow rate but without bubble formation and to a predetermined level. A schematic of a typical fast fuel fill system is shown in Figure 7.52. 7.5.4.4
Depth indicator
Many of the small and medium blasthole drills are not equipped with a computerized drilling system. In such event this item, which numerically shows the blasthole depth, becomes essential. These numerical readings are far more accurate than those judged by a driller through his visual observations or perhaps by some marking made on the mast. Instead of the old mechanical depth indicators that worked through a cable and took measurements from the position of a drill head within the mast, many drills are now provided with electronically operated depth indicators that indicate blasthole depth inside cab after sensing the drill pipe addition. More details about such depth indicators have been given in chapter 6. 7.5.4.5
Fast retract system
In rotary blasthole drilling very high feed force is required to be exerted on the drill string by the drill head. While doing so, the speed with which the drill head can travel downwards is often very low because progress of blasthole drilling itself is very slow.
Fuel Tank
Fuel Vacuum
Fuel Level Sensor Fuel Inlet
Fuel Gun or Bayonet
Fuel
Figure 7.52 Fast fuel fill system.
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For ease of controlling the feed rate the maximum feed rate at which the drill head can travel has to be kept limited. Since maximum feed rate and maximum retract rate are related to each other, due to the characteristics of actuators like electric or hydraulic motors or hydraulic cylinders, the retract speeds are also limited. It is desirable to withdraw the drill string rapidly so as to increase overall drilling speed. As blasthole depth is very shallow, the force required to withdraw a drill string from a blasthole is also very low and it is possible to retract the drill string at fast speed by using some specialized mechanism or devices. Many rotary blasthole drills provide such a device or mechanism as an optional item that can be chosen by the customer. It is called a fast retract system. A fast retract system supplies a higher volume of hydraulic oil to the hydraulic cylinder or hydraulic motor at lower pressure, so the speed of retraction is increased considerably. 7.5.4.6
Tow hook
Blasthole drills are self-propelled and have sufficient traction power to move in rough terrain conditions. However, occasionally a blasthole drill has to be pulled or towed by another vehicle like a dozer or a tractor. To enable such pulling or towing operation one or more tow hooks must be provided on the blasthole drill at a specific position so the operators do not use incorrect positions for slinging of the wire rope. Most rotary blasthole drills have this item as an optional extra. Apart from the pull load sustaining ability, a tow hook must have self locking arrangements for safety. 7.5.4.7
Fire extinguishers
Fire gives rise to smoke, flames and heat that cause damage to equipment. Fire in a blasthole drill can be disastrous not only because a blasthole drill is an expensive piece of equipment but more so because its repair can extend over a long period. Replacement of a rotary blasthole drill by a rented piece is very difficult. In either case it results in very high production losses. An automatic fire detection and extinguishing system described in the next section is very effective, but many small blasthole drills are equipped with lightweight fire extinguishers placed at different locations in a blasthole drill that can be operated manually like the one shown in Figure 7.53. In case of fire, the nozzle of the fire extinguisher is directed towards the root of the flame and a switch on the handle is pressed. The orifice in the nozzle opens and the gas stored in the fire extinguisher under high pressure is very rapidly blown at the fire to create an inert gas surrounding the fire. Fire usually gets extinguished when oxygen concentration around the fire drops below 15%. Gases used in fire extinguishers are required to be the inert type, not only so that they should not burn themselves at high temperature or aid burning of other gases, but they must not decompose to form any component having such undesirable properties. Further, they must be safe for human beings.
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Figure 7.53 Lightweight fire extinguisher.
Inert gases are best suited for fire extinguishing purposes. Commonly used gases in a fire extinguishers are nitrogen and argon. Carbon dioxide also does not decompose very easily and hence finds use in fire extinguishers. To reduce the cost of refilling, some fire extinguishers have a mixture of the above gases. In such a mixture, a high argon proportion is very essential because it is heavier and settles down on the floor where the root of fire normally exists. Some fire extinguishers have gases in liquid form, but in most handheld fire extinguishers, chemicals that generate a high volume of specific gases are used in place of gases in liquid form. To increase their effectiveness in extinguishing Fire, some fire extinguisher also have arrangements of spraying non inflammable foaming agents. These foaming agents form a film around the gas bubbles and do not allow the oxygen in the atmosphere to mix and continue the combustion.
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7.5.4.8
Automatic fire suppression system
Sensing of fire is possible through different types of electronic sensors some of which are schematically shown in Figure 7.54. Photoelectric or ionization type smoke detectors detect fire by sensing smoke generated by fire. In photoelectric sensors a ray of light from a light source is normally not sensed by the light sensors but when smoke particles enter the chamber they scatter the light from the light source and it is sensed by the light sensor. In ionization-type sensors the air in the chamber is continuously ionized by a radioactive source. So long as the air does not contain smoke a small current flows continuously from one plate to the other due to electric conductivity of the gases in the air. When smoke enters the space between two plates the conductivity of gases reduces and electric current discontinues. Smoke sensors can give a false alarm if the dust concentration in the air increases above certain level. Heat sensors are better suited in dirty or dust-polluted environments that often surround blasthole drills. The most common heat detectors either react to a broad temperature change or a predetermined fixed temperature. Heat detectors use a set of temperature-sensitive resistors called thermistors, the resistance of which decreases as the temperature rises. One thermistor is sealed and protected from the surrounding temperature while the other is exposed. A sharp increase in temperature reduces the resistance in the exposed thermistor, which allows a large current to activate the detector’s alarm. Flame detectors are devices that look for specific types of light (infrared, visible, ultraviolet) emitted by flames during combustion. When the detector recognizes this light from a fire, it sends a signal to activate an alarm. In automatic fire suppuration system one or more large size fire extinguishers are kept together, as shown in Figure 7.55, and connected to a network of tubes. These
Ionization Type Sensor
Photoelectric Detector
Radioactive Source
Ionized Gas
Unobstructed Current
Light Source Smoke Particles
Light Sensor
Heat Detector
Sealed Thermistor
Exposed Thermistor
Figure 7.54 Fire sensors.
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Sensor Electric Line
Nozzle
213
Fire Sensor
Alarm Gas Line
Control Cylinders Filled with Fire Extinguishing Gas
Figure 7.55 Fire extinguishing set up.
tubes have outlets at many places on the blasthole drill where the possibility of fire exists. One or more types of sensors are also provided at such places. When fire is sensed by sensors at a particular Location, the electronic device gives an audible alarm and automatically operates the fire extinguisher nozzle opening mechanism, and the gas contained in the extinguisher flows through the particular tubing to one or more outlets. When the gas comes out, the oxygen concentration at that point reduces below 15% and the fire is put out. An automatic override to the system is also provided when the heaters start heating the air or the atmosphere becomes very hot. 7.5.4.9
Welding outlet
In blasthole drilling practice, sometimes it becomes necessary to resort to welding for temporarily overcoming difficulties. Welding requires direct current of high amperage. Some drills, particularly the extra large all-electric type, have DC electric power for driving the DC electric motors. An outlet made available for such current is the welding outlet. 7.5.4.10
Automatic leveling system
Nowadays automatic leveling systems are being made available, as an option, on more and more rotary blasthole drills. Precision and speed in leveling a blasthole drill as well as relief to the operator are the main reasons for which automatic leveling systems are preferred by more and more users. Automatic leveling systems have different ways of working. They have sensors to sense hydraulic oil pressure or the actual inclination of the blasthole drill frame and its direction as the leveling operation begins. The logic circuit then determines the quantity
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of hydraulic oil to be pumped into or out of each of the four leveling jacks. The flow to the jacks is controlled by electrohydraulic control valves. This operation is carried out in a cyclic iterative manner so the level of the blasthole drill is corrected in stages. In every stage the blasthole drill attains a more and more precisely leveled position. The operation is stopped when the main frame of the drill is within about 0.1° from a perfectly level position. The operator is given an audio alarm after the leveling operation by the automatic leveling system is completed. He then can lock the hydraulic cylinders and continue with the next step in the operational sequence. Automatic leveling systems have an override so an operator can take charge of the drill leveling operation instead of the leveling system. 7.5.4.11
Remote propel control
Placing a blasthole drill on the exact location of a blasthole is very important from the viewpoint of effectiveness of subsequent blasting. Attaining exact placement of the drill is relatively easy in the case of vertical blastholes, but when it comes to inclined blastholes the task is difficult and a good bit of judgment is required in the operation. When a blasthole drill propels to the location of a new blasthole, the blasthole position mark gets masked by the dusthood and becomes invisible to the operator. Guidance of an assistant standing on the ground is inevitable. Small size blasthole drills are usually propelled in this manner i.e. under the guidance of an assistant. Some more sophisticated large blasthole drills have a viewing window in the dust hood and lighting arrangements so the site of the blasthole marking is visible to the operator sitting in the cab. Thus, he can position the drill quite easily. In any case it is much safer to be away from the drill while it is being propelled to the new location. To enable propelling a blasthole drill by standing on the ground, some blasthole drills have an option of remote propel. Remote controls are either cable-connected type or wireless communication type . A remote control includes a small handheld box, as shown in Figure 7.56. The box contains a device for emitting different types of digital signals based on which lever
Figure 7.56 Box for remote propel control.
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has been operated and how much it has been pressed by the driller. These signals are transmitted to a receiver on the drill through a cable or wireless transmission, depending upon the type of remote control. When the receiver on the blasthole drill receives these signals the hydraulic pumps in the drill supply hydraulic oil to the propel motors in appropriate directions. If both the hydraulic motors rotate in the same direction at same speed the drill moves in a forward (or reverse) direction. If the motors rotate in the same direction but at different speeds the drill moves in a curved path. It is also possible to rotate the drill around a central axis without moving by rotating the motors at the same speed but in reverse directions. With the help of a remote control box, a driller can stand on the ground and move the drill precisely onto the location of the blasthole collaring point. Many remote propel control boxes have automatic leveling control. The signals sent from the box are received by the automatic leveling system and the drill is automatically leveled at the collaring point. The operator can then get into the cab and start drilling operations 7.5.4.12
Air conditioners
The operator’s cab and machinery house are often equipped with air conditioners. The air conditioners are of the same type that are used for air conditioning passenger buses. Electric power to the air conditioners is supplied from a small generator in the case of diesel engine-operated drills or from the transformer in the case of electric drills. 7.5.4.13
Heaters
Heaters become essential when the drill has to operate in cold weather. Apart from the operator’s cab and machinery house, heaters may have to be fitted on the rotary head gear case and the cases of gears that reduce the speed of the propel motors. The heaters can be infrared type or heating coil type. In most cases a fan is provided near the heater so it spreads the hot air within all the space of the enclosures like cab, machinery house etc., or the enclosures provided around the gear cases. 7.5.4.14
Hydraulic test station
A rotary blasthole drill may contain only 2 to 5 hydraulic pumps but it has many hydraulic circuits meant for different operations. The pressure in each of the circuits may have to be known under certain circumstances. For this purpose an hydraulic test station, as shown in Figure 7.57, is provided on some blasthole drills. The hydraulic test station has several end points of hydraulic tubes, each emanating from a certain hydraulic circuit. A special pressure gauge can be inserted in the tube at the point and the pressure in the circuit can be easily read. 7.5.4.15
Video camera system
A driller or his assistant sitting in the operator’s cab of a blasthole drill have very very limited view of the surroundings. This is particularly true in the case of large rotary blasthole drills.
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Figure 7.57 Hydraulic test station.
From the viewpoint of safety of the blasthole drill as well as other equipment and the respective operators or uninvited intruders, it is necessary that the driller’s assistant keeps an eye on the surroundings of the blasthole drill. To enable the driller’s assistant to view the surroundings a video camera system is provided on some blasthole drills. The system comprises video cameras fixed in the blasthole drill at appropriate points. The display emanating from these cameras is shown on a screen provided in the operator’s cab as illustrated in Figure 7.58. The number of cameras is usually up to four but depending upon the necessity it can be reduced or increased.
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Figure 7.58 Display of video camera system.
7.6
EXTREME COLD OPERATION DEVICES
For operations in extremely cold weather conditions and high altitudes many devices, some of which are described here, can be specially provided on a rotary blasthole drill.
7.6.1
Engine starting aid
Diesel engines do not have any sparking mechanism. The heat generated by the compression of the air and diesel mist in the engine cylinders takes the temperature to flash point and ignition is caused to generate mechanical energy in the form of shaft rotation of the engine. When a diesel engine has to operate in extremely cold weather the temperature of the mixture itself is so low that the heat generated by compression is insufficient to take the mixture temperature to flash point. Thus, the engine does not start.
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To overcome such situations, normally an engine is provided with glow plugs within its cylinders or an alcohol vaporizer at the air intake of the engine. 7.6.1.1
Glow plugs
Glow plugs are located in the pre-combustion chamber of each of the engine cylinders. They generate heat from the electric supply and heat the air within the chamber. Thus, the temperature of the air and diesel sprayed into the chamber becomes sufficiently high for its ignition after compression. It is important to use glow plugs of specific rating only, otherwise the results can be disastrous. The glow plug electric supply is started with a switch so the glow plugs can be kept on only at the time of engine starting. 7.6.1.2
Manifold flame heater
On some engines a manifold flame heater is used for starting the engine in extremely cold weather. It consists of a housing, spark plug, flow control nozzle and two solenoid control valves. The flame heater ignition unit energizes the spark plug. The nozzle sprays fuel under pressure into the intake manifold assembly. The fuel vapor is ignited by the spark plug and burns in the intake manifold. The heat from this fire warms the manifold and the combustion chamber. Thus, the temperature of the air entering in the combustion chamber rises and the oil mist in the combustion chamber can ignite easily.
7.6.2
Batteries
In cold weather batteries lose the charge very rapidly. For extreme cold weather operation batteries of high AH rating and specialized construction may have to be used to minimize the troubles in starting a diesel engine.
7.6.3
Double wall machinery house
For protection of the electrical equipment from snow and rainwater, a machinery house is provided on most of the electric drills and all-electric drills.
Figure 7.59 Double wall of machinery house.
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Diesel hydraulic drills are rarely equipped with a machinery house as diesel engines are not so vulnerable to rainwater. Further, a machinery house on such drills can prove to be a hindrance in frequent maintenance needed by the engine and other components. When the drills have to be used in extremely cold weather even diesel hydraulic drills also have to be equipped with a machinery house by suitably directing the air intake to and air exhaust from the diesel engine. Usually a machinery house is constructed from steel plates of adequate thickness. For operations in extremely cold weather the machinery house can be made from panels that have two plates welded on square tube steel sections as shown in Figure 7.59. Air entrapped within the steel plates acts as an insulator. To reduce the heat conduction further, the plates and tubing are coated with special paints.
7.6.4
Machinery house heater
Whether single wall or double wall, a machinery house is required to be heated so the temperature of the air inside the enclosure is comfortably high for machines to work and service personnel to carry out maintenance procedures. Heaters can be provided in machinery house for this propose. These heaters are heating coil or induction type. They have fans for pushing the hot air in different directions. Depending upon the range of low temperatures and the size of the machinery house, heaters can have ratings even up to 50 kW. The heater has a control to increase or reduce the quantity of heat generated.
7.6.5
Operator cab heater
For use in cold weather, an operator’s cab is often equipped with a cab heater. For extremely cold weather this heater proves insufficient, so another heater is provided. The rating of this additional heater can be up to 10 kW. This additional heater can be switched off during the daytime or in summer when the temperatures are not at extremes. Cab heaters can also be in the form of a combination of an air conditioner and heater. As they are expensive compared to the normal heaters, they are chosen on very few occasions.
7.6.6
Other enclosures and heaters
Apart from the aforesaid specific cold weather protection components, many other assemblies in a blasthole drill store or circulate water or oil. They are not located within the machinery house. If the temperatures in cold weather are not at extremes, insulation of these assemblies can prove sufficient, but for operations in extreme cold weather these assemblies are required to have special devices in the form of enclosures and heaters. Heaters are heated coil or infrared type. Hot air generated by the heaters is circulated by means of fans in the enclosures so the temperature around the assemblies is kept at acceptable level. The following are some of the assemblies that are provided with such enclosures and heaters: 1 2
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3 4
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Propel Gear Case Rotary Head
7.6.7
Protective coating on steel components
Heat is always generated in many components of a blasthole drill when it is working. This heat keeps most of the drill components and structural members at temperatures at which the mechanical properties of the materials remain at an acceptable level. To reduce the flow of heat from the drill components and the structural members to the extreme cold weather around them a coating of special chemicals and paints are can be used. Manufacturers coat the components and structural members of their drill with such insulating chemicals/paints when a user orders such coating at the time of its purchase.
7.7
COMPARISON OF TYPES OF BLASTHOLE DRILLS
As seen earlier, blasthole drills work either on internally generated diesel engine power or externally available electric power. The power is then distributed to various actuators in the drill through a hydraulic or electric medium. The difference in the mode of power distribution makes it necessary to have different devices on respective drills as given in Table 7.8. The following are some comparative comments in respect of the power source: 1 2
Drills with diesel engines have unrestricted mobility, whereas the mobility of the electric drills is restricted by the length of power supply cable in the cable reel. Additional items required with all-electric drills are costly so all-electric drills cost more than equivalent diesel drills.
Table 7.8 Need for some devices on the drill.
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Aspect
Diesel engine power source
External electric power supply
Fuel Tank Radiator Air Cleaner Battery Starting Transformer Converter Cable Reel Fuel Oil Engine Oil Transformer Oil
Needed Needed Needed Needed Not needed Not needed Not needed Needed Needed Not needed
Water Coolant Enclosure
Needed Not needed
Not needed Not needed Not needed Not needed Needed Needed Needed Not needed Not needed Not needed for Dry Transformers Usually not needed Needed
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5
6 7 8
221
The cost of power generated by diesel engines is very high as compared to that of an external electric supply. Consumable requirements of diesel powered drills is very high compared to allelectric drills. Transformers used on modern drills are of dry type and do not require transformer oil. Reliability of power supply from an engine is excellent provided the regular maintenance procedures are followed. External electric supply is also equally reliable or perhaps better, but without much maintenance. Exhaust of a diesel engine causes air pollution to some extent. External power supply does not cause any local air pollution. Maintenance requirements of various components in the case of diesel engine power are far greater as compared to external electric power. Starting a diesel engine in cold weather is difficult. In extremely cold weather a diesel engine may have to be run all the time irrespective of the use of power. In case of electric motors and other devices that operate on external electric supply the difficulties in starting and running are much less.
A comparison between hydraulic power distribution systems and electric power distribution on a rotary blasthole drill is given in Table 7.9. Specific elaborations in this regard are as follows: 1
2
3
4
Hydraulic oil used in hydraulic power distribution systems is a consumable item. It is required to be changed periodically and incurs expenditure. No such consumable exists in electric power distribution. Hydraulic power distribution systems require many components to carry and control the hydraulic oil flow as compared to the electric power distribution system. Power cables used for electric power transmission from source to the actuator are virtually jointless but the tubes and hoses used for carrying hydraulic oil contain bents, tees and other similar components. Once the electric cables are laid properly they or components within them may not have to be changed for as much as two decades. This is because the electric flow does not exert any physical pressure on the component hence no question of contraction, expansion and resulting fatigue or rupture.
Table 7.9 Comparison between hydraulic power distribution and electric power distribution.
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Aspect
Hydraulic power distribution
Electric power distribution
Hydraulic Oil No. of Components Type of Control Distribution Losses Component Weight Control or Motor Speed Maintenance Component Life
Needed Many Electrohydraulic Very High Motors are Lightweight Excellent High Maintenance Relatively Short
Not needed Fewer Electric Low Motors are Very Heavy Good Very Low Maintenance Very Long
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5
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In hydraulic systems physical pressure is exerted on tubes, hoses, valves etc. Apart from their susceptibility to rupture or leakage the fatigue resulting from the pressure changes does reduce the life of the component. Further physical movement of parts and hydraulic oil does result in a small wear of the component. As the hydraulic motors have very low weight and low inertia they stop rotation almost instantaneously once the power is switched off. Electric motors keep on running due to inertia for a very short duration after the power is switched off. Thus, hydraulic motors can be controlled precisely. Linear movement with heavy force can be easily generated by hydraulic means but not easily with electric means.
Above point nos. 5 and 6 make it desirable to have hydraulic power distribution systems even on all electric blasthole drills for leveling jacks and propel motors. In summary it can be said that all-electric drills are very economical to use as compared to equivalent diesel hydraulic drills. However, circumstances as under dictate the use of diesel drills: 1
2
Several small and medium size surface mines use small and medium size equipment. They choose to operate the equipment on diesel power as the expense of providing a high voltage electric grid are very high. The economy achieved by use of electric machines like draglines, shovels, drills etc. is not enough to offset the cost of laying electric grid. Electric supply is usually not available at new mine sites in the beginning of mining operations so there is no alternative to using diesel hydraulic blasthole drills.
If satisfactory electric supply is expected at a later Date, it is wise to choose diesel drills in the initial phase, which at a later date can be converted to electric drills by replacing the diesel engine with an electric motor. Once the electric supply is fully functional all-electric or electric blasthole drills can be procured.
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Chapter 8
Compressed air and air compressors
8.1
INTRODUCTION
Compressed air is the invariable medium for flushing the blasthole. It plays a very important role in the blasthole drilling process because in many cases the power consumed by the air compressor on a rotary blasthole drill is more than the power consumed by any other of its devices like rotary head, feed mechanism, propel mechanism etc. This chapter deals with many aspects related to the generation of compressed air in an air compressor. Blasthole flushing is covered in a separate chapter due to its specific importance.
8.2
COMPRESSED AIR
Compressed air is nothing but atmospheric air that is pressurized in a device called a compressor. A compressor sucks the atmospheric air and applies pressure to it. In this process the air becomes thicker i.e. its density increases. To keep the density at the desired level the air absorbs and stores energy supplied by the compressor. If the valve of a container containing compressed air at high pressure is opened and compressed air is allowed to flow to atmospheric air, the compressed air flows till such time as the pressure of air in the container reduces to the atmospheric pressure. Energy stored in the compressed air in the container is used for such flow. If compressed air is made to flow through pipes, the energy stored by the compressed air also flows with it. Thus compressed air can be used as a medium of energy transmission. Every atom/molecule of the atmospheric air is being pulled towards the center of gravity of the earth as per the laws of gravitation. The weight of the atom/molecule is being passed on to the atom/molecule just below i.e. nearer to the center of the earth. Thus atmospheric air also exists at a certain pressure. The atmospheric pressure is highest at the points on the surface of the earth that lie nearest to its center and keeps on reducing as we move away from the center. Thus atmospheric pressure at an altitude is always lower than the atmospheric pressure at sea level. If we imagine a container taken far away – say 5000 km from the surface of earth – there will be no air outside the container. Naturally, the level of pressure experienced at such height will be the minimum possible. Such pressure is called absolute zero pressure.
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A barometer is used to measure the pressure of atmospheric air. Pressure gages are used for measuring the pressure of compressed air above the barometric pressure. Therefore, Absolute Pressure = Barometric Pressure + Gage Pressure. In the International Standard (SI) System of units of Measurements, pressure is measured in terms of pascals, abbreviated as Pa. One pascal equals to a force of one newton applied on one square meter area. One kilogram of mass on the earth exerts 9.80665 newton force on the earth because the gravitational constant of the earth is equal to 9.80665 m2/s. Every system in the world contains energy. A fundamental law in this regard states that energy can neither be created nor destroyed but can only be transferred and converted. Energy has to be spent in doing every form of work. As a corollary it can be said that the amount of work done on or by a system is equal to the amount of energy transferred to or from the system. Like every other physical system that exists in the world, atmospheric air also has a certain temperature. Temperature is an indicator of the intensity of heat energy contained in the system. In the International Standard (SI) System of units of Measurement, temperature is measured in Celsius and is abbreviated as °C. Originally the same unit of measure was called centigrade. When the purest form of water is about to start freezing its temperature level is considered to be 0 and when it was about to start boiling it is considered to be 100. One hundredth part of the difference between these two levels is considered equal to 1°C. If, in some way, energy from a system is removed completely, the system will attain the minimum possible temperature called absolute zero temperature. The level of absolute temperature is 273.16°C below the freezing temperature of water. When a temperature of any system measured in °Celsius is TC, we have the absolute temperature of the substance as TC + 273.16. Kelvin is a unit of temperature measurement on the absolute scale basis. Magnitude of one degree Kelvin is the same as one degree of Celsius. Therefore, TK = TC + 273.16 Figure 8.1 clarifies the concept of absolute and gage pressure as well as absolute temperature measured in °K and thermometer temperature measured in °C.
Pressure Measured by Gauge
Temperature Measured by Thermometer
Gage Pressure
Thermometer Temperature in °C
Absolute Pressure
Absolute Temperature Atmospheric Pressure Measured by Barometer
Lowest Possible Pressure
273.16°C
Lowest Possible Temperature
Figure 8.1 Concept of absolute pressure or temperature.
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Pressure, volume and temperature are the fundamental properties of compressed air. The subject that deals with these is called Thermodynamics. When air expands or gets compressed the expansion or compression can take place in two manners viz. isothermal or isoentropic. In isothermal expansion or compression, the change in volume takes place without any change in the temperature of the system. In isoentropic expansion or compression the change in volume takes place without gaining or losing the heat in the system.
8.2.1
Gas laws
Behavior of gases is in accordance with the following laws: 1 2 3 4 5 6 7
Boyle’s Law Charles’ Law Guy Lussac’s Law Joule’s Law Poisson’s Law Amagat’s Law Avogadro’s Law The principles elaborated by these laws are as given below.
8.2.1.1
Boyle’s law
Boyle’s law states: “The absolute pressure of a given mass of perfect gas varies inversely as its volume when the temperature remains constant.” i.e. p ∝ 1/v. In other words, if the absolute pressure and volume of a gas system before the process were p1 and v1, and became p2 and v2 after the process, then p1 * v1 = p2 * v2 = Constant As the temperature of the gas remains constant the process is called an isothermal process. 8.2.1.2
Charles’ law
Charles’ law states: “The volume of a given mass of perfect gas varies directly as its absolute temperature when the pressure remains constant.” i.e. v ∝ T. In other words, if the absolute temperature and volume of an air system before the process were T1 and v1 and became T2 and v2 after the process, then v1/T1 = v2/T2 = Constant As the absolute pressure of the gas remains constant the process is called an isobaric process.
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8.2.1.3
Guy lussac’s law
Guy Lussac’s law states: “The absolute pressure of a given mass of perfect gas varies directly as its absolute temperature, when the volume remains constant.” i.e. p ∝ T. In other words, if the absolute pressure and absolute temperature of an air system before the process were p1 and T1 and became p2 and T2 after the process, then, p1/T1 = p2/T2 = Constant As the volume of the gas remains constant the process is called an isochoric process. 8.2.1.4
Joule’s law
Joule’s law states: “The change in internal energy of a perfect gas is directly proportional to the change of temperature.” i.e. ΔE ∝ ΔT. In other words, if the temperature of a mass of a gas before the process was T1 and it became T2 after the process, then ΔE = m * C * ΔT = m * C * (T2 − T1) Where m is the mass of the gas and C is a constant of proportionality called specific heat. 8.2.1.5
Poisson’s law
Poisson’s law states that in the case of a gas, for a process without any heat exchange with the surroundings, the relationship between pressure and volume is as per the following equation p1 * v1(Cp/Cv) = p2 * v2(Cp/Cv) In the equation p1 and v1 represent pressure and volume before the process and p2 and v2 represent pressure and volume after the process. Since no heat is exchanged with surroundings, the process is isoentropic compression or expansion. The quantities represented by Cp and Cv are the specific heat capacities for the gas at constant pressure and constant volume. Often the ratio Cp/Cv is represented by k and has values as follows: k = 1.66 for monoatomic gases k = 1.40 for diatomic gases k = 1.30 for triatomic gases Since more than 98% of the gases contained in atmospheric air are diatomic i.e. those gases where two atoms of the gas form a stable molecule (e.g. nitrogen, oxygen), for atmospheric air the ratio k is taken as 1.40. The ratio Cp/Cv is fairly constant at low pressures.
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Compressed air and air compressors
8.2.1.6
227
Amagat’s law
Amagat’s law states that the volume of a mixture of gases is equal to the sum of the partial volumes that the constituent gases of the mixture would occupy if each of them existed alone at the total pressure of the mixture. 8.2.1.7
Avogadro’s law
Avogadro’s law states that under the same conditions of pressure and temperature, equal volumes of all gases contain an equal number of molecules. In other words 1 m3 of hydrogen, in its molecular form H2, will contain the same number of molecules as contained in 1 m3 of oxygen in its molecular form O2, when the temperature and pressure of both systems are the same. In such cases, since the molecular weights of hydrogen and oxygen are 2 and 32 respectively, the densities of hydrogen and oxygen will be in proportion of 2:32 i.e. 1:16. A mole is a scientific term given to a numerical quantity of molecules. The number of molecules contained in a mole is 6.02257 * 1027 and is often termed as Avogadro’s Number. It can be shown that at the NTP conditions i.e. temperature of 0°C and pressure of 101.325 kPa, the volume of one mole of any gas equals to 22.3933 m3. 8.2.1.8
General gas law
By combining Boyle’s law and Charles’ law we can write an equation p * v/T = Constant. This formula is followed in all the relevant gas calculations. Following example illustrates its use. EXAMPLE
Initial conditions of a gas before compression are volume = 1 m3, gage pressure = 300 kPa, temperature = 27°C. It is compressed to attain a volume = 0.2 m3, temperature = 90°C. What will be the final gage pressure? SOLUTION
The initial conditions of volume, absolute pressure and absolute temperatures are, v1 = 1 m3,
p1 = 300 + 101.325 = 401.325 kPa,
T1 = 27 + 273.16 = 300.16°K Post compression conditions are, v2 = 0.2 m3,
T2 = 90 + 273.15 = 363.16°K
Applying the general gas law we get, (401.325) *(1)/(300.16) = p2 * (0.2)/(363.16)
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This gives p2 = 2427.79 kPa. Since p2 is the final absolute pressure the final gage pressure will be 2427.79 − 101.325 = 2326.465 kPa.
8.2.2
Power required for air compression
An air compressor is a device that takes power from an external power source like a diesel engine or electric motor and apples it mechanically to the air so as to reduce its volume and increase its pressure. In such compression processes heat is neither gained nor lost by the system. The compression process is therefore adiabatic. A schematic illustration of a typical piston compressor is shown in the upper drawing in Figure 8.2. The lower drawing shows an indicator diagram of the compressor work in a complete cycle. The piston moves to left and right in the cylinder to compress the air. When the piston is at the beginning of the compression stroke (point C, pressure p1 and volume v1)
Outflow Pipe
Outflow Valve
Inflow Inflow Valve S
E
Piston at the end Piston at the of Compression beginning of Stroke Compression Stroke D pvn = Constant Δp
V P2 V2
T
B
A
C
P1 V1
Figure 8.2 Cycle for isothermal compression of air.
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the intake valve is open but as soon as it moves a little to left the air in the cylinder gets compressed and the inflow valve closes. From here the compression starts as per Poisson’s law along the curve CD. When the piston is at a certain intermediate position (point D) the pressure inside the cylinder becomes equal to the pressure in the outflow pipe and the outflow valve opens. From this point onwards as the air flows into the outflow pipe the pressure of the air does not increase. This continues till the piston reaches the end of the compression stroke (point E, pressure p2 and volume v2) As the piston starts moving to the right the outflow valve closes and the pressure of the air contained in the cylinder starts decreasing. At a certain intermediate position (point B) of the piston the pressure inside the cylinder becomes lower than the pressure in the inflow pipe and the inflow valve opens. The piston continues its movement towards the right till the end of the stroke but the pressure in the cylinder remains the same. The volume of air contained in the cylinder when the piston is at the end of compression stroke is called the clearance volume. For air compression or expansion process a general equation, p * vk = K, can be applied. For isothermal compression or expansion k = 1 as per Boyle’s Law. For isoentropic compression or expansion k = 1.4 as per Poisson’s Law. In actual air compression by an air compressor the compression process is associated with a change in temperature as well as a change in energy. Thus the process is neither isothermal nor isoentropic but called polytropic. For positive displacement compressors the value of k ≈ 1.3. The net power required by the compressor to compress air is equal to the work done in compressing the air. Most of the compressors used on rotary blasthole drills are single stage rotary vane or rotary screw compressors i.e. positive displacement compressors. For such compressors the net power requirement can be found from the following equation for work done in one cycle of compression. W = [k/(k − 1)] * p1 * v1 * [(p2/p1)[(k−1)/k] − 1] If the compressor is rotating at a speed of N rpm i.e. N/60 revolutions per second then the power requirement will be P = {k/[(k − 1) * 60]} * p1 * v1 * N * [(p2/p1)[(k−1)/k] − 1] In the above equation, W = Work done per cycle in Nm k = Exponent for Polytropic Process p1 = Initial absolute pressure in kPa p2 = Final absolute pressure in kPa N = Compressor rotary speed in rpm v1 = Initial volume sucked in compressor m3 P = Net power required by compressor kW The volume v1 is taken as the discharge capacity of the compressor.
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In some compressors a small volume of the compressed air does not go into the high pressure section but returns to the atmosphere. If this returned volume is taken as vs then the volumetric efficiency of the compressor is taken as ηv = (v1 – vs)/v1 If the compressor on a rotary blasthole drill is a two-stage compressor, it is normally equipped with an intercooler which cools the air compressed in the first stage and then sends it to the second stage for further compression. The cooling of the air in the intercooler can either be complete i.e. the temperature of the air compressed in the first stage is lowered to the atmospheric temperature or incomplete i.e. the temperature of the air compressed in the first stage is not lowered to atmospheric temperature. For compressors with complete intercooling the equation for the power requirement is P = {k/[(k − 1) * 60]} * p1 * v1 * N * {(p2/p1)[(k−1)/k] + (p3/p2)[(k−1)/k] − 1} and for compressors with incomplete intercooling the equation for power requirement is P = {k/[(k − 1) * 60]} * N * {p1 * v1 * [(p2/p1)[(k−1)/k] − 1] + p2 * v2 * [(p3/p2)[(k−1)/k] − 1]} In both the above equations, p1 = Initial absolute pressure in kPa p2 = Interstage absolute pressure in kPa p3 = Final absolute pressure in kPa all the other terms have the same meanings as in the earlier case of single stage rotary vane or rotary screw compressor. Usually the compressor characteristics state the compressed air volume delivered along with final pressure of compressed air. In such cases the initial pressure is treated to be the one defined by the rating conditions. For such situations the whole volume is considered to have occurred in one cycle and, therefore, the value of N is taken as 1 while evaluating the power requirement by the above equation. The following is a sample example that calculates the power required by a single stage rotary vane compressor. In many cases the conditions of temperature and pressure are expressed as NTP or STP rather than expressing specific values of temperature and pressure. In such instances NTP means temperature of 273°K and pressure of 101.325 kPa and STP means the standard atmospheric condition defined by the particular standard (SAE, BS, DIN etc.). These are specified for some standards in Table 11.3 of chapter 11. Two other terms that are often heard in connection with compressed air are SCFM and ACFM. Here CFM means the discharge from compressor in imperial unit cubic feet per minute. S stands for standard conditions.
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Due to the presence of moisture in the air the actual discharge is a little less than that given by the SCFM rating. This actual discharge is termed ACFM. Pressure is often given in terms of MPa, kPa, bar or psi. In such case 1 MPa = 10 bar = 1000 kPa = 145.038 psi. Similarly, for flow 1 m3/min = 35.31466 ft3/min. = 0.01666 m3/s = 16.666 L/s. EXAMPLE
A vane compressor on a blasthole drill delivers 30 m3 compressed air at a gage pressure of 6.89 bars at NTP condition. What is the power required by the compressor to compress the air? SOLUTION
In this case we presume that the compression is polytropic and the value of k = 1.3. For NTP conditions we have p1 = 101.325 kPa. Final gage pressure is 6.89 bars i.e. 689 kPa. Thus, absolute pressure p2 is 689 + 101.325 = 790.325 kPa. Volume delivered in one minute is equal to 30 m3. Evaluating the above equation with these values gives the net power requirement as P = 133.138 kW.
8.3
FLOW OF COMPRESSED AIR
The compressed air generated by a compressor on a rotary blasthole drill has to flow through a specific flow path. Conduits of circular cross section are far better in withstanding the high pressure of compressed air without undergoing distortion than conduits with other shapes of cross section. Therefore, almost invariably a circular cross section is chosen for the flow path. As far as rotary blasthole drilling is concerned, compressed air is used for flushing. For this purpose it is made to flow through circular pipes and pipe fittings as shown in Figure 8.3. If the flow path schematic in Figure 8.3 is minutely observed one can easily comprehend the following. After a little drop in elevation the compressed air flow path horizontally goes up to the mast. Steel tubes containing valves, bends etc. are used as a flow path. The length of this flow path can be from 4 to 17 m depending upon the size of the blasthole drill. A further part of the flow path, that takes compressed air up to to the steel tubes on the mast, has to be curved and must have flexibility because the mast may have to be occasionally lowered and raised. Therefore, this part of the flow path is made up of hose pipe. The length of this part is between 1 to 3 m and a small part of the hose is vertical. Further flow path is formed of straight steel tubes. It remains vertical during the drilling operation. The length of this part can be between 6 m to even 20 m depending upon the drill size.
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Figure 8.3 Flow path in a typical blasthole drilling process.
From the vertical tubing the flow path to the moving rotary head of the blasthole drill is in the form of a hose. The length of the hose is between 8 to 20 m. This hose is always bent through a 180° curve. Naturally the air flows through it vertically, in the up as well as down direction. In the drill head again some steel tubes and other components take the compressed air to the drill pipes. The length of this part is between 1 to 3 m. Some of the portion is horizontal and some is vertical. The drill pipes take the compressed air to the tricone bit. This flow path can have a length from 6 to even 60 m. In this section air flows in a vertically downward direction. In order to clean the bottom of the hole very effectively, the flow of air is passed through the three nozzles of the tricone bit. In this process the pressure energy stored
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in the compressed air is converted into kinetic energy. Thus the speed of air increases considerably and the jet is able to clean the blasthole bottom in an effective manner. Finally the compressed air flows upward in the annulus along with cuttings of rock and escapes to the blasthole top. The large cuttings fall on the ground around the blasthole, whereas the air mixed with small sand and dust size cuttings is made to flow through a dust collector. After the filtration dustless air escapes to the atmosphere. The flow of compressed air is studied through the subject called ‘Fluid Mechanics’. From the above description of flow path in a blasthole drill, it is obvious that following topics of fluid mechanics must be considered. Flow through steel pipes Flow through fittings Flow through hose pipe Flow through nozzles Flow through annulus. Of the above five topics the last i.e. flow through the annulus is considered in chapter 11. When the pressure of flowing compressed air remains almost constant, the flow is considered as incompressible flow. In the compressed air flow path described above the pressure drop between the compressor to the top of the drill bit is usually very small, of the order of 35 to 50 kPa. Thus, the flow of compressed air through pipes comes under the realm of incompressible flow. When the pressure at a point on the flow path differs significantly from that at other point on the flow path, the flow is treated as per the technical formulation of compressible flow. Thus, the flow of compressed air through nozzles or orifices is compressible.
8.3.1
Compressed air flow in steel pipes
When compressed air flows through a flow path, energy stored in the compressed air is lost due to the friction of air molecules with the walls of the flow conduits. As a result, the pressure of compressed air reduces. It is important to know the quantum of pressure lost in the passage of compressed air up to the drill bit, or ideally up to the point where the compressed air emerges out of the blasthole i.e. at the mouth of the blasthole. The properties of air, the velocity of air flow, and the roughness of wall that comes in contact with the atoms of flowing air, make the state of the flow either smooth or disturbed or transient (when it is changing from smooth to disturbed). These states are referred to as laminar or turbulent or transitional, respectively. In laminar flow the molecules of air move in a straight line parallel to the axis of the pipe, but in turbulent flow the molecules move in a very haphazard manner, and in addition to the axial velocity also have velocity components in a plane perpendicular to the axis of the pipe. As far as the roughness of pipe is concerned, on a microscopic dimensional basis even the best polished surface of any pipe has ups and downs in the surface as shown
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in Figure 8.4. The average distance between the extent of these ups and downs is called absolute roughness of the pipe. It is commonly expressed by symbol ε. Further, a pipe has some inner diameter D. Relative roughness is the ratio ε/D. It can be easily calculated for any pipe from the values of absolute roughness of different pipe materials given in Table 8.1. The most important indicator of the type of flow is called Reynold’s Number. Reynold’s number is actually a dimensionless number, usually denoted by Re. It is related to some variables of the pipe, flowing fluid and the flow as follows: Re = (D * W * ρ)/ν where, Re = Reynold's number D = Inner diameter of pipe in m W = Flow velocity in m/s ρ = Density of fluid in kg/m3 ν = Dynamic viscosity of fluid in Pa · s Centreline of the pipe Relative roughness works out to ε/D Radius of the pipe = D/2 D is the diameter of the pipe Absolute roughness ε
Figure 8.4 Absolute and relative pipe roughness. Table 8.1 Absolute roughness values (ε) of some pipes.
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Material and type of pipe
Absolute roughness in mm
Copper, Lead, Brass, Aluminum (new) PVC and Plastic Pipes Stainless steel Steel commercial pipe Stretched steel Weld steel Galvanized steel Rusted Steel New Cast Iron Worn Cast Iron Rusty Cast Iron Sheet or Asphalted Cast Iron Smoothed Cement Ordinary Concrete Coarse Concrete
0.001–0.002 0.0015–0.007 0.015 0.045–0.09 0.015 0.045 0.15 0.15–4 0.25–0.8 0.8–1.5 1.5–2.5 0.01–0.015 0.3 0.3–1.0 0.3–5
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Dynamic viscosity of compressed air at various temperatures and pressures is given in Table 8.2. When Reynold’s number of flow is less than 2300 the flow is laminar. For Reynold’s number of 4000 or more, the flow is turbulent and in between 2300 and 4000, the flow is transitional. The most widely used equation for determining pressure loss in pipes is the DarcyWeisbach equation, viz. Δp = f * (L/D) * 10−3 * {(ρ * W2)/2} where, Δp = Pressure loss in kPa f = friction factor L = Length of pipe in m D = Hydraulic diameter of pipe in m ρ = Density of flowing air in kg/m3 W = Average flow velocity in m/s The assumptions made while deriving the above equation from fundamentals are 1 2 3
The pipe is without any curvature. Pipe roughness is uniform throughout its length. The pipe is horizontal. Thus the gravitational forces acting on the mass of flowing air do not change and can, therefore be neglected.
In the above equation all the other factors except D and f are either directly or indirectly known. The factor D in the Darcy-Weisbach equation, i.e. hydraulic diameter of the pipe, can be found by an equation viz. D = 4 * A/P where A = Cross sectional area of flow path. P = Wetted perimeter of the flow path.
Table 8.2 Dynamic viscosity of dry air.
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Absolute pressure in kPa
Value of dynamic viscosities of dry air in Pa · s ( i.e. kg/(m · s)) at different temperatures in °C −25
−20
−10
0
10
20
25
30
40
50
100 300 500 700
15.9000 15.9339 15.9700 16.0083
16.1504 16.1824 16.2166 16.2530
16.6348 16.6648 16.6965 16.7299
17.1000 17.1296 17.1600 17.1913
17.5492 17.5792 17.6091 17.6390
17.9856 18.0161 18.0458 18.0748
18.2000 18.2304 18.2600 18.2887
18.4124 18.4426 18.4718 18.5001
18.8328 18.8612 18.8893 18.9170
19.2500 19.2743 19.3000 19.3270
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For a circular tube it is 4πr2/πr = 2r i.e. diameter of the pipe. For the annular flow path between a blasthole of radius r and a drill pipe of outer radius r1 it is 4π(r2 − r12)/2π(r + r1) = 2(r − r1). For other shapes it can be easily calculated. When the flow conduit is in the form of a pipe with circular cross section the Darcy-Weisbach equation can be written as, Δp = f * (L/D) * 10−3 * {(8 * ρ * Q2)/(π2 * D4)} This form is particularly suitable for pneumatic calculations because here Q is in terms of m3/s and the compressor discharge, which is usually given as m3/min can easily converted into m3/s by dividing the discharge value by 60. If the flow rate of air is given in terms of mass of compressed air then the above equation gets converted into Δp = f * (L/D) * 10−3 * {(8 * m2)/(ρ * π2 * D4)} Where m is the mass flow in kg/s. In aforesaid Darcy-Weisbach equation the factor {(ρ * W2)/2} is often called dynamic pressure. Obviously it is equal to {(8 * ρ * Q2)/(π2 * D4)} or {(8 * m2)/ (ρ * π2 * D4)} as the case may be. The value of the friction factor f has a complex relationship with Reynold’s number of the flow and relative roughness. The approximation that is considered to be the most accurate is given by the Colebrook Equation, which is (1/f0.5) = −2.0 * log[{(ε/D)/3.7} + {2.51/(Re * f0.5)}] This equation is implicit. Therefore, the evaluation of f from this equation can be done only through an iterative process. In the days when electronic calculators or computers did not exist, a chart called Moody’s chart, shown in Figure 8.5, was used for evaluating f. It is almost a graphical representation of the Colebrook equation. In the era of calculators i.e. during 1970 to 2000 some simplified and explicit formulae were suggested for determining the friction factor f. Now, in the present era of very fast personal computers, generally f is calculated by using the Colebrook equation. Computer software, some of which is totally free, is available on the internet and can be downloaded for the purpose of calculation of pressure loss. It must be remembered that the friction factor f is sometimes referred to as Moody’s Friction Factor. The pressure loss measurements taken in the field and those calculated by the above equations do tally acceptably well in the case of liquids, but for air and gases the differences are quite high. In almost all blasthole drills, the capacity rating of the compressor and the diameters of the pipes used for carrying the compressed air to the drill head, the drill pipe, and the annulus formed by the walls of the hole and the drill pipes, are such that the flow through them is turbulent. The following example will clarify the method of calculating pressure loss in a pipe.
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Moody Friction Factor
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Reynold's Number Figure 8.5 Moody’s chart showing relation between friction factor, Reynold’s number and absolute roughness. 11/22/2010 2:38:16 PM
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EXAMPLE
A straight galvanized pipe of circular cross section with a diameter 140 mm and length 75 m is used for carrying discharge from a compressor having a capacity of 25.49 m3/min at a pressure of 1034 kPa. The temperature of the compressed air is 15°C and atmospheric pressure is 101.325 kPa. Find out the pressure loss. SOLUTION
Since the atmospheric pressure is 101.325 kPa, we get the actual air flow density ρ = (1034 + 101.325)/ (287.1 * (273.16 + 15)) = 13.723134. As the compressor discharge is 25.49 m3/min, the actual volume flowing through pipe will be = (25.49/(13.723 * 60)) * = 0.0309575 m3/s. A = 3.14159 * 0.142/4 = 0.0153937 m2 Therefore, W = 0.0309577/0.0153937 = 2.011033 m/s. With the above, the dynamic pressure ρ * W2/2 = 13.723134 * 2.0110662/2 = 27.749934. Since the pipe is galvanized, ε = 0.15 mm and relative roughness ε/D = 0.15/140 = 0.0010714. Dynamic viscosity at STP conditions μ will be 1.802 * 10−5. Reynolds Number Re will be ρ * D * W/μ 13.723 * 0.14 * 2.011033/1.802 * 10−5 = 214410. From the Moody chart the friction factor f will be 0.0195 and the pressure loss will work out to Δp = f * (L/D) * [(ρ * W2)/2] Δp = 0.022 * (75/0.14) * 27.749934 Δp = 313.96 Pa. EXAMPLE
Find out the air pressure loss experienced in a drawn circular galvanized pipe of diameter 152 mm and length 50 m when compressed air at the mass flow rate 3.7 kg/s, gauge pressure of 448 kPa and temperature of 20°C is flowing through it. The atmospheric conditions are STP. SOLUTION
From thermodynamics fundamentals we have p*v=m*R*T In this case p = 448 + 101.325 = 549.325. m = 3.7 and T = 20 + 273.16 = 293.16. Therefore, v = 3.7 * 287.1 * 293.16/(549.325 * 103) = 0.556905 m3/s.
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Since the diameter of the pipe is 152 mm, the cross sectional area of the pipe A will be A = π * 1522/(4 * 106) = 18145.8 * 10−6 m2 and the velocity of the flow W will be W = v/A = 0.55690/(18145.8 * 10–6) = 31.24195 m/s As p * v = m * R * T, p = ρ * R * T or ρ = p/(R * T). In this case, ρ = 549.325 * 103/(287.1 * 293.16) = 6.52667 kg/m3 The dynamic viscosity of air for 20°C temperature and 448 kPa pressure, is 1.825 * 10−5. Reynold’s number, therefore, will be Re = 0.152 * 31.24 * 6.52667/(1.825 * 10−6) = 1698264. As the pipe is a galvanized pipe the absolute roughness will be 0.00015 m and relative roughness will be 0.00015/0.152 = 0.00098684. From Moody’s chart shown in Figure 8.5 we get the friction factor as 0.0195. With all the data evaluated above the pressure loss works out to Δp = 0.0195 * (50/0.152) * (6.52667 * 31.242/2) i.e. Δp = 8172 Pa. 8.3.1.1
Simplified formulae
As stated earlier, many explicit equations have been set forth for determining the pressure loss in the flow of compressed air in circular pipes. Popular amongst them are: 1 2
Harris Formula Engineering Toolbox Formula
These formulae are semi-empirical in nature. They are not derived on a purely theoretical basis but contain some part that is based on experimental findings. None of these contain any reference to the absolute or relative roughness of the pipe. Both these formulae involve simple calculations of some readily measurable parameters. 8.3.1.1.1
Harris formula
The original Harris formula is in imperial units as given below. Δp = 0.1025 * Q2 * L/(d5.31 * p)
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where Δp = Pressure loss in kg/cm2 Q = Flow volume in ft3/s L = Length of pipe in ft d = Inner diameter of pipe in inches p = Compression ratio The following is an SI conversion of the formula. Δp = 8.333307 * 1013 * Q2 * L/(d5.31 * p) where Δp = Pressure loss in Pa Q = Flow volume in m3/min L = Length of pipe in m d = Inner diameter of pipe in mm p = Compression ratio 8.3.1.1.2
Engineering toolbox formula
Δp = 728 * Q1.85 * L * 1012/(d5 * p) where Δp = Pressure loss in Pa Q = Flow volume in m3/min L = Length of pipe in m d = Inner diameter of pipe in mm p = Initial gauge pressure of air in Pa It can be easily seen that the above formula does not take into consideration the temperature of the flowing air. The above equation is based on the absolute roughness for standard seamless carbon steel pipes of schedule 40 or schedule 80. More information and the air pressure loss values for steel pipes are given in Appendix 4 at the end of this book.
8.3.2
Compressed air flow in hose pipes
A pipe is meant to carry fluid flow from one point to the other. If the distance between these two points and their position remains unchanged it is highly beneficial to use steel pipe for reasons of cost, safety, long life etc. However, if the distance between the points and/or their position with respect to each other changes, the only alternative is to use a flexible hose pipe. In past two decades several new materials have been used in the manufacture of hoses. Due to their superior characteristics such as durability, flexibility etc., they are often preferred to the rubber hose pipes used earlier.
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Most of the modern generation of hose pipes that are used for carrying fluids at high pressure are braided, either with nylon strings or steel wires. This braiding prohibits their enlargement in diameter and allows them to carry fluids at high pressures without losing their flexibility. On a rotary blasthole drill it becomes essential to use hose pipes at two distinct places as follows: 1
2
Where the flow conduits have to cross from the main frame to the mast, which has to change its position with respect to the frame. This section is not very critical because the length of hose required for this crossover is usually rather short. Besides this, the mast is not raised or lowered frequently hence hose bending takes place only rarely. Where the flow conduit on the mast has to go to, and remain attached to, the moving drill head. This section is very critical because the distance between the points changes by a very great length. Therefore hose pipes have to be very long. Further, the relative position of the points with respect to each other changes every fraction of a second during most of the time when the drilling is carried out.
On some modern blasthole drills a hose tray as shown in Figure 8.6 is provided to carry hydraulic hoses, electric cables and air hoses from the mast to the drill head. This device greatly reduces the maintenance and ensures longer life to all the air hoses, hydraulic hoses and electric cables carried through it. Every hose has some straight portion and some curved portion. No separate fitting is involved in a hose except the end connectors. For a straight portion of the hose, pressure loss can be calculated with the same equations given in the last section for steel pipes but by treating the pipe as having very low roughness. Appropriate values of absolute roughness to be used are to be obtained from the manufacturers of the hose pipe. Since hose pipes are made from plastic or rubber with a special smooth inner lining, absolute roughness of 0.0015 mm can also be used. Appendix 5 at the end of this book gives pressure loss experienced when compressed air flows through hose pipes.
8.3.3
Compressed air flow in pipe fittings
Whenever a pipe fitting is built into a pipe line it creates an obstruction in the flow path. This results in additional pressure loss in the pipe lines. Two identical pipe lines are shown in Figure 8.7. The upper pipeline has a valve, whereas, the lower pipeline is without a valve. Other parameters of the two pipelines are same. Since the part of the pipelines on the left hand side of the plane AA is identical, the average pressure P1 across the pipe cross section in the plane AA is the same in either case. However, the pressure at the pipe cross sections in the plane BB differs in the two cases. At this plane the pressure P2 in the pipeline without a valve is higher than the pressure P3 in the pipeline with a valve.
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Figure 8.6 Hose tray to carry hose pipes from mast to rotary head.
A
B
Pressure P3 Pressure P1 Pressure P2
A
B
Figure 8.7 Pipe line with and without pipe fitting.
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This additional pressure loss (P2 – P3) is solely due to the presence of the valve. Let this difference be BF. It can be shown that BF = KL * (ρ * W2/2) where, BF = Additional pressure loss due to valve KL = Resistance coefficient ρ = Density of flowing fluid W = Average flow velocity Fittings in pipeline are generally divided into four main categories on the basis of their purpose as follows: 1 2 3 4
To change the direction of flow To control the flowing volume To connect pipelines of different diameters To divide the flow
Due to the complexity of flow, theoretical calculations of the resistance coefficient are virtually impossible. Hence, the values of resistance coefficient for different types of pipe fittings are expressed numerically. If a pipeline has length L and diameter D, and contains three fittings having resistance coefficients of KL1, KL2, KL3 then the total pressure loss in the pipeline can be equated as, ΔP = (f * (L/D) + KL1 + KL2 + KL3) * ρ * W2/2 In many cases, particularly in the case of valves, the resistance coefficient is expressed as the equivalent length Le. Equivalent length means the additional length of the pipe that would have given the same additional pressure loss occurring due to the presence of the fitting. In other words, KL1 = f * (Le1/D) In such cases the pressure loss equation will become ΔP = (f * (L/D) + f * (Le1/D) + f * (Le2/D) + f * (Le3/D)) * ρ * W2/2 Appendix 6 also gives the values of equivalent lengths for different types of valves.
8.3.4
Effect of ups and downs in the flow path
In addition to the horizontal part, a flow path often contains sections that are not horizontal. They may be inclined downward or upward or can even be vertical in upward or downward directions. This is very much so in a blasthole drill. An inclined pipe is shown in Figure 8.8. The fluid flow in this pipe is in an upward direction as shown by the arrow. It can be proved from fundamentals that if ΔPo is the pressure loss in the fluid over the length L then because of the rise in the pipeline the pressure loss is equated as
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ΔPo = ΔP + ρ * μ * L * sinθ For a flow path inclined downwards, the angle θ will be negative, therefore the equation works out to ΔPo = ΔP − ρ * μ * L * sinθ If the pipes are in a vertical direction and the flow is in an upward or downward direction, then the pressure losses will be ΔP + ρ * μ * L and ΔP – ρ * μ * L respectively. The above considerations of ups and downs in the flow path are very important in the case of flow of liquids but are negligible in the case of air flow because the density and viscosity of air are of very small values. In blasthole drills the pressure loss due to ups and downs in the flow path can be safely neglected because the length is only of the order of 30 m or so.
8.3.5
Effect of curvature in the flow path
Pipe fittings, like elbows, bends etc., used for changing the direction of the air flow, have fixed dimensions. The dimensions of their curvature are invariably dependent upon their diameter. In such case the values for additional pressure loss due to their presence is calculated by use of a resistance coefficient as stated earlier. However, when a hose pipe is used as a conduit for air flow, the dimensions of the curvature are not fixed and bear no relation to the hose diameter. Therefore, the pressure loss has to be evaluated on an individual basis. The resistance coefficient of the curved hose, denoted by ζ is very similar to the resistance coefficient of pipe fittings. Additional pressure loss incurred by the presence of curvature in the hose is given by BH = ζ * (ρ * W2/2) where, BH = Additional pressure loss due to curvature of the hose ζ = Resistance coefficient ρ = Density of flowing fluid W = Average flow velocity The resistance coefficient is dependent upon the hose diameter and radius of hose curvature.
L H Point B
θ
Figure 8.8 Forces acting on fluid in inclined pipe.
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Referring to the dimensions of the hose in a curved state as shown in Figure 8.9, the values of the resistance coefficient ζ are dependent upon the diameter of hose, the radius of curvature of the hose path and the angle through which the curved path passes. Table 8.3 gives values of the resistance coefficient for different values of the ratio R/d and different angle of curvature. It is to be clearly understood that the resistance coefficients are for a circular curved path of the pipe in a horizontally-curved alignment. Effect of height is to be treated separately.
8.3.6
Compressed air flow through nozzles
Compressed air always flows through an opening of any shape from a high air pressure container to a low pressure container. An orifice is simply a hole, made in the wall of a high pressure container that allows fluid to flow. It is shown in the right hand side drawing in Figure 8.10. A nozzle, as shown in the left hand drawing in Figure 8.10, is also an opening and allows fluid to flow, but as can be seen in the drawing, it is specially designed to streamline the flow. Therefore, it allows the fluid to flow at high efficiency. Considering the open atmosphere as the low pressure container, a schematic of the conditions is shown in Figure 8.10. An equation that correlates various influencing factors is as follows: M = α * ψ * p1 * A * [2/(R * T1)]0.5
L
d
θ R
Center of Hose Curvature
Figure 8.9 Forces acting on fluid in curved pipe.
Table 8.3 Values of resistance coefficient ζ applicable for curved hose pipes.
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Different values of the angle of curvature
For different values of the ratio R/d 1
2
4
6
10
30° 60° 90° 120° 150° 180°
0.1440 0.2520 0.3600 0.4500 0.5400 0.6120
0.0880 0.1540 0.2200 0.2750 0.3300 0.3740
0.0680 0.1190 0.1700 0.2125 0.2550 0.2890
0.0600 0.1050 0.1500 0.1875 0.2250 0.2550
0.0520 0.0910 0.1300 0.1625 0.1950 0.2210
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p1T1
pnTn
p1T1
pnTn p2T2
p2T2 D should D be more D should
than 2.5 dn dn
be more
dn
than 2.5 dn
Figure 8.10 Nozzle and orifice.
where M = Mass flow rate in kg/s α = Nozzle efficiency ψ = Flow coefficient p1 = Upstream absolute pressure in Pa T1 = Upstream absolute temperature in K A = Area of nozzle opening in m2 R = Gas constant in J/(kg · K) Some basic points to be understood clearly while using this equation are given below. The shape of the opening influences the flow rate to a very great extent. In practice, almost all man-made openings are of circular cross section. Whatever the type of opening it may be, there is always friction between the molecules of air and the walls or edges of the flow conduit. Due to this, resistance is caused to the flow and the actual flow decreases. Very well-shaped and highly polished circular nozzles can have an opening efficiency α as high as 0.97, whereas well-shaped circular orifices can have opening efficiency α of 0.61 at best. Flow volume also depends upon the ratio of downstream absolute pressure p2 to upstream absolute pressure p1. If the downstream pressure p2 is reduced, the pressure ratio p2/p1 also reduces and the flow through the nozzle keeps on increasing till the ratio reduces to a certain value. At this value of the pressure ratio, the flow is at maximum because for any lower value of the ratio the flow remains the same as that of the maximum. This pressure ratio p2/p1 is called the critical pressure ratio. Thermodynamically it can be shown that in the case of compressed air this critical pressure ratio is 0.5282, and for this value the flow coefficient ψ is equal to 0.4841. For air, the gas constant R is equal to 287.1 J/kg · K. In this equation the flow rate is given in terms of the mass and not volume of the compressed air, because the mass of flowing compressed air remains constant while the volume of compressed air changes at every point in the flow path. Here the mass of the flow is with respect to a pressure of p1. As the mass flow is in terms of kg/s, to convert it into conventional volume flow in standard conditions used for air compressors i.e. m3/min, we must multiply by 60 and divide by 1.225. This multiplier works out to 48.97959.
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EXAMPLE
From a high pressure vessel containing compressed air at 3 MPa gauge the air flows through a well made circular nozzle of diameter 5 mm to a low pressure vessel having air pressure of 600 kPa gauge. The temperature of the air in the high pressure vessel is 35°C. What will be the mass flow rate? SOLUTION
In this case the pressure ratio is (600 + 101.325)/(3000 + 101.325) = 0.2336. Therefore, the flow will remain maximum and is to be calculated by using the flow coefficient ψ = 0.4841 Since the nozzle is well-made we can assume the nozzle efficiency to be 0.95. We also have R = 287.1 J/kg · K. p1 = 3000000 + 101325 = 3101325 Pa T1 = 35 + 273.16 = 308.16°K A = π * (5/1000)2/4 = 0.00001963 m2 Hence M = 0.95 * 0.00001963 * 0.4841 * 3101325 * [2/(287.1 * 308.16)]0.5 or M = 0.133151 kg/s or Q = 6.5217 m3/min.
8.3.7
Leakage of compressed air
Leakage of compressed air from the system is to be Avoided, as leakage of compressed air means waste of energy that has been used in its compression. Pipes used for the flow of compressed air are coupled either by welded or by threaded connections. A carefully welded connection does not have any leakage. Good threaded connections can have some leakage but being small it is treated as no leakage. How much compressed air is lost in leakage through leak paths depends upon many factors such as size and shape of the path, atmospheric pressure, compressor pressure etc. Table 8.4 gives the estimated volume of compressed air leaked in NTP atmospheric conditions through apertures of various sizes. The table also gives energy wasted in the leakage. Calculations are based on compressed air pressure of 700 kPa and yearly operation of the compressor for 2000 hrs.
8.4
COMPRESSORS USED ON BLASTHOLE DRILLS
To a rotary blasthole drill, a compressor is as good as a heart to a human being. Rotary blasthole drills use compressors with discharge capacities ranging between 12.5 to 107.6 m3/min and pressures ranging between 345 to 862 kPa. In some blasthole drills, compressors are used capable of delivering air at high
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Table 8.4 Air leakage and energy wastage in air system. Aperture diameter in mm
Quantity of air lost in leaks in liters/s
Annual energy wasted in kWh
0.4 0.8 1.6 3.2 6.4 12.7
0.2 0.8 3.2 12.8 51.2 204.8
133 532 2128 8512 34040 136192
pressures – even to 3500 kPa – so that the drills are also capable of DTH hammer drilling. In the early days, rotary blasthole drills used to have piston type reciprocating compressors. By the 1960s many of the drills changed over to sliding vane type rotary compressors. From 1980 onwards preference has shifted to rotary screw compressors for the many advantages that they offer.
8.4.1
Sliding vane rotary compressors
The working of sliding vane rotary compressors is quite easy to understand. A schematic illustration of a sliding vane rotary compressor is in Figure 8.11. The working of a vane compressor can be explained from Figure 8.12. The vanes of the compressor are made from a tough polymer and are free to slide in and out from the slots in the rotor. The rotor rotates in the direction shown and the centrifugal force keeps the vanes out of the slots to the extent made possible by the steel housing in which the rotor and vanes are confined. The rotor is mounted in the housing eccentrically in such a manner that at the lower end the gap between the rotor and the housing is almost nil. Usually there are 8 vanes in 8 slots of the rotor. During rotation when the vane is at point A, it creates an enclosure of air between the rotor, housing and the previous vane that has moved to the point B. This enclosure opens in the suction port as shown and the air rushes into the enclosure. As the rotation continues both the vanes move outward and the volume of the enclosure increases. More air is sucked into it from the suction. When the vane moves from point A to point C the enclosure loses the contact with the suction port and the air in the enclosure is trapped. Further rotation forces the vane to slide inside and the volume of the enclosure keeps on reducing. The entrapped air in the enclosure is thus compressed. When the vane originally at point B reaches point D the enclosure gets opened to the discharge port and the compressed air moves into the discharge port. The volume of the enclosure keeps on reducing till it becomes virtually nil. Thus, compressed air is completely forced into the discharge port.
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Rotor
Vanes
249
Cooling Medium
Air Inlet
Air Outlet
Steel Body
Cooling Medium
Figure 8.11 Schematic of sliding vane rotary compressor.
C
Suction Port B A
Discharge Port
Figure 8.12 Working of sliding vane rotary compressor.
Some advantages of sliding vane compressor are: 1 High volumetric efficiency, up to 99.8% is retained through the life of the compressor. 2 Very compact construction, require less space.
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3 Can be driven by V belts. 4 The component alignment is only at one place and that too is not very critical. 5 Lubrication is required at three places viz. the contact of slots and vanes, vanes and housing, and the bearings of the rotor shaft. This can be accomplished by force feeding the lubricant at appropriate points. 6 Lubricant mixes with compressed air only to a negligible extent. 7 Very simple construction through usually 8 vanes and two roller bearings. Therefore fewer wearing parts. 8 Lower manufacturing cost and hence relatively low price. 9 No cold weather starting problems. 10 Low noise level. 11 Negligible pulsation of air in discharge side. 12 Low heat generation as the vanes are made from non metallic materials. The main drawbacks of the sliding vane rotary compressors are: 1
2 3
Due to the limitations of the strength of vanes made from polymer, the eccentricity of the rotor cannot be increased beyond a certain limit. Therefore, vane compressors cannot achieve a very high compression ratio, and hence it is difficult to achieve pressures beyond about 400 kPa in the first stage. It is usually not possible to change the pressure discharge characteristics. Compressed air becomes hot as it absorbs the heat generated by the friction between the vanes and the housing.
Due to these disadvantages, most of the blasthole drills that were equipped with vane compressors have now switched over to screw compressors.
8.4.2
Rotary screw compressors
The working of rotary screw compressors can be explained with a series of figures. A screw compressor consists of one male and one female rotor as shown in Figure 8.13. The male rotor has four lobes, whereas the female rotor has six grooves. The male rotor is driven in the direction as shown and the female rotor follows it in the opposite direction. When the groove of the female rotor is opened to the suction port, atmospheric air fills the groove enclosure formed by the casing of the rotors and the end flanges. As the male rotor rotates and the female rotor follows, the lobe of the male rotor intrudes into the female rotor groove at the front flange end. The male and female rotors are so machined that the gap between male and female rotors is very small. Further, it is filled with lubricant circulated through the compressor housing for the purpose. Thus, air cannot escape through this gap. As the male rotor rotates further and female rotor follows, the contact between suction port and the female rotor groove is lost and the air in the female rotor is entrapped. Further rotation causes the reduction of volume of the groove between female rotor, male rotor and the rear end of the flange as shown in Figure 8.14. Air is thus compressed further as shown by the dark gray color in the figure. Further rotation causes contact between the groove filled with compressed air and the discharge
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Atmospheric Air Female Rotor in the Female Groove
Gap Between Rotors
Front End Flange
251
Suction Port Rear End Flange
Male Rotor
Figure 8.13 Screw compressor working 1.
port at the lower side of the compressor casing as shown in Figure 8.15. The compressed air thus rushes completely into the discharge port. It is quite easy to visualize that the compression ratio depends upon the length of the male and female rotors. The length of these rotors cannot be increased beyond a certain limit that is imposed for many reasons such as difficulties in manufacture, alignment etc., but in actual practice it can be made sufficiently long to give pressures of the order of about 1500 kPa in one stage. The compressed air received in the discharge port is always mixed with oil. Rotor profiles can be machined very precisely so the need for a film of oil between the male and female rotors is eliminated. Most of the screw compressors used on rotary blasthole drills are large oil-flooded type. The proportion of oil in the air is quite excessive even for air flushing operations. For this reason and for the purpose of economy of operations the oil must be separated and retrieved. Oil separation of the compressed air is carried out in a subassembly called an oil separator and is shown in Figure 8.16. The oil-mixed air discharged by the compressor flows to the oil separator where the air is filtered to remove the oil. Usually 99.5 to 99.8% of the oil is removed from the compressed air and recirculated back to the compressor as shown in the figure. Oil that flows with compressed air to the drilling string and eventually the bit, serves the purpose of bit lubrication. When blasthole drills, earlier equipped with vane compressors, were modified with a screw compressor a significant increase in bit life was observed. Compressed air discharged from the compressor loses pressure while flowing through the piping up to the oil separator and subsequent oil filtering operation. Depending upon the design of the oil separator assembly, the pressure loss is about 60 to 100 kPa. Thus the pressure developed at the compressor is higher by this differential than the air discharged to the outlet for flow to the drill string.
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Compressed Air
Figure 8.14 Screw compressor working 2.
Suction Port at Top
Compressed Air Discharge Port at Bottom
Figure 8.15 Screw compressor working 3.
The main advantages of screw compressors are: 1 2 3
4 5 6
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The moving parts are only the two rotors, two roller bearings and two thrust bearings. Screw compressors can attain a very high compression ratio – even up to 20:1. Thus, it is possible to get high pressures up to even 1500 kPa in the first stage. Since oil is injected in the compressed air for the purpose of sealing and lubrication, it also reduces the temperature of compressed air sent into the discharge pipes. The size of the compressor is very compact. During operation the noise and vibration levels are very low. The pressure discharge characteristics can be easily varied by means of a sliding valve control placed at the bottom of the compressor.
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Radiator Oil Cooler
Minimum Pressure Orifice
Fan Motor
Scavenging Line Air Outlet
Inlet Regulating Valve
Air Inlet
Strainer
Oil Separator Strainer
Heavy Duty Inlet Air Filter Sight Glass Orifice Oil Filter
Thermostatic Flow Control Valve Screw Compressor
Oil Level Gage Air/Oil Receiver
AIR FLOW DIAGRAM
Oil Pump Main Oil Filter
Bearing Oil Filter
Air Air/Oil Oil
Figure 8.16 Oil separation in screw compressors.
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7
8 9
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In screw compressors it is also possible to reduce capacity without making any change in the discharge pressure by means of a sliding valve which causes recirculation of the air between adjacent female grooves while the air is being compressed. This decreases power requirements significantly. Maintenance of the screw compressors is very easy. It is required only to a negligible extent. Pulsation levels of compressed air are very low. The main drawbacks of screw compressors are:
1 2 3
Since the compressed air contains a significant quantity of oil, an oil separator is essential. As the temperature of oil is increased during air compression, the screw compressors require an oil cooler. Precise profiling of the male and female rotors increases manufacturing costs and so the price of screw compressors is higher than equivalent sliding vane compressors.
In the early days, screw compressors were designed for operating speeds of for 3550 with two-pole electric motors, but later their design was changed for operation at about 1800 rpm. With this they could be used on blasthole drills. Today, most of the blasthole drills have screw compressors mainly for their power-saving features. Screw compressors on very large blasthole drills operate at about 450 kPa air pressure whereas those on smaller drills operate at higher pressures ranging from 689 to even 758 kPa. Some two-stage screw compressors can also operate at either low pressure high discharge mode for rotary drilling, or high pressure low discharge mode for DTH drilling.
8.4.3
Discharge and pressure control
From the viewpoint of rotary drilling the compressor must discharge an adequate volume of compressed air at sufficiently high pressure. How much discharge and pressure is sufficient depends upon several factors related to blasthole flushing. Due to its importance this topic has been elaborated in a separate chapter. It is essential that compressed air generated by the compressor must be controlled to give the desired discharge at the desired pressure. For this purpose one of the three types of control systems viz. load/no load, modulation, or combination of load/no load and modulation, are used. 8.4.3.1
Load/no load control
This system consists of a valve at the suction of the compressor (Valve1), a pressure sensor that senses the pressure at the discharge port of the compressor and another valve at the discharge port of the oil separator (Valve2).
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When the compressor drive motor or diesel engine is started initially, Valve1 is closed and Valve2 is open. As Valve1 is closed no air is sucked into the compressor and the compressor starts without any load. Thus, no overloading or overheating of the drive motor or engine occurs. As the rotors start rotating, a very small quantity of air accumulated in the housing of the compressor does get compressed and is discharged. When the pressure sensor senses this pressure, Valve1 is opened and Valve2 is closed. The compressor then starts fully loaded. Pressure in the piping and air separator sub-assembly starts increasing. When the pressure sensor senses that pressure has reached an adequately high level and the oil is being forced to the compressor, Valve2 is also opened and Valve1 is closed. Thus, the compressor works in unloaded condition. When the operator opens the valve in the discharge pipe located beyond the oil separator Valve1 is opened again and the compressor starts working in loaded condition. 8.4.3.2
Modulation control
This system consists of a control valve (CV1) at the suction of the compressor, a regulator with pressure sensor and a control valve (CV2) at the discharge port of the oil separator. The controller senses the pressure at the discharge port of the compressor and partially closes both the control valves CV1 and CV2 as needed depending upon the pressure sensed. When the motor or diesel engine driving the compressor is started initially, CV1 is fully closed and Valve2 is open. As CV1 is fully closed no air is sucked into the compressor and the compressor starts without any load. Thus, no overloading or overheating of the drive motor or engine occurs. As the rotors start rotating, a very small quantity of air accumulated in the housing of the compressor does get compressed and is discharged. When the pressure sensor senses this pressure, the CV1 is fully opened and Valve2 is closed. The compressor then starts fully loaded. Pressure in the piping and air separator sub-assembly starts increasing. When the pressure sensor senses that pressure has reached an adequately high level and the oil is being forced to the compressor, Valve2 is also opened and CV1 is closed. Thus, the compressor works in unloaded condition. When the operator opens the valve in the discharge pipe located beyond the oil separator the pressure in the system drops down. At this stage the regulator comes into action and partially opens CV1, depending upon the pressure differential. Modulation control is very effective in maintaining a stable discharge pressure. 8.4.3.3
Load/No load and modulation control
This is a combination of the two control systems described above. When the compressed air volume required to reach the pressure level is more than about 70%, the control is as per the Load/No Load System described above, but the volume requirement is less and the control is achieved by modulation. When the compressor is working in unloaded condition it consumes about 25% to 30% of the full load power. In all-electric drills the system may also consists of
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Percent of Full Load Power
120 B
100 80
E
D
A
C G
F H
60 I
40
J
20 K 0
0
20
40 60 80 100 Percent of Capacity Load Unload Control - ABCDEF Modulation Control - AGHIJK Combination Control - ABHIJK
Figure 8.17 Comparison of control system.
an adjustable timer. If this timer senses that the compressor is working in unloaded condition for a long time the electric motor driving the compressor is automatically switched off, after closing Valve1 or CV1 as the case may be. Figure 8.17 illustrates a comparison of the energy consumption with each control system.
8.5
MEASUREMENTS OF COMPRESSED AIR
As can be seen from the earlier elaboration, the calculations in respect of compressed air are rather complex and are not very accurate. For this reason many manufacturers rely more upon measurements of flow volumes, temperatures, pressures etc. than finding theoretically calculated values of such parameters. For this purpose a measuring kit, like the one shown in Figure 8.18, is used. It consists of a unit containing a pressure gage and a temperature gage, a short cylinder with a polished inner surface, precisely shaped sharp orifices of several opening diameters, and an adapter to connect the cylinder at the bottom of the drill head or drill pipe. If pressure loss is to be measured, one more pressure gage is required to be attached in the pipeline on the deck at a location where the pipeline is connected to the compressor discharge. By fitting the orifices of different sizes one after another in the kit assembled at the bottom of the drill pipe, and allowing the compressor discharge to flow through the pipes, the reading of pressure and temperature can be taken on the gages provided in the kit. The reading should be taken at such time when the readings have stabilized on the gages. This may take several minutes. At this time the reading on the additional pressure gage installed in the beginning of the flow passage is also taken.
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Figure 8.18 Compressed air measuring kit.
By using tables provided with the kit, the discharge of compressed air through the particular nozzle can be found from the measured readings on the gage. Thus, the pressure loss for a particular flow volume through the pipeline is known from the pressure difference in the two pressure gages.
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Chapter 9
Mechanics of rock fracture under a drill bit
9.1
INTRODUCTION
It is important to understand the process of rock fracture with good conceptual clarity because one can get to know the advantages, disadvantages and limitations of the equipment, and procedures used for blasthole drilling. Rock breakage takes place when energy is applied to a rock mass. The intensity and rate at which the energy is applied to the rock mass plays an important role in the rock fracture process. In drilling as well as subsequent blasting, rock is broken and moved. The mechanics of rock fracture, however, is radically different in drilling than in blasting. In this chapter both these types of rock fracture mechanisms are elaborated to the necessary extent.
9.2
ROCK FRACTURE IN DRILLING
Energy in many different forms can be applied to a rock mass for creating a blasthole. To ensure that subsequent blasting is very effective, a blasthole must have some particular dimensional properties. As a blasthole is required to have set dimensions, energy has to be applied in a very controlled form. Three forms of energy can be applied to a rock mass in controlled manner. They are 1 2 3
Mechanical Energy Heat Energy Chemical Energy
Based on these three forms of energy application, many drilling methods as shown in Figure 9.1, are recognized. Of these three forms of energy, mechanical energy is almost exclusively used in actual drilling practice. The other two forms viz. heat and chemical energy are yet to prove their potential from an economic angle. Therefore, the methods that use mechanical energy for formation disintegration are called conventional drilling methods. Methods that use other two forms are non-conventional drilling methods.
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Drilling Methods Conventional Drilling Methods Rotary Drilling
Non Conventional Drilling Methods
Cable Tool Drilling
Rotary Percussion Drilling
Top Hammer Drilling
Down the Hole Hammer Drilling
Uncased Drilling Drill Rod Drilling
Full Face Tunneling
Water Jetting
Cased Drilling
COPROD Drilling
OD Drilling
Rotary Drilling
Large Diameter Rotary Drilling
ODEX Drilling
Augur Drilling
Calyx Drilling
Diamond Drilling
Jet Piercing
Induction Drilling
Figure 9.1 Drilling methods.
Microwave Drilling
Laser Drilling
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About five decades ago, when the metallurgy of drilling tools had not sufficiently developed, a drilling method called jet piercing, where heat energy was used for the disintegration of the formation, was adapted for drilling blastholes in extremely hard formations such as taconites. Soon thereafter great advances were made in the concepts, metallurgy and machining of drilling tools. This resulted in a huge reduction in the cost of drilling with these tools, and now the jet piercing method is totally discarded from blasthole drilling practice. Use of chemical energy for formation disintegration is still in the experimental stage. It is very unlikely that any drilling method based on heat or chemical energy application will become economically viable in next two decades or so.
9.2.1
Modes of fragmentation
Whichever may be the type of energy used, in the drilling process rock is fragmented to create a hole. The three forms of energy that have potential in being used for drilling are described here for academic purposes. 9.2.2.1
Mechanical energy
When mechanical energy is applied to a rock mass it causes various types of stresses. These stresses cause strains in the rock mass and subsequently it breaks into small pieces as the cracks propagate. Different modes of passing on the mechanical energy to the formation are practiced. These are described below. 9.2.2.1.1
Crushing
Crushing takes place when heavy and somewhat steady force is exerted on the formation through a hard drilling bit. The strain required to cause crack propagation can be adequately high only when the stress generated by the drilling tool is high. This can be achieved by having a very low contact area between the rock and the drilling tool or by exerting a very high force on the drilling tool. Figure 9.2 shows the disintegration pattern of the rock mass caused by the exertion of steady force on it. It must be noted that just along the contact between bit and
Increased Feed Force
Feed Force
Indenter
Zone of Very Fine Cuttings
Zone of Fine Cuttings Zone of Coarse Cuttings
Zone of Coarse Cuttings
Figure 9.2 Crushing of rock under an indenter.
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the formation and within some portion beneath it, the stresses developed are compressive, but along the fracture lines the stresses are predominantly shear stresses and the fracture takes place due to shear stresses. Crushing is the basis of material disintegration in rotary drilling using tricone bits. 9.2.2.1.2
Impact crushing
When a hard material piece impinges on the rock mass, the impact force is increasing for a very short duration and then it reduces suddenly. Thereafter the contact between the tool and rock mass is lost. As the impact force is increasing, the particles of the rock mass are compressed and a compression wave is generated when the particles near the drilling tool compress the particles a little farther from the drilling tool. As the force exerted by the drilling tool suddenly reduces, the particles in contact with the drilling tool try to regain their original position by moving back. Thus, a tension wave is also generated. Rocks by nature are made up of such materials that their tensile strength is very low. Since the tensile forces in the vicinity of the drilling tool are high, cracks are generated in the vicinity of the drilling tool. Further impacts of the drilling tool propagate the cracks and eventually a piece of rock is separated from the rock mass. Impact crushing is the basis of material disintegration in percussion drilling. 9.2.2.1.3
Scratching
Scratching of rock mass by a drilling tool is schematically illustrated in Figure 9.3. For scratching to take place, a vertical force that tries to penetrate the drilling tool into the rock mass and a horizontal force that tries to move the drilling tool along the surface of the rock mass, are necessary. Cracks are generated and propagated by the horizontal force along a shear surface as shown in the figure. The vertical force is necessary to initially cause penetration of the drilling tool into the rock mass and then keep the drilling tool in the penetrated position. If the vertical force is removed the drilling tool slides up along the fractured surface and continues its movement on the top of the rock surface without generating cracks and scratches.
Vertical Force
Separated Material Horizontal Force
Surface Where Potential Cracks Exist
Figure 9.3 Scratching of rock by a plow tool.
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Scratching action for rock disintegration takes place in rotary drag bit drilling or diamond bit drilling. More details of rock fragmentation by application of mechanical energy are given later in this chapter. 9.2.2.2
Heat energy
When heat energy is transferred to the rock mass, disintegration takes place either by spalling or fusion and vaporization. 9.2.2.2.1
Spalling
A rock mass consists of varied constituent minerals. These minerals have different coefficients of linear expansion. When formations having such minerals are rapidly heated, due to different levels of thermal expansion of the mineral particles, high stresses develop in the rock mass and result in disintegration. Most of the rock-forming minerals are poor conductors of heat. If a heat source with high potential such as a high temperature flame is brought very near the rock surface, the rock particles in a layer near to the flame get heated very quickly to high temperature. Since the heat is not effectively transmitted by these hot particles to the particles situated a little deeper in the rock mass, there is a great disparity between the thermal expansion of the particles in these two (or more) layers. This gives rise to flakes. The phenomenon of flaking caused by heat is called spalling. If a rock mass consists of minerals that are better conductors of heat and have uniform coefficient of thermal expansion, it does not spall easily. Limestone is one such rock that does not spall easily. Spalling is the basis of material disintegration in jet piercing. 9.2.2.2.2
Fusion and vaporization
If a rock mass is heated to very high temperature, it melts. Further heating causes vaporization of the rock mass. In vapor form the formation can be displaced easily by the gases used for the combustion. As the vaporized minerals move away from the flames with the exhaust gas flow, the cooler surroundings turn the mineral back to solid form and the cuttings formed in the solidification come out of the hole with the exhaust gas flow. This type of heat energy transfer to formation for the purpose of drilling is as yet under laboratory testing. 9.2.2.3
Chemical energy
Some minerals that form rock easily react with certain chemicals. A very common example of this is the reaction of marble with carbonic acid. As marble is mostly composed of calcium carbonate, i.e. CaCO3, it easily reacts with carbonic acid i.e. H2CO3. The chemical equation of the reaction is CaCO3 + H2CO3 = Ca(HCO3)2
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The minerals that result from such a reaction are in loosely bonded crystal form and can be easily removed from the place of formation. Many such reactions of different chemicals with a variety of minerals are known. It is, however, not very likely that chemical energy will be used for primary rock fracture in drilling for the following two main reasons. 1 2
The chemical reactions are invariably slow The composition of rock mass varies and even a small layer of mineral that does not react with the chemical can halt further progress.
The subject of this book being rotary blasthole drilling, all further discussion about rock fracture in this chapter is in respect of crushing and scratching as these mechanisms fracture rock in rotary drilling.
9.3
BASIC THEORY OF SOLID FRACTURE
When external or internal forces act on a solid body, stresses are generated at virtually every imaginable surface within the solid body. There are three basic types of stresses viz. compressive stress, tensile stress and shear stress. Figure 9.4 shows the three types of stresses. If a surface under stress is considered, compressive stress tries to push the atoms in the solid body towards each other in the direction perpendicular to the surface. Tensile stress tries to pull the atoms in the solid body away from each other in the direction perpendicular to the surface. Shear stress tries to move the atoms in the solid body away from each other along the surface. Whenever stresses are generated in a body it changes its shape. This is called strain. Solid bodies can be perfectly elastic, perfectly plastic or elasto-plastic i.e. combination of the two types. In a perfectly elastic body strain goes on increasing with increasing stress but when the external or internal forces that cause the stress are
External Force External Force
External Force
External Force Compressive Stress
Tensile Stress
Shear Stress
Figure 9.4 Three basic types of stresses.
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reduced to zero magnitude the strain also proportionally reduces to zero magnitude. In perfectly plastic bodies the strain goes on increasing so long as the body is under stress. Later when the external or internal forces that cause strain are reduced to zero the increase in strain stops but the magnitude of strain does not become zero as in elastic bodies. In other words the body remains in deformed shape. In reality, all materials are elasto-plastic. Atoms in a solid body are bound together by bonding forces. When the atoms of a material move away from each other and the bond between adjacent atoms breaks down, fracture is said to have taken place. Fracture generates cracks. When cracks propagate and reach the surface of the body, the body disintegrates into separate small pieces. Since the bonding forces between atoms must be broken for fracture to take place, in any fracture it is essential that the atoms move away from each other. This happens only in case of tensile strain or shear strain. Thus, a solid body can fail by tensile or shear strain but not by compressive strain where the atoms move near each other and bonds do not break. Experiments carried out by Bridgman and Crossland have proved that if a solid body is subjected to external compressive forces perpendicular to it in all directions, it does not disintegrate even when the forces are extremely high. Some properties of the material, however, change dramatically. When a sample of material subjected to compressive load fails, the failure occurs through the tensile and shear stresses generated in the sample as a consequence of the compressive load. Many theories about fracture of solids have been suggested so far. For brittle materials like rocks, the Mohr-Coulomb Theory for failure in shear and the Griffith Theory for failure in tension are very widely accepted. A complete description of both the theories is beyond the scope of this book. The following are brief summaries of the two failure theories.
9.3.1
Mohr-Coulomb theory of shear failure
While studying the magnitude of pressure exerted on a retaining wall by the soil behind the wall, Coulomb proposed in 1773 that a shear stress is always associated with a compressive stress. The equation for the shear stress proposed by him is S = C + σ tan φ where S = Shear stress C = Cohesion between the soil particles σ = Normal (compressive) stress φ = Angle of internal friction A graphical representation of the above equation is shown in Figure 9.5. As Coulomb was dealing with earth pressures he concluded that failure will occur when the shear stress exceeds shear strength. Thus the line AB in Figure 9.5 is the envelope that determines failure criteria.
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B S σ tan φ A C
φ σ
Figure 9.5 Mohr Coulomb failure criteria.
Compressive Stress mechanically applied on the two circular surface of a cylindrical rock sample.
σr
Compressive Pressure i.e. stress hydraulically applied from all the sides on the curved surface of a cylindrical rock sample
Figure 9.6 Triaxial compression test.
Mohr considered the concept further and in 1882 concluded that failure within a solid body was not caused by normal compressive stress alone reaching a certain maximum or yield point, or by shear stress alone reaching a maximum, but a combination of both shear and normal stress. Even though the failure is essentially caused by shear stress, the critical shear stress is governed by the normal stress acting on the potential surface of failure. The critical combination of shear and normal stress, when plotted on a graph will give a line known as Mohr’s envelope of failure. For most rocks the line is a curve. There no equation for plotting the line. The line is actually drawn on the basis of results on the rock samples obtained in a triaxial compression test. Forces applied on the cylindrical rock sample in a triaxial compression test are shown in Figure 9.6. During the test the hydraulically applied pressure is kept constant but the stress due to mechanical force applied on the end of the cylindrical rock specimen is slowly increased till the failure takes place. Typical results of three tests carried out on three cylindrical samples of the same rock with different hydraulic pressures are given in Table 9.1. These results can be plotted to give Mohr’s envelope as shown in Figure 9.7.
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Table 9.1 Results of triaxial compression test. Mechanically applied comp. stress σc in MPa
Hydraulic pressure σr i.e. comp. stress in MPa
Difference (σc − σr) in MPa
500 900 1400
100 300 600
400 600 800
Mohr’s Envelope
The diameters of the circles are equal to the difference between the stresses. The centers of the circles are at a distance equal to the sum of stresses divided by 2.
300 600 1000
Figure 9.7 Mohr’s envelope plotted from test results.
The Mohr Coulomb theory was originally developed for soils, which by nature have very weak bonding forces amongst the particles. It can, therefore, be applied with reasonably good accuracy for slope failures in soils and rocks. The theory, however, does not go to the desired depth in the actual mechanism of fracture of the rock.
9.3.2
Griffith theory of tensile failure
This theory was proposed by Griffith in 1920. It was developed for the actual mechanism of fracture in glass which is a highly brittle material. Griffith considered that brittle materials do have very fine randomly located and oriented cracks and discontinuities within them. For the material to break these cracks must propagate i.e. extend in their length till they join each other and create a failure surface that extends up the outer surface of the material body. The parts of the body on the two sides of failure surface must also separate. The assumption made by Griffith is very near to reality because if we consider a perfect material having a lattice of atoms in the atomic structure as shown in Figure 9.8, the formula for theoretical strength of the material is σt = (Eγ/b)0.5
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σ
σ b
Figure 9.8 Atoms and bonds in an atomic lattice of material.
where σt E γ b
= = = =
Theoretical Strength of the Material Modulus of Elasticity of the Material Actual Inter-atomic Separation Inter-atomic Distance at Stable State
For quite a few materials the inter-atomic distance γ required for breaking the bonds is equal to Eb/40, hence the theoretical strength of the material works out to E/6.3245. Since E is very high the theoretical strength also works out to an unrealistically high value. Thus, the actual low strength of the material can only be explained by presuming cracks in the material. Griffith’s theory has been extended and improved by many, but the basic assumption of the existence of cracks and imperfections in materials has been recognized in each of such extensions. Therefore before going ahead with Griffith’s theory one must understand the characteristics of a crack. If we presume a circular hole in a perfect material that is subjected to tensile stress σ as shown in Figure 9.9, it can be shown from the theory of elasticity that a stress −σ (negative sign shows that actually the stress is compressive) exists at the top and bottom of the hole, and a tensile stress of magnitude 3σ exists at the sides of the hole. The stress concentration arises from the lack of load-bearing capacity of the hole. The magnitude of the stress concentration is determined by the geometry of the hole but is irrespective of the size of the hole. If the hole is of elliptical shape with dimensions b and c as shown in Figure 9.10 then the stress concentration factor x is given by an equation, x = (1 + 2c/b) and the tensile stress at the narrow ends of the elliptical crack will be σe = σ(1 + 2c/b) Griffith considered the energy balance during propagation of cracks. The left side illustration in Figure 9.11 shows an elastic body that contains a crack of length 2c.
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σ
−σ 3σ
Figure 9.9 Concentration of stress around a circular hole.
σ
2b xσ 2c
Figure 9.10 Concentration of stress around an elliptical hole.
σ
Us = 2Cy
δc U 2C U
−Ue
Figure 9.11 Energy balance in crack propagation.
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The internal strain energy of the body is Ue. The body is loaded by a force that causes tensile stress. If the crack extends by a distance δc due to the action of external force, then work W will be done by the external force. A new surface will be generated. Energy spent in generating the new surface will be Us. The total energy of the system U can be equated as U = (Ue–W) + Us The term (Ue–W) in the above equation is often called mechanical energy. It can be shown that mechanical energy Um is related as Um = πc2σ2/E where c = Crack length σ = External force E = Young’s Modulus in the crack plane Similarly, as two surfaces will be formed, the surface energy will be Us = 4cγ If the cohesion between the incremental extension surfaces δc did not exist, the crack would grow outward without much work being done. Thus the mechanical energy would reduce. But at the same time the surface energy would increase. For a condition of equilibrium dU/dc must be 0. With this condition we get the stress σf at the time of failure, 0 = −πc2σf2/E + 4cγ σf = (2Eγ/πc)0.5
or
This is considered as the Griffith criteria of failure of a brittle body under tensile stress. The most important extensions to Griffith’s Theory are by Obrlemoff (1930), Wastergaard (1939), Irwin and Orowan (1948) and Rice and Cherepanov (1965).
9.4
FRACTURE OF ROCK IN DRILLING
Rotary blasthole drilling is carried out either by tricone bits or on fewer occasions by drag bits. In either cases the cutting edge of the teeth of the bit must penetrate the rock mass. A clear concept of penetration of teeth in rock is therefore essential.
9.4.1
Penetration of an indenter
When an indenter is forced into rock, stresses are generated in the rock. Even if the contact surface between the rock and the indenter is perfectly plane, the stress distribution in the rock is usually uneven. The nature of stress distribution depends upon the rigidity of the indenter. For a perfectly rigid indenter the stress distribution
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is shown in Figure 9.12. The isobar curves show the lines (or surface sections in case of three dimensions) where stresses generated are equal. As can be seen, the maximum stress is always generated near the edge of the indenter. The distribution of pressure, called contact pressure, under a perfectly rigid indenter is shown in Figure 9.13. If the rock mass has infinite strength the distribution of pressure will be as shown by the dotted line.
Vertical Force Indenter with Radius r 0
2r
r
r 1
0.7 0.55 0.5 0.4
r
2r
0.6
0.3 2r 0.2 3r 0.1 4r
Figure 9.12 Isobars created by an indenter.
Vertical Force Rigid Rock Mass
Pult
Practical Rock Mass
Ideally Rigid Formation
Figure 9.13 Contact pressure below an indenter.
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In actual practice the rock has definite ultimate strength and the maximum contact pressure is as per the continuous line shown. When an indenter is forced into rock, in the initial stage the point contacts between the indenter and the rock disintegrate the rock surrounding the points. The displaced rock particles act as a very thin distribution layer and force further disintegration of the rock. Then the particles below the indenter edge disintegrate. As the particles get loose they are unable to resist the force of penetration and the particles towards the centerline of the indenter take the load and disintegrate. Ultimately a small cone of fragmented material is formed below the indenter as shown in Figure 9.14. An indenter is not able to generate cracks in rock. It creates more fine particles of the rock mass. That is the reason why a wedge shape indenter is better from the viewpoint of formation fracture.
9.4.2
Penetration of a wedge
Considerations about penetration of a wedge in the rock are more important than that of an indenter because in actual drilling practice all the teeth of a tricone bit or even the cutting edge of a drag bit have a wedge shape. This is because a wedge has a very small contact area and more importantly an ability of exerting side pressures on the rock. With this, a wedge is capable of forming larger chips on the side and is far more effective in disintegrating the rock as compared to an indenter with a large contact area and inability of generating side pressure on the rock. When a wedge penetrates rock, in the very initial stage the contact area of the edge of a wedge disintegrates a small area underneath and a very small bulb of disintegrated rock particles is formed. This is shown in Figure 9.15 on the left side sketch of the wedge. As the wedge is forced further into the rock the sides of the wedge break the rock at their contacts. Owing to the friction between the wedge surface and the particles formed in the process, the particles have a natural tendency of moving towards the tip. The pressure inside the small bulb at the tip increases. This causes the particles in the bulb to break rock on the surroundings, and the size of the bulb increases. As the pressure in the surroundings is increased further cracks are formed at the discontinuities. This is shown by the middle sketch of the wedge in Figure 9.15. Further penetration of the wedge results in higher growth of the cracks near the surface because the rock mass on their upper side is unable to offer necessary resistance
Figure 9.14 Cone of fractured material below an indenter.
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Figure 9.15 Process of chip formation by a wedge.
to the crack growth and finally chips are formed, as shown on the right side sketch of the wedge in Figure 9.15. A wedge generates more coarse chips of the rock mass than an Indenter. When a wedge is forced to penetrate rock at slow constant speed, the resistance of the rock to penetration varies. The resistance keeps on increasing until fracture takes place; immediately after fracture, chips are formed and the resistance decreases. Upon separation of the chips the resistance starts increasing once again and keeps on increasing till the fracture takes place once again and another new set of chips is formed. This is clearly visible in the plot of resistance to wedge penetration shown in Figure 9.16. The curves presented in the figure are for wedges with different wedge angles. It can be observed from the figure that in most of the cases the depth of wedge penetration is inversely proportional to the wedge angle. How big is the volume of crushed zone and how much is the depth of cracks, depends upon the engineering properties of the rock being penetrated. In brittle rocks that have high hardness and low toughness, the volume of the crushed zone is less but the length of cracks is longer. For such rocks very high feed force is needed to initiate fracture because of their high compressive and shear strength. When a wedge penetrates rocks like basalt that have higher toughness but comparatively lower mineral hardness, the volume of the crushed zone is slightly higher but the length of cracks is somewhat shorter. Very high feed force is needed for chip formation in such rocks in view of their high toughness that gives limited extent of crack propagation. Many rocks e.g. limestone, sandstone etc., have a plastic nature. They have much lesser compressive strength. Such rocks form a large crushed zone and comparatively much shorter crack lengths. A careful observation of the experimental feed force vs. penetration curve for a wedge shown in Figure 9.16 gives clear indication that chips are formed in succession as the wedge penetrates more and more into the rock mass. The magnitude of the feed force required for formation of successive chips increases more and more. Several researchers have tried to tackle the problem analytically and have arrived at a relationship between different geometrical aspects of the wedge with depth of
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Feed Force in Thousand Pounds
274
14 13 12 11 10 9 8 7 6 5 4 3 2 1 0
Decrease to resistance shows that chips are being formed under the wedge 120° Wedge Angle Increase in resistance shows that chips are not being formed under the wedge
105°
90° 75° 60°
0
0.01 0.02 0.03 Wedge Penetration in Inches
0.04
Figure 9.16 Static feed force vs. penetration of a wedge.
wedge penetration based on some rock properties. Finite element analysis has also been introduced in establishing the trends of feed force vs penetration relationship. Peeping into such mathematical treatment of the wedge penetration has been avoided here because the solutions obtained have not yet reached a stage where the calculations give better results than the experimental procedures or empirical formulas used for predicting penetration rates in drilling.
9.4.3
Formation fracture below a drill bit
Whether a drag bit or a tricone bit is used, fracture of the formation takes place by a combination of the shear and tensile stresses. As shown in Figure 9.3, a tooth of a drag bit must initially penetrate the rock through the action of vertical force and later cause shear-cum-tensile failure by the horizontal force. In the case of drag bits the penetration of the tooth causes very little formation fracture. It is the horizontal force that mainly contributes in the formation fracture by shear. Diamond bits, though not much used in blasthole drilling, are a type of drag bit. In the case of old style diamond bits where the diamonds jut outward, failure surfaces are almost wholly formed by shear stress as shown in Figure 9.17 because the horizontal force on the diamond tends to push the formation within it. In the case of PDC bits the diamond compact is placed somewhat like a plow, as shown in Figure 9.18. Since in this case the horizontal force tries to move the formation outward, the major cause behind generation of failure surface is tensile stress. As most of the rocks are very weak in tension, PDC bits give better penetration rates than the old style diamond bits under the same operating conditions.
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Bit Matrix
Embedded Diamond Formation
Second Chip
First Chip
Figure 9.17 Formation fracture by old style diamond bit.
Bit Matrix
PDC i.e. Polycrystalline Diamond Compact
Formation
First Chip Second Chip
Figure 9.18 Formation fracture by PDC bit.
In a tricone bit the cones of the bit rotate around their respective axes as the bit rotates around the axis of the drill string. The teeth on the cone come in contact with the rock mass and later move away from it. Therefore the load exerted on a tooth increases initially and then decreases. Of the many stages in which this happens, five stages are shown in Figure 9.19. As shown in Figure 9.19a, a tooth of the bit touches and penetrates the rock surface at one stage with low feed force because most of the feed force is borne by the adjacent tooth that is in the vertical direction. The volume of crushed zone created by the tooth is very small. Due to rotation of the cone, the feed force on the tooth increases while the feed force on the earlier adjacent tooth is being released. As a result of this, penetration of the tooth increases and the volume of crushed zone also increases initially. Later shear cracks are formed in the rock as shown in Figure 9.19b. At this stage the length of the cracks in the direction of feed force, which is still slightly off the vertical, is highest. As the cone rotates further, due to lateral shift of the tooth, which in this case is presumed to be towards the right hand side of the picture, the cracks in the right hand direction propagate to a considerable distance. The cracks in the left hand direction do not propagate, or propagate to a very short distance. This can be seen in Figure 9.19c. Stresses generated in the rock due to further rotation create chips as shown in Figure 9.19d. Laterally loosened fine particles in the crushed zone are removed at this stage by flushing air. Finally due to lateral shift and rotation of the tooth the chips are separated from the rock mass and move away with flushing air as shown in Figure 9.19e.
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Feed Force Wedge
Feed Force Wedge
Feed Force Wedge Tooth Height
Crushed Zone
Rock
Rock
Rock Initial Cracks a
Tooth Penetration b
Feed Force Wedge
Propogating Cracks
c
Wedge
Rock
Rock Chip
Separated Chip d
e
Figure 9.19 Formation fracture underneath the tooth of a tricone bit.
Due to the shifting position of the tooth and consequent rock fragmentation, as explained above and shown in Figure 9.19, very high torque is needed for rotating tricone bits. Tricone bits meant for drilling in very hard formation have lesser offset. Such offset imparts only limited shift to the tooth. This is done so because, if the offset is kept large, rather than rock fracture, the teeth of the bit tend to break or wear out rapidly. For this reason the shape of teeth meant for hard rock drilling is with a larger wedge angle in the case of steel tooth bits, or nearly hemispherical in the case of TC inserted bits. In other words this means that for bits meant to drill in very hard formations, the height of the teeth is less, and for those meant to drill in soft formations the height of insert is greater. Technical common sense tells us that as long as the penetration of the tooth in the formation does not exceed the height of the tooth, the following trends are likely to occur. 1
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Since the teeth of the bit will penetrate in the formation to a larger depth, the penetration rate is likely to increase with increasing feed force.
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277
If the speed of rotation of the bit increases, the teeth of the bit will penetrate into the formation more frequently. Hence the penetration rate is likely to increase with increasing rotary speed.
In actual drilling with tricone bits, there are four factors that largely affect the process of drilling. Mathematically it can be expressed as P = f(W, N, C, Q) where P W N C Q
= = = = =
Penetration rate Feed force on the bit Rotary speed of the bit Height of tooth of the bit Compressed air flow volume used.
Of the above factors the volume of compressed air i.e. Q does not take part in the formation fracturing process but is very important from the viewpoint of removing the cuttings from the bottom of the hole. Therefore it is elaborated in a separate chapter. George Shiveley took into consideration a large spectrum of blasthole drilling data obtained in field practice. Drilling was carried out by tricone drill bits in limestone quarries. In all the cases the tooth penetration in the formation was less than the tooth height. Some plots and the conclusions drawn by him are described below. A plot shown in Figure 9.20 clearly indicates that for a constant rotary speed the penetration rate varies as per the feed force exerted on the tricone bit. Mathematically it can be stated that, P∝W The plot in Figure 9.21 is for penetration rates at 100 and 200 rpm rotary speed for different feed forces on the bit. The data used to draw this plot were from a coal mine. This plot also reconfirms that rotary speed remaining the same, the penetration rate varies as per the feed force exerted on the bit. More importantly the plot also indicates that by doubling the rotary speed (i.e. from 100 rpm to 200 rpm) the penetration rates also doubles. Yet another plot is shown in Figure 9.22. It represents data from a copper mine. Here it can be seen that till the penetration rate was about 2.4 ft/min (which corresponded with feed force of 27000 lb) the curve was a straight line, but when higher and higher feed force was exerted on the drill bit while keeping the same rotary speed, the penetration rate became successively lower than the one predicted by the straight line. Here, it may be noted that the bit used for drilling had a tooth height of 0.37 inches and was rotated at 79 rpm. This means that while drilling at the penetration rate of 2.4 ft/min the penetration of the tooth was equal to the tooth height.
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5 N = 100 RPM N = 155 RPM
Penetration Rate ft/min
4 N = 220 RPM N = 285 RPM 3
2
1
0 0
10
20
30
Weight on the Bit in lb/1000 Figure 9.20 Plot of penetration rate vs. feed force on the bit at various rotary speeds.
5 N = 100 RPM N = 200 RPM
Penetration Rate ft/min
4
3
2
1
0 0
10
20
30
Weight on the Bit in lb/1000 Figure 9.21 Plot of penetration rate vs. feed force on the bit at 100 and 200 RPM rotary speed.
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5 N = 79 RPM
Penetration Rate ft/min
4
3
2
1
0 0
20
40
60
80
Weight on the Bit in lb/1000 Figure 9.22 Plot of penetration rate vs. feed force on the bit at rotary speeds of 79 RPM.
A tricone bit is said to have bogged down when the teeth of the bit penetrate in the ground to the fullest extent and additional feed force on the bit does not increase the bit penetration in the formation. Theoretically, in certain cases, such increase in feed force on the bit can actually reduce the penetration rate because the process of formation fracturing gets hampered. Insufficient flushing can occur even before bogging down of a bit. In such cases the cuttings formed are not removed from the hole bottom and secondary crushing is required before they are lifted off the hole bottom. The formation fragmenting action of a tricone bit can be divided into three different phases. If the weight on the bit is very low the penetration of teeth is very small, and whatever fragmentation takes place is without forming cracks in the formation as shown in Figure 9.19a. Such fragmentation takes place mainly due to the abrasion between the teeth and the formation. At higher feed force cracks are formed, as shown in Figure 9.19b and 9.19c. With successive crack formation the material gets removed but without any lateral displacement. Such fragmentation is called fatigue fragmentation. When the bit has a large cone offset and the load applied on the bit is significantly high the formation fracturing takes place due to the laterally applied loads as illustrated by Figure 9.19d and 9.19e. This type of fragmentation is called scraping. While using appropriate tricone bits in drilling soft, medium-hard and hard formations all the three types of fragmentation modes are present but the contributory proportion changes as shown in Figure 9.23. In soft formations much of the
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Increasing Penetration Rate
280
Soft Formations
Medium Hard Formations
Hard Formations
Abrasive Fatigue Scraping Bit Bogging
Increasing Feed Force on the Bit Figure 9.23 Contributory proportions of different modes for formation fracture under tricone bits.
fragmentation is due to scraping action but as the formations drilled are harder the contribution of abrasive and fatigue type fragmentation becomes significantly higher. In tricone bits the limitations on feed force on the bit are mostly due to the bearings in the bit. With excessive feed force the balls or rollers make a dent in the cone or bit leg surface with which they are in contact. Such dents result in vibrations that accelerate the process of dent-making and the bit fails prematurely. Small diameter tricone bits have less space for accommodating the bearings and therefore, the bearings of small diameter bits are proportionally smaller than those of large diameter bits. For this reason the maximum recommended weights on small diameter bits are also less than those for large diameter bits.
9.4.4
Specific energy
As a matter of fact, drilling is an excavation process. The difference between excavation, as it is normally visualized, and drilling is that drilling takes place in a particular cross sectional area in the rock to create a hole. The specific energy of rocks can be defined as the work done in the excavation of a unit volume of rock. Naturally, specific energy is dependent upon the properties of the rock to a very great extent. In rotary blasthole drilling the excavation is carried out within the boundary of the blasthole and for this purpose energy in the form of thrust and rotation of the drill bit is applied by the blasthole drill. An equation that correlates various
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parameters of energy transfer through a drill bit in rotary drilling was set forth by Teale as follows: S = (F/A) + (2 * π/A) * (NT/P) where S F A N T P
= = = = = =
Specific energy Thrust applied on the drill bit Area of blasthole cross section Rotary speed Torque applied to the drill bit Penetration rate
As can be seen the equation has two components. The first component viz. (F/A) does not contribute to the ‘excavation’ process in a significant manner. This is because the energy used for the penetration of the teeth of the bit into the formation is spent in crushing the rock underneath the teeth and the formation of cracks. The second component viz. (2 * π/A) * (NT/P), however, causes formation fracture to a very large extent because it loosens and displaces the fragments made by the cracks. The cross sectional area of the blasthole is almost constant as it is being drilled hence the first term, (2 * π/A), is a constant. In the initial stage of drilling when the thrust on the bit is low, hardly any penetration takes place. In such circumstances the specific energy is very high. But when thrust is increased the torque required for rotating the drill bit also increases. This increases the penetration rate. Thus, the specific energy decreases. The minimum value of specific energy can be considered as an indicator of drillability. However, such correlation has to be used very cautiously.
9.5
DRILLABILITY OF ROCKS
Drillability of rock means the real or projected rate of penetration in a given rock by a particular type of drilling bit. Sometimes it is considered as a property of rock but actually it depends not only upon the type of rock and its characteristics, but also upon many other factors. Factors that affect drillability of rock are of four categories. Factors Related to the Rock Factors Related to the Blasthole Drill Factors Related to the Drill Bit Factors Related to the Rock Mass From the above description of the process of rock fracture under a drill bit, it can be easily surmised that the fracture process is very complicated. It is possible to establish a qualitative relation of drillability with many factors as has been done in Table 9.2. However, as the factors are interrelated with each other in a complex manner, it is very
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Table 9.2 Correlation of drillability with factors in drilling. Category
Factor
Dependence of drillability with factor
Rock Factors
UCS Tensile Strength Density Abrasiveness Toughness Shear Strength Particle Cohesion
Decreases with increasing UCS Decreases with increasing Tensile Strength Decreases with increasing Density Decreases with increasing Abrasiveness Decreases with increasing Toughness Decreases with increasing Shear Strength Decreases with increasing Particle Cohesion
Machine Factors
Feed Force Rotary Speed Torque Vibrations Air Flow Air Pressure
Increases with increasing Feed Force Increases with increasing Rotary Speed Increases with increasing Torque Decreases with increasing Vibrations Increases with increasing Air Flow Increases with increasing Air Pressure
Bit Factors
Teeth Height Teeth Apex Angle Number of Teeth Hardness Abrasiveness Diameter
Increases with increasing Tooth Height Decreases with increasing Tooth Apex Angle Increases with increasing Number of Teeth Increases with increasing Hardness Increases with increasing Abrasiveness Decreases with increasing Diameter
Rock Mass Factors
Joint Spacing Joint Orientation Homo Water Contents
Decreases with increasing Joint Spacing Difficult to describe qualitatively Difficult to describe qualitatively Decreases with increasing Water Contents
difficult to develop a mathematical formula that can take into account all the factors mentioned.
9.6
ESTIMATION OF PENETRATION RATE
Many terms such as penetration rate, net penetration rate, gross penetration rate, drilling rate, drill performance etc. are used in connection with the progress of blasthole drilling. Since the terminology is not standardized, often these terms are intermixed. Here in this book such terms are used with specific meanings specified in Table 9.3. Many researchers have attempted to correlate penetration rate with one or more easily measurable properties of rock. Some of these are elaborated in following sections.
9.6.1
Protodyakonov approach
One of the earliest attempts of correlating penetration rate with rock properties was by Protodyakonov, who in 1926 proposed a toughness index of rock based on a simple test.
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Table 9.3 Definitions of penetration rate category factor. Category
Factor
Penetration Rate
Measured in m/min. Calculated by dividing the length of the blasthole drilled by the time interval during which the formation fracturing activity through application of feed force and bit rotation are continuous.
Instantaneous Penetration Rate
Measured in m/min. Calculated in the same manner as penetration rate mentioned above but over a very short interval of time.
Net Penetration Rate
Same as penetration rate mentioned above.
Gross Penetration Rate
Measured in m/min. Calculated by dividing the length of complete blasthole drilled by time taken for all the activities required for reaching completion of drilling the hole. Here time spent for moving the drill to the location of the hole and positioning it is not included but time spent in adding and removing the drill rods is included. Drilling Rate Measured in m/h. Calculated by dividing the length of complete blasthole drilled by time taken for all the activities required for reaching completion of drilling the hole. Here time spent for moving the drill to the location of the hole and positioning as well as time spent in adding and removing the drill rods is included.
Drill Performance
Measured in m/shift. Calculated by dividing the total length of all the blastholes drilled in one shift by the duration of shift i.e. 8 hours. It includes time taken for all the activities required for drilling the blastholes during the shift but does not include pauses for non availability of the power etc. In this book this term is used as a cost effectiveness indicator as stated in the beginning of this chapter.
In actual test, samples of the rock measuring about 15 to 20 mm were kept in the pan of the test apparatus shown in Figure 9.24, so that they filled the pan up to a height of 40 mm. A 2.4 kg weight, was dropped on these rock specimens by removing the retaining pin so that the drop height was 600 mm. After 10 drops (or more if required) of the weight the sample was removed and sieved through a wire mesh having square openings of size 0.5 mm. Crushed material passing through sieve was collected and poured in the measuring apparatus shown in Figure 9.24. After lightly tapping the cylinder, a piston was inserted in the cylinder to measure the height of column of crushed material in the measuring cylinder. The Protodyakonov number, that represented the dynamic strength of the rock, was found by using an empirical formula as under, f = 20(n/h) where n was the number of drops of the weight and h was the height of crushed rock, measured in mm, in the cylinder. It was realized later that Protodyakonov number correlated with properties of rock as f = 1.887 * (σc2/(2 * E))
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All Dimensions in mm.
Figure 9.24 Apparatus for measuring protodyakonov index.
where σc = UCS in kg/cm2 E = Modulus of elasticity in kg/cm2 Based on the tests described above Protodyakonov proposed a classification of rocks as shown in Table 9.4. The indices presented give indications of likely penetration rates only on relative basis rather than absolute basis. Therefore, they are not of much use in actual estimation.
9.6.2
Paone and Bruce approach
An equation set forth by Paone and Bruce for penetration rates in diamond core drilling is δ = 2 * π * (T – μ * F * r)/(S * A – F) where δ T μ F r S A
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= = = = = = =
Depth of penetration per revolution of the bit in m Torque applied to drilling bit in Nm Coefficient of friction between bit and rock Feed force applied on the bit in N Mean radius of the drilling kerf in m Resistance of rock to drilling Cross sectional area under bit in m2
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Table 9.4 Rock classification based on protodyakonov indices. Category
Hardness level
I
Hardest
The hardest, toughest and most dense quartzite and basalt, other exceptionally hard rocks
20
II
Very hard
Very hard granitic rocks, quartz porphyry, very hard granite, siliceous schist, weaker quartzite, hardest sandstone and limestone
15
III
Hard
Granite (dense) and granitic rocks, very hard sandstone and limestone, quartz bearing veins, hard conglomerate, very hard iron ore
10
IIIa
Hard
Limestone (hard), weaker granites, hard sandstone, hard marble, dolomite and pyrites
8
Description of rock
Rock hardness f
IV
Rather hard
Ordinary sandstone, iron ore
6
IVa
Rather hard
Sandy shale, schist, schistose sandstone
5
V
Moderate
Hard argillaceous shale, non-hard sandstone and limestone, soft conglomerates
4
Va
Moderate
Various types of shale (non-hard), dense marl
3
VI
Rather soft
Soft shale, very soft limestone, chalk, rock-salt, gypsum, frozen soil, anthracite, ordinary marl, weathered sandstone, cemented shingle and gravel, rocky soil
2
Vla
Rather soft
Detritus soil, weathered shales, compressed shingles and rock debris, hard bituminous coal, hardened clay
1.5
VII
Soft
Clay (dense), soft bituminous coal, hard alluvium
1.0
Vlla
Soft
Soft sandy clay, loess, gravel
0.8
VIII
Earthy
Vegetable earth, peat, soft loam, damp sand
0.6
XI
Dry Substances
Sand, talus, soft gravel, piled up earth, extracted coal
0.5
X
Flowing
Shifting sands, swampy soil, saturated loess and other saturated soils
0.3
This equation applies to diamond core bits that have small and uniform size diamonds studded in a sintered matrix. While calculating the depth of penetration per revolution by using above equation, all the other parameters, except S, are easily known from the bit geometry and machine characteristics. Figure 9.25 shows the experimental and theoretical results for penetration rates for diamond core drilling. In theoretical calculations the coefficient of friction for the rock, i.e. μ, was presumed to be 0.4 and resistance to rock drilling i.e. S, was presumed to be equal to the compressive strength of rock. In tests the three quantities viz. penetration rate, coefficient of friction and compressive strength were measured. Figure 9.25 indicates very good correlation between experimental and theoretical results.
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0.060
Penetration Per Revolution in inches
0.055 0.050 0.045
Experimental Curve Drawn From Test Results
0.040 0.035 0.030 0.025
Theoretical Curve Drawn On the Basis of Equation
0.020 0.015 0.010 0.005 0.000
0
10
20
30
40
50
60
70
Compressive Strength in 103 psi Figure 9.25 Penetration rate for diamond drilling.
For many reasons, explained elsewhere in this book, rotary drilling with diamond bits was not practiced in the blasthole drilling industry. But soon, when bits with new type of diamonds will be available, diamond drilling may be introduced for blasthole drilling. Fracture process in diamond bit and drag bit drilling is similar hence the equation can also be used for predicting penetration rates of drag bit drilling.
9.6.3
Bauer, Calder and Workman approach
In the early 1960s rotary blasthole drilling became well established in blasthole drilling activities. Very heavy blasthole drills were developed for drilling in extremely hard formations, met in the iron ore mines, with the tungsten carbide inserted tricone bits. Bauer and Calder studied the subject of working performance of TC inserted tricone bits in hard iron ores of Mesabi Range in USA and developed the following empirical equation in 1967 for prediction of the penetration rate. log(P/6) = K log(12W/SC) where P = Penetration rate in ft/hr
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SC = UCS in thousands of lb/in2 W = Feed force exerted on the bit in terms of lb/in of bit diameter K = A factor that depended on the rock and varied between 1.4 and 1.75 for rocks of UCS from 15000 lb/in2 to 50000 lb/in2 In 1971 Bauer modified the equation by adding one more factor i.e. rotary speed. His equation was P = (61 − 28 * log SC)(W/φ)(N/300) where P SC W N φ
= = = = =
Penetration rate in ft/hr UCS in thousands of lb/in2 Feed force exerted on the bit in terms of thousands of lb Rotary speed of the bit in RPM Diameter of the bit in inches
In hard iron ores the values of penetration rates projected by using this equation tallied well with the actual measurements taken in the field but gave erroneous results when the iron ores were of less uniaxial compressive strength. To overcome these shortcomings investigations were done by Calder and Workman. They used data obtained in blasthole drilling in formations of lower uniaxial compressive strength and proposed a modified equation in 1994 as under, P = 5.7 * 10−5 * (RF–28 * log(0.145 * SC))W * N where P RF SC W N
= = = = =
Penetration rate in m/hr Rock Penetration Factor UCS in MPa Feed force exerted on the bit in terms of kg/mm Rotary speed of the bit in RPM
The rock penetration factor RF can be found from Table 9.5. If the UCS of the formation lies somewhere in between the limits of each class appropriate value from the class should be taken from the curve shown in Figure 9.26.
Table 9.5 Rock penetration factors.
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Rock description
Uniaxial comp. strength psi
Uniaxial comp. strength MPa
Rock factor
Very Hard Hard Moderate Soft Very Soft Extremely Soft
More than 30000 15000–30000 10000–15000 5000–10000 1000–5000 Less than
More than 207 103–207 69–103 34–69 7–34 1000 Less than 7
61 180 100 135 200 300
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The rock factor RF was determined by using the penetration rate data shown in Figure 9.27. In the use of this equation over the past decade it has been found that this equation gives acceptably correct estimate of penetration rate to be expected in blasthole drilling. In all the above equations it has been presumed that the bailing velocity of the compressed air is sufficiently high and the blasthole drilling process is not hampered due to insufficient flushing.
Rock Penetration Factor
350 300 250 200 150 100 50 0 0
25
50
75 100 125 150 175 200 225 250 275
Uniaxial Compressive Strength in MPa Figure 9.26 Rock penetration Factor vs. UCS data curve used for interpolating rock factor. 100 90
Penetration Rate in m/hr
80 70 279 mm dia 311 mm dia
60 50 40
251 mm dia
381 mm dia
200 mm dia
30 20 10 0 1000
10000 Formation UCS in lbs
100000
Figure 9.27 Penetration rate in rock mass of different UCS and different blasthole diameters.
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It is also presumed that the formation in which drilling is carried out is dry and the dry dust collection method as described in chapter 7 is used for the dust suppression. EXAMPLE
Blasthole drilling is to be carried out in a rock formation having UCS of 200 MPa by using TC inserted tricone bit of 311 mm diameter with a feed force including the drill head weight of 30,000 kg at a rotary speed of 80 rpm. What penetration rate can be expected? SOLUTION
As per chart in Figure 9.26, for the rock UCS of 200 MPa the rock factor works out to 65. The feed force exerted on the bit is 30000/311 i.e. 96.46 kg/mm. These factors give P = 5.7 * 10−5 * (65 − 28 * log(0.145 * 200)) * 94.46 * 80 Evaluating the above equation gives P = 10.36 m/hr.
9.6.4
Specific energy approach
Following equation has been suggested for calculating penetration rate based on specific energy of the rock. P = 48 * R * C/(π * D2 * Es) where P R C D Es
= = = = =
Penetration rate in in/m Drill head rotary power in in.lb/min Energy transmission factor Blasthole diameter in inches Specific energy in in.lb/in3
The most common values of C i.e. energy transmission factor lie between 0.6 to 0.8. The value of Es can be determined by the following equation. Es = 13900 * CRS + 15500 where Es = Specific energy in in.lb/in3 CRS = Coefficient of rock strength (From Table 9.6.)
9.6.5
Microbit test approach
In this test a large size rock sample is taken from the worksite to a testing laboratory. A hole is drilled in the sample by a two or three cone drill bit of small size.
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Table 9.6 Coefficient of rock strength for some rocks. Rock name
CRS
Rock name
CRS
Humboldt Iron Silicate Hornblende Schist Granite Pegmatite Wausau Quartzite Wausau Argillite Mankato Stone New Ulm Quartzite Jaspar Quartzite Rockville Granite Charcoal Granite Diamond Gray Granite Dresser Basalt
2.39 1.64 0.77 0.78 2.28 0.45 0.75 1.01 0.84 1.21 0.82 2.86
Shelly Limestone Mount Iron Taconite Aurora Taconite Babbitt Taconite Sabbitt Diabase Ely Gabbro Trap Rock Anorthosite Duluth Basalt Marble Primax Gabbro Iron Ore
0.57 1.47 2.62 2.84 2.44 1.21 0.64 0.73 2.11 0.98 1.02 1.23
The penetration rate attained in such tests is carefully measured and then the results are extrapolated to the actual drilling conditions at the worksite. It is necessary to carefully prepare the rock sample for the test, and drilling parameters such as weight on the bit, rotary speed etc. have to be suitably adjusted so that the extrapolations, to be carried out later on, give correct results. The test is repeated 5 to 10 times with a new bit for each of the tests. Mean penetration rate measured in the tests is used for interpolation. Interpolation is done on the basis of a chart similar to the one shown in Figure 9.28. The chart in Figure 9.28 is presented for illustrating the method. Tricone bits, on the basis of which the chart was designed, are old types and have now been largely replaced by more efficient bits. The microbit drilling approach was usually adopted by tricone bit manufacturers since they had very precise knowledge of the physical and geometrical parameters of their bits and manufacturing a suitable scaled down version of the bit was much easier for them. At first glance the microbit drilling test appears to be better placed for giving more accurate results for penetration rates but actually it is not so because the scale factor is not very easy to establish. Further, the test is also very costly as it involves use of specially made small drill bits rather than readily available simple tools. In many cases a simple masonry drill bit can be used for drilling a hole in the sample but it is more akin to drag bit or diamond bit drilling. Manufacturers of tricone drill bits usually have a large database of the use of tricone bits in various types of rock formations drilled in several places. From such databases they have prepared graphs of tricone bit penetration rates as shown in Figure 9.29. In most cases these graphs give very reliable values of penetration rates that can be expected in the field.
9.6.6
Indenter test approach
This test is also called as stamp test because the indenter makes a stamp on the rock.
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100 90 80 70 60 50 40
A - Carthage Marble, Sandstone, Limestone, Gypsum B - Dolomite Lime, Dolomite, Hard Shale C - Granite, Chert, Quartzite
30
M Pe icro ne bi tr t at io n
Ra
te
in
Ft
/h
r
10.0 9.0 8.0 7.0 6.0 5.0 4.0 3.0
4.0
2.0
3.5 3.0
1.0 0.9 0.8 0.7 0.6 0.5 0.4
0.2
0.2
0.1 10
2
3
4
5 6 7 8 9
2
3
4
0.1 0.04
0.06 0.02
0.4
0.3
0.6
0.8
1.0
1.5
2.0
2.5
Penetration Rate in Ft/hr
20
5 6 7 8 9
100 Weight on the Bit in lb/inch of Bit Diameter 1000
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Figure 9.28 Penetration rates based on microbit test.
2
3
4
5 6 7 8 9 10000
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100 8 6
ne 20 Sh 00 ale ps 35 i 00 ps Do i lom Ma rb ite le 89 70 00 00 psi ps i
ds to
2 Sa n
Penetration Rate in ft/hr
4
0 8 6 4
Sandstone 21000 psi 2
Quartzite 60000 psi
Basalt 46000 psi
Granite 51000 psi 1 102
2
4
6 8 103
2
4
6 8 104
Weight on the bit in lb/inch
Figure 9.29 Penetration rates based on database with a manufacturer (at 60 RPM).
This test is used for predicting drillability of rocks in rotary percussion as well as rotary drilling. Since a small tungsten carbide insert is used as an indentor rather than a tricone bit, the test can be carried out in any material testing laboratory. In their mode, penetration of an indenter into rock sample and the tooth of a tricone drill bit into in situ rock mass are identical to each other. An equation that gives the relation between several factors encountered in the process of rotary drilling is as under: V = 0.159 * K * (P/S)–1(W *N /(R * B)) where V W N K P/S R B
= = = = = = =
Penetration rate Weight on the bit Rotary speed in rpm A parameter related to rock Parameter related to penetration of indenter into the rock Radius of the load bearing tooth of the tricone bit Number of teeth on a bit in contact with the rock and bear the load
The above equation can be derived from fundamental principles. Of the above parameters K and S are determined by tests carried out in the laboratory. Parameters B and R can be known from the details of the bit provided by the manufacturer or actual physical observation and counting of the bit teeth.
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Parameters W and N are governed in actual drilling by using the controls provided on the blasthole drill. The sample to be used for the stamp test need not be a core sample but can be of any shape. The size of the sample should be such that the least dimension is more than 100 mm. If the rock is soft, larger sample sizes of the order of 200 mm are desirable. In fact a larger sample always proves ideal. Care should be taken in selecting an intact sample without cracks. For testing, the sample is embedded in concrete and the block is then cut by circular diamond saw so the sample surface is fresh and undisturbed. In the actual test a tungsten carbide stamp of circular cross section is pressed in the rock sample as shown schematically in Figure 9.30. The stamp is chosen in such a way that it has the same contact area as that of the insert on the drill bit to be used in actual drilling. Initially the stamp is in contact with the rock surface with little force. During test the force on the stamp is increased and the penetration of the indenter in the rock is continuously measured. At one juncture a crater is formed and the load on the stamp gets suddenly released. This is one cycle of the stamp test in which the stamp makes one indenture in the rock sample. The test is repeated at four or more positions on the sample. These positions are selected in such a way that the craters formed in earlier tests remain and are unaffected by the new stamp penetration. Apart from force and penetration measurements taken during the tests, the volume of material removed in each of the indentures is also measured. For this, weight of the material fragmented in the indenture is precisely measured and its volume is calculated on the basis of density of the sample. Graphs of force vs. penetration are drawn to determine the maximum stamp force for the penetration. One such graph is shown in Figure 9.31. Once the average values of stamp force FST, stamp penetration XST and crater volume VST are known, the stamp strength is calculated by the formula σST = FST/(4 * π * B2) In actual practice the test is repeated several times and the average values of FST, XST and VST are taken for calculations. Surface Carefully Prepared for the Test
Force Exerted on the Indenter
Tungsten Carbide Indenter
Rock Sample
Concrete Embedding the Rock Sample
Figure 9.30 Schematics of stamp test.
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Fst
Xst Figure 9.31 Force penetration curves from a stamp test.
The stamp strength is found to be much higher than the uniaxial compressive strength because in the stamp test the rock sample is virtually in semi infinite state. Test results show that stamp strength lies between 5 to 15 times the uniaxial compressive strength. In the case of tricone bits each of the three cones have many teeth in the form of milled points or tungsten carbide inserts. During rotation of the bit some of the inserts are in contact with the formation. All the weight on the bit is used by these teeth for penetration into the rock. In one rotation of the bit many (not necessarily all) teeth on the cones come in contact with the formation. The weight exerted on the bit has to be sufficiently high to exceed the value FST on each of the inserts in contact with the formation. Suppose the bit used for drilling has diameter D mm and the number of teeth of the bit coming in contact with the formation in one revolution is T rpm and the rotary speed is N, then the volume of material removed per revolution will be N * VST. Thus in one minute the volume removal will be T * N * VST. Since the cross sectional area of the hole is o * D2/4 the penetration rate P in terms of mm/min can be calculated by P = 4 * T * N * VST/π * D2 where P = Penetration rate in mm/min VST = Crater volume in mm3 Nowadays this test is becoming popular for the following reasons. 1 2 3 4
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It gives more reliable estimates of the penetration rates. It needs only one or two tungsten carbide indenters and a universal testing machine. It can be completed at short notice and can be carried out at virtually any material testing laboratory. Since it does not require any specially machined micro bit the test is normally far more inexpensive compared to a microbit test.
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Chapter 10
Flushing the blasthole
10.1
INTRODUCTION
A blasthole, or for that matter any type of hole in the ground, cannot be drilled without taking out the fragmented material from the hole. Flushing is a component of drilling activity, where a flowing medium is taken to the bottom of the hole by exerting pressure on it so that it returns to the surface of the ground and brings along cuttings with it. As mentioned in chapter 2 of this book, flushing is one of the essential components of the drilling process. In rotary blasthole drilling the flushing medium happens to be compressed air. In fact the power required to be spent in compressing the air to be used for flushing is even more than that required for any other essential drilling activity. For efficient blasthole drilling, the flushing has to be carried out with a well calculated plan. With the possibility of varying the volume of compressed air generated by the compressor, a driller can accomplish significant power saving if he is well acquainted with the process of flushing. This chapter is devoted to giving essential technical knowledge of the compressed air flushing process as used in rotary blasthole drilling.
10.2 WHY COMPRESSED AIR? Way back in 1946, i.e. in the beginning of rotary blasthole drilling, initial attempts were made to drill blastholes with water and mud as the flushing fluids. Later, when compressed air was used for flushing, it was noticed that compressed air cleaned the holes very efficiently and the size of cuttings ejected from the blasthole was far larger than that possible with water or mud circulation. This resulted in a far higher penetration rate. For this reason compressed air has been invariably used as a flushing medium in rotary blasthole drilling from the year 1950. The other reason for the preference for compressed air flushing was due to the fact that it gave a dry hole in which explosives could be filled directly i.e. without need for dewatering them.
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10.3
Rotary drilling and blasting in large surface mines
SCHEMATICS OF FLUSHING
When a blasthole is flushed with compressed air, the compressed air is generated at the compressor located on the main frame of the blasthole drill. It then goes to the drill head from where it enters the rotating drill string through the swivel of the drill head. This has been well illustrated in Figure 8.3. The passage of compressed air in a blasthole is shown in Figure 10.1. When the compressed air comes out of the nozzles of the drill bit and impinges on the formation, the cuttings formed by the interaction between the drill bit and formation get lifted from the bottom and may start moving upward, depending upon the forces acting on them. Three forces acting on the cutting are shown in the inset in the Figure 10.1. These three forces are gravitational force, buoyant force and drag force.
Compressed Air Moving Down through the Central Bore of the Drill Drill Pipe Compressed Air Moving Up through the Annular Space Between Hole Walls and Drill Pipes
Drag Force Buoyant Force
Gravitational Force
Drill Bit
Figure 10.1 Flushing schematics.
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10.3.1
297
Gravitational force
Gravitational force is due to the gravitational field of the earth. It is proportional to the mass of the particle and is constant throughout the fall of the particle. For a spherical particle it is given by Fg
= 4 * π * r3 * dm g/3
where Fg = Gravitational force r = Radius of the falling spherical particle dm = Density of the material of particle g = Gravitational constant. Gravitational force always acts in a vertically downward direction.
10.3.2
Buoyant force
Buoyant force is due to the displacement of the medium in which the particle is falling. This displacement is due to the volume of the falling body and can be written as Fb = 4 * π * r3 * df g/3 where Fb = Buoyant force r = Radius of the falling spherical particle df = Density of the material of particle g = Gravitational constant. Buoyant force always acts in a vertically upward direction.
10.3.3
Drag force
When a particle is suspended in a stream of moving fluid it is always being dragged in the direction of fluid flow. The drag force is generated due to the friction between the particle surface and the medium. Besides, the particle has also to displace the medium and make room for it to move. Both these components of the drag force are dependent upon the velocity at which the relative movement between the particle and medium takes place. The component of the drag force arising out of friction can be called Fdf and the component of drag force arising out of medium displacement can be called Fdd. With this we have the total drag force Fd as Fd = Fdf + Fdd Fdf and Fdd are equated as
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Fdf = Cdf * A * df * V2/2 and Fdd = Cdd * A * df * V2/2 where Cdf = Coefficient of drag due to the friction between the particle and medium Cdd = Coefficient of drag due to the need for displacing the medium df = Density of the material of particle V = Velocity at which the particle moves A = Area of cross section of the particle in the direction perpendicular to the direction of its movement. It can also be written as Fd = (Cdf + Cdd) * A * df * V2/2 or Fd = Cd * A * df * V2/2 When the velocity of the particle is low the value of Fdf makes a much larger contribution to Fd, but at very high velocity of the particle the contribution of Fdd to Fd is far higher than that of Fdf.
10.3.4 Terminal velocity Balancing these three forces acting on the particle an equation can be written as Gravitational force = Drag force + Buoyant force or Net Force on the particle = Gravitational force − (Drag force + Buoyant force) When the particle is falling in the medium its velocity keeps on increasing due the action of the gravity. The gravitational and buoyant force remain almost the same. The drag force keeps on increasing due to increase in the velocity of the particle. At one juncture the drag force increases and together with the buoyant force balances the gravitational force. At this juncture the net force on the particle becomes zero and the acceleration of the particle also becomes zero. After this the falling particle continues downward movement at a constant velocity, which is called the ‘terminal velocity’. In chemical engineering, where sedimentation of small particles is considered, it is also called settling velocity. The equation for settling velocity for spherical particles can be derived as follows: 4 * π * r3 * dm g/3 − 4 * π * r3 * df g/3 = Cd * π * r2 * df * V2/2 4 * r * dm g/3 − 4 * r * df g/3 = Cd * df * V2/2 V = ((8/(3 * Cd)) * r * g * ((dm − df)/df))0.5
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where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid Cd = Coefficient of drag g = Gravitational constant equal to 9.80665 m/s2. V = Velocity of the falling spherical particle.
10.3.5
Laws for settlement of particles in fluid
While considering the problem of settling of small particles in fluids, Stoke formulated that the drag force acting upon a particle moving in the fluid is given by the equation Fd = 6 * π * μ * r * V Later, he went on to derive the equation for settling velocity as follows: 4 * π * r3 * dm g/3 − 4 * π * r3 * df g/3 = 6 * π * μ * r * V i.e. 2 * (dm − df) * r2 = 9 * μ * V i.e. V = 2 * r2 * (dm − df)/(9 * μ) where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid μ = Dynamic viscosity of the medium g = Gravitational constant V = Velocity of the falling spherical particle. The above equation is known as Stoke’s Law. Stoke’s Law is valid when the fluid flow is in the region of Reynold’s Number from 0.0001 and 2. When the fluid flow is in the region of Reynold’s Number from 2 to 500 the settling velocity is given by an equation called the Intermediate Law. This equation is V = (0.306 * g0.71 * r1.14 * (dm − df)0.70)/(df0.29 * μ0.43)) where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid Cd = Coefficient of drag
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g = Gravitational constant equal to 9.80665 m/s2 V = Velocity of the falling spherical particle. Finally for fluid flow in the region of Reynold’s Number from 500 to 200,000 the settling velocity is given by an equation called Newton’s Law. In this region the coefficient of drag is fairly constant and is given by Cd = 0.44. Substituting this value into the equation for settling velocity we get, V = ((8/(3 * 0.44)) * r * g * ((dm − df)/df))0.5 = (6.06 * r * g * ((dm − df)/df))0.5 where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid g = Gravitational constant V = Velocity of the falling spherical particle. The important assumptions made in arriving at the above formula are: 1 2 3 4 5
The shape of the falling body is perfectly round so the surface area is at a minimum for the given volume. The surface of the falling body is very smooth. The falling bodies are located far apart from each other so that the presence of one body does not affect the fall of other falling bodies. The air through which the body is falling is still. The extent of air column through which the body is falling is several times larger than the diameter of the falling body.
Due consideration has been given to the deviations from these assumptions made in the following treatment of the subject. For a typical rock cutting formed in the formation fracturing process, we have a particle density of 2700 kg/m3, particle diameter 5 mm or 0.005 m and air density of 1.225 kg/m3. Substituting the values we get the terminal velocity as 18.09 m/s i.e. 1085.5 m/min or 3561 ft/min. It can be said that if the spherical particle is in a large annular space, when the upward velocity Va of the air in the annulus equals the terminal velocity Vt, the particle gets suspended in the air and will not fall down. Similarly, if the upward air velocity Va is more than the terminal velocity of the body Vt then the particle will be thrown up by the air at a speed of Va − Vt. In blasthole drilling practice, compressed air forced through the drill pipes is jetted out at the bottom of the hole through the bit nozzles. This air then drags the cuttings with it through the annular space between the hole walls and drill pipe. The
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upward velocity of air in the annular space is called ‘bailing velocity’ because at that velocity of the compressed air the particles are bailed out of the blasthole. Bailing velocity can easily be calculated by dividing the volume of compressed air by the area of the annular space. The equation can be expressed as Vb = 4 * 106 * Q/(π * (D2 − d2)) where Vb = Bailing velocity in m/min Q = Discharge of compressor in m3/min D = Diameter of the drill hole i.e. drill bit in mm d = Outer diameter of the drill pipe in mm Discharge volumes of compressed air are always given as free air discharge, meaning that the air will occupy an equivalent volume when the pressure of the air comes down to that of standard atmospheric air. For this purpose the 101.325 kPa is considered as standard atmospheric pressure and the discharge is also called SCFM when it is being expressed in imperial units. When drilling is carried out with water or mud circulation, the velocity of fluid moving upward is nearly constant through the depth of the hole. In contrast, in the case of compressed air circulation used in a blasthole drilling, the bailing velocity varies continuously because as the pressure drops, the air continuously expands. It is highest at the mouth of the hole.
10.4
FORMULATION OF DESIRED BAILING VELOCITY
How much bailing velocity is appropriate for particular conditions of the blasthole is a debatable matter. Under certain conditions in a blasthole, cuttings formed may be very large but the quantum of such cuttings may be very low, i.e. less than 0.5%. In such cases it is inappropriate to have a very large compressor working all the time. It is certainly more economical to get the chips re-fragmented and taken out of the blasthole. An appropriate value of the desired annular bailing velocity Vd for particular drilling conditions can be equated as Vd = 1.1α Vt where α = A factor depending upon many conditions described hereunder. Vt = A value for terminal velocity calculated on certain assumptions. Let us say that Vt is equal to 1100 m/min as calculated earlier for 5 mm spherical particles of density 2700 kg/m3 at mean sea level with air density of 1.225 kg/m3. The
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multiplier 1.1 is included so that the desired bailing velocity is 10% higher than the terminal velocity, thereby enabling the particles to move up with a velocity equal to 10% of the terminal velocity. The equation, therefore, turns into: Vb = 1210α The value of the correlation factor α must account for all the deviations from the assumed conditions used for calculating the value of 1100 m/min on the basis of the equation for terminal velocity. Aspects to be considered while arriving at an appropriate value of the correlation factor are: 1 2 3 4 5 6 7 8
Size of Fragments Density of Fragments Roundness of Fragments Roughness of Fragments Rate of Fragmentation Annular Space Inclination of the Blasthole Amount of Water Injection (if any) In other words the correlation factor α can be equated as α = αrαdαrouαrndαfrαaαiαw
where αr = Component depending upon the size of the fragments. αd = Component depending upon the density of fragments i.e. rock. αrou = Component depending upon the roundness of fragments. αrnd = Component depending upon the roughness of fragments. αfr = Component depending upon the rate of fragmentation of fragments. αa = Component depending upon the available annular space. αi = Component depending upon the inclination of the blasthole. αw = Component depending upon presence of water injection. In formulating the above equation it has been presumed that all these components of the correlation factor are independent of each other. Values of each of the above correlation factors can be determined from the conditions actually met in drilling on the basis of the elaboration given in the following paragraphs.
10.4.1
Size of fragments
Rock chips are formed at the bottom of the hole. In order that these chips are smoothly removed from the hole, they must essentially be smaller than about 90% of the size of the smallest passage through which they have to pass. If rock chips are larger than this size, the chips have to reduce in size by undergoing re-fragmentation.
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In almost any type of drilling with any type of bit, the smallest passage through which the rock chip has to pass is the flushing groove of the bit as shown in Figure 10.2. Studies of the geometry of different types of drill bits have suggested that most of the bits have 2 to 4 flushing grooves, each with a mean dimension of about 7 to 8% of the bit diameter. It is therefore appropriate to assume that the size of largest cutting is about one fifteenth of the bit diameter. For small and medium size blastholes field observations match very well with this correlation. In the case of large blastholes of diameter 300 mm or so, the largest cuttings were found to be of 19 mm size. Since the desired bailing velocity of 1100 m/s has been calculated for the particles of 5 mm size, which is equivalent to a hole diameter of 75 mm, the correlation factor αr will have a value of 1 for this blasthole diameter. Further, as per the equation for terminal velocity, the effect of particle size will be proportional to the square root of the diameter. With this we can determine one component of the correlation factor, viz. αr, on account of size of fragment as αr = (D/75)0.5 where D is the diameter of hole in mm.
Blasthole wall
One of the three grooves of tricone bit for cutting removal
Figure 10.2 Grooves around a tricone bit.
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10.4.2
Density of fragments
The density of rock being drilled, assumed while arriving at the value of 1100 m/min, is 2.7 (2700 kg/m3). In practice, the density of rock varies between about 2 for shells to 5 for magnetite ore. The correlation factor αd on account of particle density can be written as αd = (ρ/2700)0.5 where ρ is the density of the particles in kg/m3.
10.4.3
Roundness of fragments
While arriving at the desired bailing velocity of 1100 m/s the particles were assumed to be perfectly spherical. Roundness of fragments is very important. In almost all types of rocks the chips are far from spherical. In igneous or metamorphic rocks the chips are closer to tetrahedron or parallelepiped shapes, whereas in the case of sedimentary rocks they are close to flat disks. Coefficient of drag for non spherical chips is much higher, and therefore the bailing velocity required for lifting off such particles is much less than that required for spherical particles. Roundness or sphericity of a particle can be defined as the ratio of the surface area of a sphere of the same volume as the non spherical particle to the surface area of the non spherical particle. Table 10.1 presents sphericity coefficients of typical shapes. From the values presented in the table, it can be concluded that for particles of igneous and metamorphic rocks the value of sphericity of the cuttings will be about 0.75 and for sedimentary rocks it will be about 0.6. Very roughly, the drag coefficient can be equated to the sphericity. Therefore, the value of αrnd will be either 0.75 or 0.6.
10.4.4
Roughness of fragments
Roughness of the surface of particles does affect the terminal velocity. When the surface is rough the terminal velocity is less. As the size of the grains which form rock is coarser, the surface of rock particles becomes more rough because fracture generally takes place at the joint between two particles. However, since the particles formed in the drilling process are small, their roughness does not affect the terminal velocity
Table 10.1 Sphericity of shapes.
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Particle shape
Sphericity
Particle shape
Sphericity
Sphere Octahedron Cube Prism L × L × 2L Prism L × 2L × 3L
1 0.85 0.81 0.77 0.73
Cylinder H = R/15 Cylinder H = R/3 Cylinder H = R Cylinder H = 2 Cylinder H = 20R
0.25 0.59 0.83 0.87 0.58
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greatly. In case of large size chips, the reduction in terminal velocity on account of roughness of the particle surface can be as much as 10%. An estimated correlation factor on account of particle roughness can be written as αrou = 1/(1 + (D/1000)2) where D is the diameter of the hole.
10.4.5
Rate of fragmentation
Rate of fragmentation is directly dependent upon penetration rate. A higher penetration rate means more fragments. If the volume of fragments to be removed from the blasthole is large, a higher bailing velocity is needed. It is proposed that if the penetration rate is more than 15 m/h then for every 1 m/h increase in the penetration rate, the bailing velocity should be increased by 1%. An estimated correlation factor on account of the penetration rate of fragmentation can be written as αfr = (1 + 0.01 * (Pr − 15)) where Pr is the penetration rate in m/h.
10.4.6
Annular space
From field observations it has been noticed that for effective removal of rock cuttings the annular space should be about 17% of the cross sectional area of the blasthole. It is proposed that if the percentage of annular space is less than 17%, then for every 1% reduction in this percentage the bailing velocity needs to be increased by 2%. An estimated correlation factor on account of insufficient annular space can be written as αa = (1 + 0.02 * (17 − A)) where A is the percentage of annular space for the combination of the bits and drill rods to be used in actual drilling.
10.4.7
Inclination of the blasthole
If a hole is inclined the tendency of the fragments to dash against the wall of the hole is much greater. In such circumstances a higher bailing velocity is necessary for effective cutting removal. It is proposed that for every 1° inclination of the blasthole, the bailing velocity be increased by 0.5%.
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An estimated correlation factor on account of blasthole inclination can be written as αi = (1 + 0.005 * (I)) where I is the angle of inclination with the vertical in degrees.
10.4.8
Quantity of water injection
In many blasthole drills water is injected into the compressed air to minimize dust pollution. This method of dust suppression is not ideal but it is an inexpensive method. When water is injected into the compressed air a mist is formed. This misty air is then circulated in the hole. Lifting formation fragments in such misty air is rather difficult because particles – particularly the sand and dust size – tend to stick together and form somewhat larger chunks to be removed by the circulating compressed air. It is proposed that if water injection is used while drilling the bailing velocity should be increased by 10%. Thus, the correlation factor on account of the presence of water injection needs to be equal to 1 when water injection is not used and be equal to 1.1 when water injection is used. With all the reasoning and correlation factors mentioned above the final equation for bailing velocity takes the form as under, Vd = 1210αrαdαrouαrndαfrαaαiαw The above is the final equation for desirable bailing velocity of flushing air. It is particularly valid in blasthole drilling as the drilling depth is usually less than 30 m and almost never exceeds 50 m. In actual practice, a bailing velocity of about 900 m/min is considered as the bare minimum required, and about 1000 m/min is desirable. While drilling large diameter blastholes in soft rocks, bailing velocities of about 2000 m/min have given very satisfactory results. In heavy iron ores, bailing velocities of the order of 2500 m/min have given good results. In very abrasive formations bailing velocities higher than 2000 m/min cause very heavy abrasive wear on the drill string components. Therefore, bailing velocities beyond 2400 m/min are rarely used. This restriction has very little effect on the penetration rate. To look into the correctness of the above hypothesis and formula, bailing velocities for three typical cases have been calculated in Table 10.2. These values for desired bailing velocities, calculated by using the empirical formula for the three random cases of blastholes, tally well with results obtained in standard field practice. It thus appears that the empirical formula for bailing velocity can be used under many drilling conditions. Compressed air is far more efficient in lifting formation fragments from the hole bottom as compared to water or water-based mud used in oilwell drilling. The basic
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Table 10.2 Desired bailing velocity in m/min calculated after taking all the correlation components into consideration. Case 1
Case 2
Case 3
254 mm hole inclined at 20° in granite of density 2.6, sphericity 0.75, pen. rate 22 m/hr, annular space 18%, with water injection
406 mm vertical hole in sandstone of density 2.6, sphericity 0.6, pen. rate 35 m/hr, annular space 15%, without water injection
Factor
Aspect considered
311 mm vertical hole in iron ore of density 5.2, sphericity 0.75, pen. rate 18 m/hr, annular space 17%, without water injection
αr
Size of Fragments
2.03633658
1.840289832
2.326657116
αd
Density of Fragments
1.387777333
0.981306763
0.981306763
αrnd
Roundness of Fragments 0.75
0.75
0.6
αrou
Roughness of Fragments 0.911808929
0.939394053
0.858489951
αfr
Rate of Fragmentation
1.03
1.07
1.2
αa
Annular Space
1
0.98
1.04
αI
Hole Inclination
1
1.1
1
αw
Water Injection
1
1.1
1
2408.557682
1953.352769
1775.92082
Aspect for correlation factor
Desired Bailing Velocity m/min
reason for the efficiency of compressed air in lifting the cuttings of the bottom of the hole is that when compressed air escapes from the drill bit, it expands and the fragments are lifted from the bottom due to the drag force. Further the drag force on the particle keeps on acting for a long time due to continuous expansion of the compressed air that takes place till the particle reaches the mouth of the hole and escapes to the atmospheric air. This gives very high velocity to the fragments. As the depth of the blasthole increases, the requirement of compressed air pressure also increases. It is very important to keep the bailing velocity at the optimum as calculated in Table 10.2. Lower bailing velocity results in the following: 1 2 3 4 5
Regrinding of the chips under the drill bit. Reduced penetration rate. More chances of premature drill bit damage. Higher bit cost and hence higher drilling cost. Higher volume of fines and hence more load on the filters of the dust collector. Higher bailing velocity results in the following:
1 2
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Higher power consumption and hence higher drilling cost. High wear and tear on drill string components.
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10.4.9
Other approaches
Many writers have given deep consideration to the formulation of the desired bailing velocity required in rotary blasthole drilling. Equations put forth by them are empirical in nature and are as follows: Vd = 573 * (ρr/(ρr + 1)) * dp0.6 Bauer and Crosby have put forth an equation for balancing velocity i.e. velocity for keeping rock particles floating in the compressed air flow: Vd = 250 * ρ0.5 d0.5 r * p In both the above equations, Vd = Desired bailing velocity in m/min ρr = Density of rock chips in g/cm3 dp = Mean diameter of rock chips in mm. An equation proposed for the discharge capacity of the compressor on the blasthole drill is Qd = 224 * D0.5 where Qd = Air flow in m3/min. D = Diameter of Blasthole in m.
10.5
FORMULATION OF COMPRESSED AIR PRESSURE
What volume is adequate has been arrived at on the basis of various aspects of flushing discussed in the earlier part of this chapter. Here, in this part, the subject of discussion is how much pressure is needed. Compressed air must have a certain built-in pressure to ensure that an adequate volume of air flows through the flow passage. As stated earlier and shown in Figure 8.3, the components through which compressed air flows can be grouped into four sections. The number and type of these components changes from one blasthole drill to another. These four sections are described, and their general composition in terms of components, is given in Table 10.3. Loss in compressed air pressure occurs in each of these four sections. It is possible to calculate the pressure lost in these sections for every blasthole drill, but the calculations are very tedious and more importantly the theoretical treatment of such calculations is far from perfect. Hence it is necessary to carry out air tests. Manufacturers of rotary blasthole drills often carry out air tests on their blasthole drill models. For such tests the apparatus used, and the test procedure, has been described in chapter 8.
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Table 10.3 Different sections in compressed air flow path in a rotary blasthole drill. Section Flow section no. location
Components likely used for the flow path in the section Comments on diameter and length
1
From Compressor The flow path comprises of to the Drill Dead metal tubing, one gate valve, eight 90° bends, one 180° bend, one piece of short hose pipe, one piece of long hose pipe.
The diameter of the metal tubes and hose pipes equals to about half the diameter of the largest blasthole. Length of both the hoses together equals about 1.5 times the length of longest drill pipe. Length of metal tubing also equals to about 1.5 times the length of longest drill pipe.
2
From Drill Dead to the Bit
Apart from its length each drill pipe has one gradual enlargement and one gradual contraction.
3
From Drill Pipe The drill bit, three nozzles Side of the Bit to and the clearance kept in Blasthole Side of the bearings of the drill bit. the Bit
The diameter of the nozzles is to be selected as per procedure explained later in this chapter. The underlying principle is that 30% of the compressed air should flow through bit bearings.
4
Blasthole Bottom This section is formed to its Opening at outside the drill pipes and the Ground Level the drill bit. Hence, no components are involved.
A small part of this section is below and on the side of the drill bit. It is very irregular. Grooves are also formed around the stabilizer. Apart from these the full section is uniform and annular in cross section. The annulus is formed by the diameter of blasthole and the outer diameter of the drill pipes.
One to five drill pipes. One Stabilizer that may be roller type.
Giving due consideration to the components in the four sections and considering average values of compressed air flow, diameters and lengths of metal pipe and hose, blasthole diameters, drill pipe inner and outer diameters etc., the values of compressed air pressure loss in sections 1 + 2 + 4 have been arrived at in Table 10.4. If we consider that smaller diameter blasthole bits have smaller API pin connection and much smaller clearance in the bit bearings, the smaller bits must ensure higher pressure loss through the nozzles so as to ensure that about 30% of the compressed air flows through the bearings. On the basis of this logic, the desired pressure loss values when 70% of compressed air flows through the nozzles have been given in Table 10.4. Field observations roughly tally with the values of desired pressure loss given in the second last row, and the actual pressure provided on the compressors of the blasthole drills is given in the last row in the respective categories in Table 10.4.
10.5.1
Oil injection
As stated earlier, oil is often introduced into the compressed air flow in the form of mist through an oil injection system. The purpose of this process is bit lubrication.
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Table 10.4 Estimation of pressure loss in pa in four sections of the flow path in a rotary blasthole drill. Pressure loss for different blasthole drills having following recommended range of blasthole diameters in mm Section and component Pressure Loss in Section 1 (Pa) Metal Tubing Valves 9 × 90° Bends 1 × 180° Bend Hose Pipe
150 to 200
200 to 279
279 to 381
381 to 445
40707
28451
21948
18180
16282.8
11380.4
8779.2
7272
Pressure Loss in Section 2 (Pa) 5 Drill Pipes
6800
6938
3346
819
Pressure Loss in Section 3 (Pa) Around Bit Annulus
1500 22500
1200 21800
900 13200
600 9600
Total Pr Loss Section 1 + 2 + 4 Pa
114989.8
97521.4
61557.2
39747
Desired Pressure Loss at Bit Pa
550000
400000
280000
210000
Min. Desired Comp. Pressure in Pa in psi
664989.8 96.5
497521.4 72.2
341557.2 49.83
249747 36.2
125 to 100
110 to 100
85 to 50
65 to 50
Common Pressure Ratings of the Compressors in psi
With the introduction of oil, the resistance to flow of compressed air increases. The increase, however, is very small. Quite likely the fine film of oil formed on the walls of the flow path reduces friction factor.
10.6
CHOOSING A DRILL PIPE
The diameter of blastholes is chosen on the basis of many factors related to production requirements, blasting requirements, and parameters pertaining to loading and hauling equipment used in the mine etc. They are elaborated elsewhere in this book. When a drill pipe is to be chosen for drilling blastholes of a particular diameter, the following factors are required to be given due consideration: 1 2 3 4 5
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Drill Pipe Dimensional Parameters Drill Pipe Surface Treatment Size and Shape of Drill Cuttings Bailing Velocity in the Blasthole Drill Pipe Wall Thickness.
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The following elaboration of these factors will give ample guidelines for selection of a drill pipe. Later in this section, an example is given to clarify a systematic procedure for drill pipe selection.
10.6.1
Drill pipe dimensional parameters
As far as the dimensional parameters are concerned, drill pipes are available in many different external diameters, wall thicknesses, lengths etc. Dimensional parameters of commonly available drill pipes are given in Table 10.5. In blasthole drills, drill pipes are handled through many mechanical devices. For example, they are stored in pipe racks or carousel type pipe changers, handled by centralizer, tightened by wrenches etc. All this makes it necessary to have perfect dimensions of a drill pipe in terms of pipe length, length of threaded portion, position and dimensions of wrench flats etc., so they do not pose any problems while being handled by these devices provided in blasthole drills. Most of the suppliers of drill pipes ask for the make, model and some other relevant details of the blasthole drill with which the drill pipes are to be used. Giving such data to the suppliers is not sufficient by itself. Purchasers should ask for the external dimension drawings for the drill pipe from the suppliers, and verify that the drill pipe will fit in the blasthole drill being used by them, before finalizing an order.
10.6.2
Drill pipe surface treatment
Surface treatment given to blasthole drill pipes is of two types viz. surface hardening and anti-corrosion. In oilwell or waterwell drilling, water and mud are used as circulating fluid. The annular velocity of such fluids is much lower and the wear of drill pipes due to friction with the drill cuttings is very small. In blasthole drilling compressed air is used as a circulating fluid. The velocity of air flow and the drill cuttings is very high. This results in heavy wear on the drill pipes due to such continuous sandblasting of the external surface of the drill pipes. For this reason it is always desirable to choose surface-hardened drill pipes in blasthole drilling. This need is far more pronounced when blastholes are to be drilled in hard and abrasive rocks such as granite, quartzite etc. than in soft and non-abrasive rocks such as slate, limestone etc. In many rock masses a small quantity of groundwater seeps into the blastholes. Such groundwater may not pose any problem for drilling but generates mild acids or other chemicals by getting mixed with some minerals. When the drill pipes come in contact with such acids they start corroding. In such conditions anti-corrosion treatment is important. Many manufacturers make drill pipes for water well drilling to be carried out with DTH hammers. For waterwell drilling smaller wall thickness is desirable, so longer drill strings used in deep wells can be lifted with ease. Feed force and torque required for DTH drilling are also much lower than those in rotary drilling with tricone bits.
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Table 10.5 Dimensional parameters of commonly available drill pipes. Drill pipe OD Inch 3.5
mm
Wall thickness
Cross section area
Body weight
in
mm
in2
lb/ft
mm2
kg/m
88.9
0.75
19.05
6.48
4180.3
22.05
32.84
4
101.6
0.5
12.7
5.498
3546.9
18.71
27.87
4.5
114.3
0.375
4.86
3135.3
16.54
24.63
4.5
114.3
0.5
12.7
6.283
4053.7
21.38
31.85
5
127
0.5
12.7
7.069
4560.4
24.05
35.83
5
127
0.75
19.05
10.014
6460.5
34.08
50.76
5.5
139.7
0.5
12.7
7.854
5067.1
26.73
39.81
5.5
139.7
0.75
19.05
11.192
7220.6
38.08
56.73
6
152.4
0.75
19.05
12.37
7980.6
42.09
62.7
6.25
158.75
0.5
12.7
9.032
5827.1
30.73
45.78
6.25
158.75
0.75
19.05
12.959
8360.7
44.1
65.69
6.25
158.75
1
25.4
16.493
10640.9
56.12
83.6
6.5
165.1
0.75
19.05
13.548
8740.7
46.1
68.67
6.625
168.28
0.864
21.946
15.637
10088.5
53.21
79.26
6.75
171.45
0.75
19.05
14.137
9120.7
48.11
71.66
7
177.8
0.5
12.7
10.21
6587.2
34.74
51.75
7
177.8
0.75
19.05
14.726
9500.8
50.11
74.64
7
177.8
1
25.4
18.85
12161
64.14
95.55
7.5
190.5
1
25.4
20.42
13174.4
69.49
103.51
7.625
193.68
0.75
19.05
16.199
10450.8
55.12
82.11
7.625
193.68
0.875
22.225
18.555
11971
63.14
94.05
7.625
193.68
1
25.4
20.813
13427.7
70.82
105.5
8.625
219.08
0.906
23.012
21.97
14174.5
74.76
111.37
8.625
219.08
1
25.4
23.955
15454.6
81.51
121.42
8.625
219.08
1.5
38.1
33.576
21661.7
114.25
170.19
9.25
234.95
0.75
19.05
20.028
12921
68.15
101.52
9.25
234.95
1
25.4
25.918
16721.3
88.19
131.38
9.25
234.95
1.5
38.1
36.521
23561.9
124.27
185.12
9.25
234.95
1.5
38.1
36.521
23561.9
124.27
185.12
9.625
244.48
1.5
38.1
38.288
24702
130.29
194.08
10.75
273.05
1
25.4
30.631
19761.6
104.23
155.26
10.75
273.05
1.25
31.75
37.306
24068.6
126.94
189.1
10.75
273.05
1.5
38.1
43.59
28122.2
148.33
220.95
10.75
273.05
2
50.8
54.978
35469.5
187.08
278.68
11.75
298.45
1.25
31.75
41.233
26602.1
140.31
209.01
9.525
(Continued)
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Table 10.5 (Continued) Drill pipe OD
Wall thickness
Cross section area 2
2
Body weight
Inch
mm
in
mm
in
mm
lb/ft
kg/m
12.25
311.15
1
25.4
35.343
22801.8
120.26
179.15
12.75
323.85
1
25.4
36.914
23815.2
125.61
187.11
13.375
339.73
1
25.4
38.877
25082
132.29
197.06
13.375
339.73
1.25
31.75
47.615
30719.1
162.02
241.35
13.375
339.73
1.5
38.1
55.96
36102.9
190.42
283.65
14
355.6
1.5
38.1
58.905
38003
200.44
298.58
14
355.6
2
50.8
75.398
48643.9
256.56
382.18
14.375
365.13
1.5
38.1
60.672
39143.1
206.45
307.54
14.375
365.13
2
50.8
77.754
50164
264.58
394.13
15
381
1.25
31.75
53.996
34836.1
183.74
273.7
This makes the thin-walled drill pipes acceptable in waterwell drilling. Further, in water well drilling dimensional inaccuracies of the drill pipes are tolerable to a greater extent, as drill pipes are handled manually and stored on the ground and not in a pipe changer. For blasthole drilling such drill pipes do not have adequate wall thickness and are without any surface hardening treatment so they should be avoided.
10.6.3
Size and shape of drill cuttings
In the formulation of bailing velocity in the earlier part of this chapter, the following two assumptions were made in respect of the largest drill cuttings that are formed while blasthole drilling. 1 2
While drilling in igneous and metamorphic rocks the size of the largest cuttings is about 1/15 times the diameter of the blasthole. While drilling sedimentary rocks the size of the largest cuttings is about 1/12 times the diameter of the blasthole.
Reasoning behind these assumptions is that drill cuttings in sedimentary rocks are flaky and have sphericity of 0.6 as against 0.75 for the cuttings in igneous and metamorphic rocks. Observations about fluid flow patterns around objects of different shapes reveal that if a cutting has to pass smoothly through the annulus, the annulus should have thickness equal to 1.25 times the dimension of the cutting size. Thus the desired thickness T for the annulus can be equated as T ≥ D/12 for igneous or metamorphic rocks T ≥ D/9.6 for sedimentary rocks.
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With this hypothesis, the optimum outside diameter d for the drill pipe can be equated as d = D − 2 * (D/12) = 5 * D/6 = 0.833 * D for igneous and metamorphic rocks or d = D − 2 * (D/9.6) = 3.8 * D/4.8 = 0.792 * D for sedimentary rocks. In actual practice it may not be possible to get a drill pipe with outside diameter calculated as above. In such an event a drill pipe having the nearest, and preferably slightly smaller, OD should be chosen for blasthole drilling. Dimensional parameters for standard blasthole drill pipes are given in Table 10.4. An appendix at the end of this book gives more details.
10.6.4
Bailing velocity
In order to clean the hole effectively the compressed air flow in the annulus must have sufficient bailing velocity. This topic has already been discussed earlier in this chapter. Accordingly the desirable bailing velocity Vd can be calculated. Once the outside diameter of a drill pipe is determined, and a standard drill pipe is chosen as stated in previous subsection, it is necessary to check if the compressor on the blasthole drill has adequate discharge capacity to give the desired bailing velocity in actual operation. The desired capacity for the compressor can be calculated by the following equation: Qd = (π * (D2 − d2) * Vd)/(4 * 106) where Qd is the desired discharge capacity of the compressor on the blasthole drill in m3/min, D is the diameter of blasthole in mm, d is the diameter of drill pipe in mm and Vd is the desired bailing velocity in m/min. In almost all the cases it will be found that the compressor on the blasthole drill has more than adequate discharge capacity because compressors on blasthole drills are most usually chosen to drill blastholes of largest rated diameter.
10.6.5
Drill pipe wall thickness
Drill pipes with larger wall thickness are always advantageous in blasthole drilling. The reasons are: 1
2
3
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Drill pipes with larger thickness are heavier and their heavy weight adds to the feed force exerted by the feed mechanism to give even higher feed force on the drill bit. Unlike other purposes such as oilwell or waterwell drilling, in blasthole drilling the drill string is very short, so the heavier weight of the drill pipes does not pose a limitation on lifting the drill string. Drill pipes in oilwell or water drilling are always in tension, whereas those in blasthole drilling are in compression. Thus, blasthole drill pipes are susceptible to buckling. When a blasthole drill string is exerted with a feed force just above its
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315
critical buckling load, the drill string suddenly starts deflecting in the center. Such deflection substantially increases the possibilities of blasthole deviation. Drill pipes with greater wall thickness have a much higher section modulus and have a much higher critical load against buckling. Thus they do not buckle and contribute to hole deviation and other undesirable factors such as rubbing against the wall of the blasthole. A drill pipe with large wall thickness may be good in blasthole drilling for long life and higher bit loading, but a drill string composed of such drill pipes is frequently required to be lifted by using power. Consumption of power for lifting heavier drill strings is always higher. In this regard the following observations are quite self explanatory. 1
2
3
4
Very heavy drill collars are essential in the oilwell drilling industry as the drills normally do not have any mechanisms for exerting positive feed force on the bit. Bit weight is solely attained by use of heavy drill collars. Integral drill pipes are the same as drill collars with some dimensional difference. Integral drill pipes have very large wall thickness – in most of the cases far higher than an equivalent fabricated drill pipe, as can be seen from details given in one of the appendices at the end of this book. If larger wall thickness was the only criteria for selecting the best drill pipe for blasthole drilling, everyone would have chosen integral drill pipes rather than fabricated drill pipes. Actually the trend is otherwise. Manufacturers who solely deal in blasthole drilling products rarely attempt to manufacture integral drill pipes. Integral drill pipes are manufactured only by those who manufacture drill collars for oilwell drilling.
The above elaboration sufficiently indicates that in addition to the outside diameter due consideration must also be given to appropriate wall thickness of the drill pipe to be used for blasthole drilling. One important factor to be kept in mind while choosing a drill pipe for blasthole drilling is that the drill pipes are stored in pipe racks, which are vertically placed in the mast. If the drill pipes are very heavy the stability of the blasthole drill is seriously affected. When a drill pipe with a specific outside diameter is to be chosen from the table of available drill pipes given in one of the appendices at the end of this book, there may be many options of wall thickness for such drill pipes. Further, even if such an option does not exist, it is necessary to check the susceptibility of the available drill pipe to buckling. Critical buckling load can be defined as the axial compressive load at which the drill string starts deflecting somewhere in the midst of its length. In the case of a drill string with uniform cross section, the point at which deflection starts happens to be at the midpoint of the drill string. As per Euler’s formula, critical buckling load can be equated to properties of the drill string as under. P = 1000 * C * π2 * E * I/(1000 * L)2
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where P is the critical buckling load in newtons, C is the coefficient dependent upon the end fixing conditions for the drill string, E is the modulus of elasticity for the material of the drill string in GPa, I is the moment of inertia for the drill string in mm4 and L is the length of the drill string in m. Since components of the drill string are made from either low carbon or alloy steel tubes, E is equal to 200 GPa. Since most of the drill string length consists of drill pipes, the moment of inertia for the drill pipes can be used in calculations. For a drill pipe with outside diameter D in mm and inside diameter d in mm, the moment of inertia I in mm4 is given by the equation: I = (π/64) * (D4 − d4) If both the end points of the drill string are purely pin type, as shown in Figure 10.3A, then the value of C equals to 1. If both the drill string ends are rigidly fixed, as shown in Figure 10.3B, then the value of C equals to 4. In actual practice the drill string is restrained from buckling at the top by the deck bushing, and at the bottom it is usually restrained from buckling by the combination of drill bit and stabilizer. For such conditions the value of C is estimated at nearly 3.7. While using the length of drill string in Euler’s formula, the following points can be made. 1
When drilling is to be carried out by using only the base drill pipe, the drill string length is equal to the length of base drill pipe plus an additional length of about 2 m due to the drill bit, stabilizer, saver sub at the upper end and some length of
B
A
C
Column Length
Both Ends Pinned
Both Ends Fixed
Both Ends Semi Fixed
C=1
C=4
C = Variable (1 to 4)
Figure 10.3 End fixing conditions of drill string.
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2
317
the shock absorber. Therefore, the length L to be used for calculations works out to L = Lp + 2, where Lp is length of base drill pipe. When drilling is to be carried out by using many drill pipes, and the maximum depth of the blasthole, the length L to be used for calculations works out to L = Lh − 1, where Lh is depth of the blasthole. Deduction of 1 m is done because a stabilizer effectively restrains the drill string from buckling.
In actual calculations the larger of the two values should be used. Once the buckling load for the chosen drill pipe is calculated by using Euler’s formula it must be verified that the drill pipe can withstand the maximum rated feed force of the blasthole drill (and not the maximum feed force likely to be exerted on the drill string) with a factor of safety more than 1.25 to allow for the vibrations etc. Drill pipes affect the stability of a blasthole drill considerably. When the diameter and wall thickness etc. of the chosen drill pipe differs from the original, it is essential to verify if the stability of the blasthole drill will remain unaffected. This can be done by checking with the blasthole drill manufacturer.
EXAMPLE
An example of systematic selection of a drill pipe for rotary blasthole drilling operations is illustrated through Table 10.6. In the table, cell A1 states what is in the cell C1. Cell B1 specifies whether contents of cell C1 is data input or the results of the calculation. If the contents are data input, cell B1 mentions “data input”. If the contents are a formula cell B1 gives the formula used for calculations. Repetitions with the next higher wall thickness may be necessary if pipe thickness in cell C35 is found to be inadequate. In the example shown in Table 10.6 the depth of blasthole was taken as 20 m. This was because the drill was to be used in a copper mine where the primary loading equipment was an electric shovel and hauling was to be carried out by dumper. Had the drill to be used in a gigantic coal mine where a dragline was the primary equipment for removal of overburden, the blasthole depth would have been considerably greater, say about 50 m. In such a case the length for calculating the critical buckling load would have increased from 22 m to 49 m. As a result of this the critical buckling load would have reduced considerably.
10.7
DISCARDING A DRILL PIPE
As drilling is carried out by using a drill pipe the drill pipe goes on wearing; In particular its outer diameter reduces considerably. As the outer diameter reduces, the annular area increases and the bailing velocity decreases. In actual practice this decrease can be compensated for by increasing the volume flow from the compressor. Most of the modern screw compressors have the facility of increasing or decreasing volume flow by a simple control. Obviously, increasing the volume flow is possible up to the maximum discharge capacity of the compressor.
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Table 10.6 Choice of suitable drill pipe. Aspect
Formula used (if any)
Input/output
Cell ref.
Make and Model of the Blasthole Drill Used
Data Input
Bucyrus 49RIII
C1
Method of Drilling
Data Input
Rotary
C2
Maximum Rated Pulldown of the Blasthole Drill in kg
Data Input
54800
C3
Maximum Discharge of Compressor in m3/min
Data Input
103
C4
Pressure oh the Compressed Air in kpa
Data Input
448
C5
Maximum Rated Depth Capacity in m
Data Input
85.5
C6
Single Pass Length in m
Data Input
21.3
C7
Type of Formation to be Drilled
Data Input
Granitic Copper Ore
C8
Nature of Formation Igneous or Metamorphic or Sedimentary
Data Input
Igneous
C9
Feed Force Required on the Bit for Drilling the Formation in kg/mm
Data Input
90
C10
Blasthole Diameter in mm
Data Input
406
C11
Actual Feed Force on the Drill String in kg
C12 = C10 * C11
36540
C12
Can the Blasthole Drill Exert Necessary Feed Force?
C13 = IF(C12 < C3, "Yes", "No")
Yes
C13
Depth of Blasthole in m
Data Input
20
C14
0
C15
Inclination of the Blasthole in Degrees Data Input 3
Data Input
2700
C16
Data Input
12
C17
Annular Space Datum in %
Data Input
20
C18
Presence of Water Injection
Data Input
No
C19
Desired Drill Pipe Diameter in mm
338.198 C20 = IF(C9 = "Sedimentary", 0.792 * C11,0.833 * C11) Data input 339.72
C20
Length of Drill String for Calculating Critical Buckling Load in m
C22 = IF(C14 > (C7 + 2), C14 − 1,C7 + 2)
23.3
C22
Factor for Size of Fragment
C23 = (C11/75)^0.5
2.326657116
C23
Factor for Formation Density
C24 = (C16/2700)^0.5
1
C24
0.75
C25
0.858489951
C26
Density of the Formation in kg/m
Expected Penetration Rate in m/hr
Nearest Available Drill Pipe OD in mm
Factor for Roundness of Drill Cuttings C25 = IF(C9 = "Sedimentary", 0.6, 0.75) Factor for Roughness of Drill Cuttings
C26 = 1/(1 + (C11/1000)^2)
C21
(Continued)
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Table 10.6 (continued) Aspect
Formula used (if any)
Input/output
Factor for Rate of Fragmentation
C27 = 1 + 0.01 * (C17 − 15) C28 = 1 + 0.02 * (17 − C18)
1.07
C27
0.94
C28
C29 = 1 + 0.005 * C15 C30 = IF(C19 = “No”, 1, 1.1)
1
C29
1
C30
Calculated Desired Bailing Velocity in m/min
C31 = 1210 * C23 * C24 * C25 * C26 * C27 * C28 * C29 * C30
1823.164542
C31
Discharge Required from Compressor in m3/min
C32 = (PI/4) * (C11^2 − C21^2) * 70.77398847 C31/10^6
C32
Does the Compressor on the Drill Give Sufficient Air Discharge?
C33 = IF(C32 < C4, “Yes”, “No”) Yes
C33
Type of Drill Pipe (Alloy Steel or Mild Steel)
Data Input
Alloy Steel
C34
Available Thickness of the Chosen Drill Pipe in mm
Data Input
25.4
C35
Factor for Annular Space Factor for Hole Inclination Factor for Water Injection
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Cell ref.
ID of the Drill Pipe in mm
C36 = C21 − 2 * C35
288.92
C36
Moment of Inertia for the Drill Pipe in mm4
C37 = (PI/64) * (C21^4 − C36^4)
311771256.3
C37
Critical Buckling Load in kg
C38 = (3.7 * 1000 * PI^2 * 200 * C37/(1000 * C22)^2)/9.80665
427695.1119
C38
Factor of Safety Against Buckling
C39 = C38/C3
11.70484707
C39
Is Chosen Drill Pipe Appropriate for Blasthole Drilling ?
C40 = IF(C39 > 1.25, “Yes”, “No”)
Yes
C40
Actual Area Ratio (Annular Space Area/Hole Area)
C41 = (C11^2 − C21^2)/C11^2
29.98514985
C41
Outside Diameter of the Chosen Drill Pipe in mm
C42 = C21
339.72
C42
ID of the Drill Pipe in mm
C43 = C35
288.92
C43
Maximum Bit Loading in Blasthole Drilling in kg
C44 = C12
36540
C44
Minimum Desired Factor of Safety Against Buckling
Data Input
3
C45
Critical Buckling Load to Which the Drill Pipe Must Withstand in kg
C46 = C44 * C43
109620
C46
Minimum Acceptable Diameter of the Drill Pipe After Abrasive Erosion on the Basis of Buckling Load Criteria
C47 = ((64/PI) * C12 * C45 * (1000 * C22)^2 * 9.80665/(3700 * PI^2 * 200) + C36^4)^0.25
304.489999
C47
Minimum Acceptable Diameter of the Drill Pipe After Abrasive Erosion on the Basis of Desired Bailing Velocity By Using Max. Compressor Discharge
C48 = (c11^2 − (4 * c4 * 1000^3/(3.14159 * c31 * 1000)))^0.5
304.8016539
C48
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As the diameter of the drill pipe reduces the moment of inertia decreases, and the possibility of buckling of the drill pipe also increases because of decreased critical load. It is possible to find out what is the minimum outside diameter to which the selected drill pipe can be allowed to wear down, on each of the above considerations. This has been done in rows 41 to 48, where calculations reveal the following. Minimum acceptable diameter of the drill pipe after abrasive erosion on the basis of buckling load criteria = 304.49 mm. Minimum acceptable diameter of the drill pipe after abrasive erosion on the basis of desired bailing velocity by using maximum compressor discharge = 304.80 mm. It is somewhat strange that the minimum acceptable diameters are almost the same in both cases. This happens rarely. When the diameters differ the higher value should be chosen so both the criteria, i.e. adequate bailing velocity and safety against drill pipe buckling, are satisfied. Therefore the drill pipe can be allowed to wear from the original 339.72 mm diameter to 305 mm diameter in actual drilling. Once the diameter of any part on its length becomes equal to 305 mm, it should be discarded. Particular attention must be given to the desired factor of safety against buckling i.e. data input in cell C45. About 2 to 3 can be considered as adequate.
10.8
CHOOSING NOZZLES FOR TRICONE BITS
A tricone bit releases compressed air through three nozzles as illustrated in Figure 10.4. The nozzles of the bit are separate from the bit and have different bore diameters. As shown in Figure 8.10 they are shaped so as to efficiently convert air pressure into kinetic energy without turbulence loss. Nozzles can be fixed into the bit by means of a circlip like the one shown in Figure 10.4, or by means of screwing into the threads provided, or by means of a pin as shown in Figure 10.5. The advantage of use of the nozzle is that when the air jet is directed towards the blasthole bottom, the cuttings (formed by the interaction of the bit teeth with the formation) are separated from the formation very rapidly and are sent to the annulus for bailing from the blasthole. The main reason for the quick separation is that when air goes in the cracks formed in the formation it exerts drag forces on the cutting formed by the cracks developed on all the sides of the bit. Similarly the small quantity of air that fills the cracks starts expanding. It has been experienced that use of appropriate bit nozzles gives a faster penetration rate than is possible by directing the air flow towards the blasthole bottom through the central hole, shown in Figure 5.10. It is also true that the penetration rate increases with the increase in velocity of the air emerging from bit nozzles. Nozzles for a tricone bit must be selected meticulously to get the fastest penetration rate but at the same time they must also give long life to the tricone bits. This can be achieved by keeping the rotary speed as well as bit weight within the limits stated in
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Nozzle
321
Air Passage to Bearings
Air Jet Emerging from Nozzle
Figure 10.4 Jetting of the compressed air by bit nozzles.
Figure 10.5 Fitting a nozzle in tricone bit by a pin.
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chapter 5 and ensuring that 20%–30% of the compressed air volume passes through the bearing of the tricone bit. In chapter 8 the equation for the flow of compressed air was given as Q = (60/1.225) * α * ψ * p1 * A * [2/(R * T1)]0.5 where Q = Volume flow rate in m3/min α = Nozzle efficiency ψ = Flow coefficient p1 = Upstream absolute pressure in Pa T1 = Upstream absolute temperature in K A = Area of nozzle opening in m2 R = Gas constant in J/(kg ⋅ K). This equation can be taken as the basis for nozzle selection in a tricone bit. Common sizes of nozzles available for tricone bits are given in Table 10.7. If a compressor has a discharge capacity Qc at a pressure rating of pc then the compressor keeps on giving compressed air to the air flow conduits so long as the pressure loss in the conduits is less than or equal to the pressure rating. Now let us assume that for a particular diameter of blasthole and the diameter of the drill pipes used for blasthole drilling the required volume flow is Qr and the compressor provided on the rotary blasthole drill is adequate. This means Qc > Qr. Since modern screw compressors have discharge controls to control the output of the compressor, we also assume that the compressor is maintaining a discharge of only Qr to the flushing system. Therefore each nozzle should be of such size that it passes a discharge of 0.70 * Qr through it. Table 10.7 Commonly available nozzle sizes. Nozzle Size No.
7
8
9
10
11
12
13
Nozzle Dia (in.)
7/32
1/4
9/32
5/16
11/32
3/8
13/32
Nozzle Dia (mm)
5.55625
6.35
7.14375
7.9375
8.73125
9.525
10.31875
Nozzle Size No.
14
15
16
17
18
19
20
Nozzle Dia (in.)
7/16
15/32
1/2
17/32
9/16
19/32
5/8
Nozzle Dia (mm)
11.1125
11.90625 12.7
13.49375 14.2875
15.08125 15.875
Nozzle Size No.
21
22
23
24
25
26
28
Nozzle Dia (in.)
21/32
11/16
23/16
3/4
25/32
13/16
7/8
Nozzle Dia (mm)
16.66875 17.4625
18.25625 19.05
19.84375 20.6375
Nozzle Size No.
30
32
34
36
38
40
Nozzle Dia (in.) Nozzle Dia (mm)
15/16 23.8125
1 25.4
11/16 26.9875
11/8 28.575
13/16 30.1625
11/4 31.75
22.225
Some manufacturers make nozzles for their tricone bits in sizes ranging from 5 mm to 28 mm in steps of 1 mm.
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To fully utilize the pressure developed by the compressor, a pressure loss to be achieved while this volume of compressed air flows through the flushing system, can be found from following equation: pr = pc − (pd + pa) where pr = Pressure loss to be achieved through use of nozzles pc = Pressure of compressed air at air compressor pd = Pressure lost while compressed air flows to the tricone drill bit pa = Pressure required at the blasthole bottom to ensure that cutting laden air passes through the annulus. It can be said that under most circumstances (pd + pa) is less than 100 kPa. Table 10.8 gives the diameter of each of the three nozzles that together will pass the compressed air flow given in column 1. Nozzle diameter values in this table are calculated for a temperature of 35°C. This table can be used for choosing nozzles of appropriate size. The following example will clarify the method of choosing nozzles.
EXAMPLE
A rotary blasthole drill is being used for drilling blastholes of diameter 311.15 mm with a 228.6 mm drill pipe. The formations being drilled are such that bailing velocity should be about 2200 m/min for efficient hole cleaning. The compressor on the blasthole drill is capable of giving discharge up to 85 m3/min at a governed pressure of 448 kPa. Find the diameter of each of the nozzles to be used with the tricone bit. Atmospheric temperature and pressure at the worksite are 35°C and 101.325 kPa.
SOLUTION
From Table 3.1A in appendix 3 it can be seen that for getting a bailing velocity of 2200 m/min in a blasthole of diameter 311.15 being drilled with drill pipes of diameter 228.6 mm, the volume of compressed air required will be 76.99 m3/min. It is, therefore, presumed that the operator will govern the discharge of the compressor to 77 m3/min. Of this flow volume it is necessary that 70% or 53.9 m3/min flow must pass through the nozzles. From Table 10.8, for a compressor discharge of 78 m3/min and compressor pressure of 448 kPa, the appropriate diameter of nozzle is 21.97 mm. The nearest size nozzle available for the bit is 20.6375 or 22.225 mm. In almost all the cases it is better to select the nozzle with smaller diameter than larger, because a smaller nozzle builds up higher pressure but discharges less air through the nozzles and ensures more compressed air flows through the bearings of the tricone bit. Verification that the nozzle does give about 30% of air flow through bearings is needed.
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Table 10.8 Nozzle selection aid. Pressure pc
1034
862
758
689
448
414
379
350
Pressure pd + pa
100
100
100
100
100
100
100
100
Pressure pc − (pd + pa )
934
762
658
589
348
314
279
250
Volume flow
Nozzle sizes for dissipating above pressure pc − (pd + pa ), when the compressed air flow volume is as shown on left cells
m3/min
cfm
20
706.29
7.32
8.02
8.55
8.97
11.12
11.57
12.09
12.58
22
776.92
7.68
8.41
8.97
9.41
11.67
12.14
12.68
13.20
24
847.55
8.02
8.79
9.37
9.83
12.19
12.68
13.25
13.79
26
918.18
8.35
9.15
9.75
10.23
12.68
13.19
13.79
14.35
28
988.81
8.66
9.49
10.12
10.61
13.16
13.69
14.31
14.89
30
1059.44
8.97
9.82
10.48
10.99
13.62
14.17
14.81
15.41
32
1130.07
9.26
10.15
10.82
11.35
14.07
14.64
15.30
15.92
34
1200.70
9.54
10.46
11.15
11.70
14.50
15.09
15.77
16.41
36
1271.33
9.82
10.76
11.48
12.04
14.92
15.53
16.23
16.88
38
1341.96
10.09
11.06
11.79
12.37
15.33
15.95
16.67
17.35
40
1412.59
10.35
11.34
12.10
12.69
15.73
16.36
17.10
17.80
42
1483.22
10.61
11.62
12.40
13.00
16.12
16.77
17.53
18.24
44
1553.85
10.86
11.90
12.69
13.31
16.50
17.16
17.94
18.67
46
1624.47
11.10
12.16
12.97
13.61
16.87
17.55
18.34
19.09
48
1695.10
11.34
12.43
13.25
13.90
17.23
17.93
18.74
19.50
50
1765.73
11.57
12.68
13.52
14.18
17.59
18.30
19.12
19.90
52
1836.36
11.80
12.93
13.79
14.47
17.94
18.66
19.50
20.29
54
1906.99
12.03
13.18
14.05
14.74
18.28
19.01
19.87
20.68
56
1977.62
12.25
13.42
14.31
15.01
18.61
19.36
20.24
21.06
58
2048.25
12.47
13.66
14.57
15.28
18.94
19.71
20.59
21.43
60
2118.88
12.68
13.89
14.81
15.54
19.27
20.04
20.95
21.80
62
2189.51
12.89
14.12
15.06
15.80
19.59
20.37
21.29
22.16
64
2260.14
13.09
14.35
15.30
16.05
19.90
20.70
21.63
22.51
66
2330.77
13.30
14.57
15.54
16.30
20.21
21.02
21.97
22.86
68
2401.40
13.50
14.79
15.77
16.54
20.51
21.34
22.30
23.20
70
2472.03
13.70
15.01
16.00
16.78
20.81
21.65
22.63
23.54
72
2542.66
13.89
15.22
16.23
17.02
21.11
21.96
22.95
23.88
74
2613.29
14.08
15.43
16.45
17.26
21.40
22.26
23.26
24.21
76
2683.91
14.27
15.64
16.67
17.49
21.69
22.56
23.58
24.53
78
2754.54
14.46
15.84
16.89
17.72
21.97
22.85
23.88
24.85
80
2825.17
14.64
16.04
17.11
17.94
22.25
23.14
24.19
25.17
(Continued)
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Table 10.8 (Continued) Pressure pc
1034
862
758
689
448
414
379
350
Pressure pd + pa
100
100
100
100
100
100
100
100
Pressure pc − (pd + pa )
934
762
658
589
348
314
279
250
Volume Flow cfm
Nozzle sizes for dissipating above pressure pc − (pd + pa ), when the compressed air flow volume is as shown on left cells
82
2895.80
14.82
16.24
17.32
18.17
22.52
23.43
24.49
25.48
84
2966.43
15.00
16.44
17.53
18.39
22.80
23.71
24.78
25.79
86
3037.06
15.18
16.63
17.74
18.60
23.07
24.00
25.08
26.10
88
3107.69
15.36
16.83
17.94
18.82
23.33
24.27
25.37
26.40
90
3178.32
15.53
17.02
18.14
19.03
23.60
24.55
25.65
26.70
92
3248.95
15.70
17.20
18.34
19.24
23.86
24.82
25.94
26.99
94
3319.58
15.87
17.39
18.54
19.45
24.12
25.09
26.22
27.28
96
3390.21
16.04
17.57
18.74
19.66
24.37
25.35
26.50
27.57
m3/min
98
3460.84
16.20
17.76
18.93
19.86
24.62
25.61
26.77
27.86
100
3531.47
16.37
17.94
19.13
20.06
24.87
25.88
27.04
28.14
103
3637.41
16.61
18.20
19.41
20.36
25.24
26.26
27.45
28.56
106
3743.35
16.85
18.47
19.69
20.65
25.61
26.64
27.84
28.97
110
3884.61
17.17
18.81
20.06
21.04
26.09
27.14
28.36
29.51
For this purpose the nozzle flow equation given in chapter 8 has to be used. In this case we have p = 348 kPa α = 0.95 ψ = 0.4841 R = 287.1 J/kg ⋅ K p1 = 348000 + 101325 = 449325 Pa T1 = 15 + 273.16 = 288.16°K A = π * (22.25/1000)2/4 = 0.00033451 m2 Hence M = 0.95 * 0.00033451 * 0.4841 * 449325 * 2/(287.1 * 288.16)]0.5 or M = 0.33986 kg/s or Q = 16.6464 m3/min Therefore, discharge from three nozzles will be 49.93 m3/min and balance 77 − 49.93 = 27.07 m3/min or 27.07/77 = 35.15% air will flow through the bearings. If under the same conditions, nozzles of diameter 22.225 were chosen then the discharge through each nozzle would have been 19.3058 m3/min and the three nozzles would have discharged 57.92 m3/min. Under such conditions the flow through the bearings would have been 77 − 57.92 = 19.08 m3/min or 19.08/77 = 24.78%.
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For ensuring 30% discharge through the bearings it would have been necessary to control the discharge from the compressor at 57.92/0.7 = 82.75 m3/min. This would have resulted in a large increase in the bailing velocity to about 2380 m/min. As stated earlier, drill manufacturers carry out air tests to verify that bearings of the tricone bit are getting adequate air supply. These tests and the method of selecting nozzles adopted by them is elaborated in an appendix.
10.9
CHOOSING THE RIGHT COMPRESSOR
The compressor is an integral part of a rotary blasthole drill. It cannot be changed without incurring very heavy expenses. It must, therefore be appropriately chosen while purchasing the blasthole drill itself. All the manufacturers provide the most appropriate compressor on their blasthole drills for normal working conditions. But when the blasthole drill is to operate at altitude or for drilling in a very heavy mineral mass, such standard compressors prove insufficient and an optional compressor has to be chosen. As this topic is more related to altitude operation it is discussed in the next chapter that deals with operation of the blasthole drill at high altitude.
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Chapter 11
Effect of altitude and severe weather
11.1
INTRODUCTION
Blasthole drilling is carried out by using a blasthole drill and many other accessories described in earlier chapters of this book. The working of some of these components does depend upon the atmospheric parameters and other environmental conditions that change with altitude. The height of a workplace from mean sea level is called altitude. As many mines are located above 1500 m altitude and some are as high as 4500 m altitude, it is important to know the effects on such components. There is a well established relation between most of the environmental parameters and the altitude of the place. Parameters such as temperature, pressure and density of the atmospheric air do change with altitude. Humidity is not dependent upon altitude but the surroundings. These changes affect the working of many items used in the drilling process. In a very strictest sense, even the process of rock fracture must also be affected due to the changes of the atmospheric parameters mentioned above. However, the effect is so negligible that it is not even researched. Apart from the above changes, the worksite often experiences severe operating conditions such as stormy winds, extreme cold, extreme heat, heavy rainfall, heavy snowfall etc. These too affect working of the drill and other components. This chapter is dedicated to giving the essential knowledge about the effect of altitude and severe weather on the blasthole drilling process.
11.2
ATMOSPHERE
The atmosphere surrounds the earth and extends to about 3000 km in height. The mass of all the matter in this zone is estimated to be about 151.7 × 106 kg. The matter in the atmosphere is collectively called air and consists of many gases as shown in Table 11.1. The percentage of component gases marked with an asterisk in this table varies to some extent due to several reasons. In addition, there are organic and inorganic impurities in the form of dust, water vapor etc. The proportion of impurities and water vapor greatly depends upon the surroundings of the place. The atmosphere itself is subdivided into five zones viz. troposphere, stratosphere, mesosphere, ionosphere and exosphere as shown in Figure 11.1. Of these, the troposphere extends up to an altitude of 11 km. As all the drilling sites in the world are located in this zone we need be concerned with this zone only.
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Table 11.1 Composition of air in the atmosphere. Component
% by vol
% by mass
Nitrogen
78.08
75.71
Oxygen
20.95
23.15
Argon
0.93
1.28
Carbon Dioxide*
0.03
0.05
Neon
0.0018
0.00125
Helium
0.00052
0.00007
1.5 * 10
−04
9.4 * 10−05
Krypton
1.1 * 10
−04
2.9 * 10−04
Carbon Monoxide*
1.0 * 10−04
2.0 * 10−04
Nitrous Oxide*
5.0 * 10
−05
8.0 * 10−05
Hydrogen*
5.0 * 10−05
3.5 * 10−05
Ozone*
4.0 * 10
−05
7.0 * 10−06
Xenon
8.0 * 10−06
3.6 * 10−06
Nitrogen Dioxide
1.0 * 10
−07
2.0 * 10−07
Iodine
2.0 * 10−11
1.0 * 10−10
Radon
−18
5.0 * 10−17
Methane
6.0 * 10
Exosphere - From 410 to nearly 3000 km
Ionosphere - From 90 to 410 km
Mesosphere - From 35 to 90 km Stratosphere - From 11 to 35 km
Troposphere - From 0 to 11 km
Figure 11.1 Different zones of atmosphere.
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Effect of altitude and severe weather
11.2.1
329
Effects of altitude on atmosphere
Pressure, temperature, as well as density of the atmospheric air, slowly reduces with increasing altitude. Changes in these parameters with respect to altitude are given in one of the appendices at the end of this book. The data presented in the appendix is applicable for latitude of 40°. The atmospheric pressure and air density remain virtually unaffected by latitude but the temperature is affected considerably. Usually precise indication of atmospheric temperature is not required for evaluation of any factor in blasthole drilling, however, and the following rule of thumb can help in this context. To estimate mean annual temperature of a place located between latitude 0° (i.e. equator) and latitude 75° (generally the northernmost or the southernmost inhabited place), add 0.7°C in the 40° latitude temperature for decrease of each 1° latitude or subtract 0.7°C from the 40° latitude temperature for increase of each 1° latitude.
11.2.2
Effects of humidity on atmosphere
Up to a certain limit, water vapor can get mixed with air and behave like a gas i.e. without condensation. The presence of water vapor in the air is called humidity. The amount of vapor in the air is most simply expressed as its weight per unit volume. There is a limit to the amount of water vapor that a given volume of air can hold. This limiting amount of water vapor makes the air saturated and any additional water vapor in the air, if present, gets condensed immediately. At higher temperatures air can hold more water vapor than at lower temperatures. Atmospheric pressure does not have any bearing on the saturation of air so long as the temperature is constant. Absolute humidity is measured as grams of water vapor per cubic meter of air. Measurements of humidity in this absolute manner, however, are very difficult hence the term relative humidity is more commonly used. For different temperatures, the absolute humidity required for saturation of the air with water vapor, are well established by means of laboratory tests. In general, air is not saturated. It contains only a fraction of the possible water vapor. This fraction, expressed in percentage, is called relative humidity. In reality the above is an oversimplified version of absolute and relative humidity, and hence a little imprecise, because these quantities are measured on the basis of pressure rather than weight. The vapor pressure of saturated air for different temperatures is given in Table 11.2. The actual water content of moist air at different temperatures and relative humidity is given in Figure 11.2. Table 11.2 Vapor pressure of saturated air for different temperatures.
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Temp in °C
Vapor pr. in kPa
Temp in °C
Vapor pr. in kPa
−10 −5 0 5 10 15 20
0.260 0.402 0.611 0.872 1.227 1.705 2.337
25 30 35 40 45 50
3.167 4.263 5.624 7.378 9.580 12.300
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11.3
EFFECT OF SCARCE AIR IN DRILLING
In rotary blasthole drilling atmospheric air is required for three functions as follows: 1 2 3
The atmospheric air gives sufficient oxygen to the burning processes inside the cylinders of a diesel engine. The atmospheric air absorbs heat from various components that generate or carry generated heat during the drilling process. The atmospheric air, in compressed form, is circulated through the blasthole for the purpose of cutting removal by flushing.
In the following sections the influence of changes in atmospheric parameters on various components used on drilling machines and in the drilling process is elaborated.
11.3.1
Diesel engines
A diesel engine generates power from the combustion of fuel within its cylinders. The combustion requires oxygen that is obtained by the engine from the atmospheric air. If, for any reasons, the oxygen available from the atmosphere is not sufficient then less quantity of fuel is burnt and the diesel engine generates less power. Every diesel engine has a rated power. The rated power is based on the tests carried out on the diesel engine on a test bed in a laboratory. Temperature and humidity in the laboratory are controlled within certain limits. Since the laboratory is located at a known altitude, the other atmospheric parameters are well established. Diesel engine manufacturers rate their engines for some well defined conditions as specified by one of the various standards. The standards also define the factors of deration to be applied to the engine power output when the engine is used under different atmospheric conditions. The best way, therefore, is to use the factors stipulated by the standards for the purpose of engine derating. Table 11.3 gives the standard conditions used for the purpose of engine derating by some of the well known standards. There are two distinct types of diesel engines viz. naturally aspirated and turbocharged. In a naturally aspirated engine the atmospheric air is simply sucked in through the engine valves. The pressure of the air, before it is sucked, is the same as the atmospheric pressure. In turbo-charged engines, however, there is a turbo charger which, by its fan, sucks the air to the engine suction mechanism. Naturally this fan acts as a very low pressure compressor, and therefore the pressure of the air before it is sucked in by the engine suction mechanism is somewhat higher than the atmospheric pressure. This results in more air being sucked into a turbo-charged engine as compared to an identical naturally aspirated engine. More often, the power output of a turbo-charged engine is purposely restricted by the amount of fuel pumped into the engine, rather than the volume of air sucked into it. Some engines can be equipped with special turbo chargers that enable engine operations in excess of 3000 m altitude without power deration. Due to these factors, a turbo charged engine is required to be derated to a much lesser extent than a equivalent naturally aspirated diesel engine.
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Relative Humidity %
10
50
20
30
40
60
80
100
40
Temperature in °C
30
20
10
0
−10
−20
1
2
4
6
8
10
20
40
3
Water Contents in g/m
Figure 11.2 Water content of air for different atmospheric temperature and relative humidity values.
60
80
100
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Table 11.3 Standard conditions for engine power output as defined by different standards. Standard
Pressure kPa
Temperature °C
Rel. humidity %
SAE (American) BS (British) DIN (German) ISO (International) IS (Indian)
100 100 100 100 100
25 27 27 27 27
0 60 60 60 60
Table 11.4 Rules of thumb for engine power deration.
Engine type Naturally Aspirated Turbo charged
Altitude in m up to which no deration is required 500 m Variable between 2000 m to 3000 m
Deration in % for each 100 m increase in altitude
Deration in % for each 10
1 1
3.6 5.4
A rule of thumb for engine power deration is given in Table 11.4. It is to be specifically noted that turbo-charged engines need not be derated up to an altitude of 2000 m, or 3000 m in some cases. As far as naturally aspirated engines are concerned a more precise formula for engine power deration, developed on a technological basis, is as follows: R = [(PO − RO ∗ PHO)/(PT − RT ∗ PHT)] ∗ [(TT + 273.15)/(TO + 273.15)]0.5 where the variables are: a
Values used for engine power rating PT = Absolute pressure in kPa. RT = Relative humidity in percent. TT = Atmospheric temperature in °C. PHT = Vapor pressure at saturation coexistent with temperature TT, in kPa.
b
Values for workplace of engine operation PO = Atmospheric pressure in kPa. RO = Relative humidity in percent. TO = Atmospheric temperature in °C. PHO = Vapor pressure at saturation coexistent to temperature TO, in kPa.
Values of PT and PO can be found from the appendix that gives properties of atmospheric air and PHT and PHO from Table 11.2.
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Effect of altitude and severe weather
333
EXAMPLE
A naturally-aspirated diesel engine has a continuous rating of 400 kW as per a standard that specifies ratings to be given at sea level under 20°C air temperature and 60% relative humidity. Find the rated power of the engine when it is operated at an altitude of 1400 m where prevailing temperature and relative humidity are 35°C and 80% respectively. SOLUTION
From the data of engine rating conditions it can be concluded that PT = 101.325 kPa RT = 0.6 as given in the example data TT = 20°C. as given in the example data PHT = 2.337 kPa (from Table 11.2) From the details of the operating conditions it is evident that PO = 85.6 kPa. RO = 0.8 as given in the example data TO = 35°C. as given in the example data PHO = 5.624 kPa (from Table 11.2) The above data yields the deration factor to be R = [(85.6 − 0.8 ∗ 5.624)/(101.325 − 0.6 ∗ 2.337)] ∗ [(20 + 273.150/(35 + 273.15)]0.5 or R = 0.7914 From the deration factor we get the power output at the altitude and the other conditions as P = 0.7914 ∗ 400 = 316.56 kW
11.3.2
Electric motors and transformers
Electric motors generate heat while converting the electrical energy to mechanical energy. In a similar manner, a transformer also generates heat while the power supply is transformed from one type to the other. The dissipation of heat from such units to the surroundings is quite a complex process, as all the three modes of heat transfer viz. conduction, convection and radiation are involved. Electric motors and transformers are rated on the basis of certain temperature rise permitted in the machine over the ambient temperature, i.e. temperature of the surrounding atmospheric air, measured by a thermometer at a distance of 2 m or more from the machine. Distance and temperature rises differ for different standards.
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In actual practice the motors and transformers dissipate about 60% of the heat through convection and the balance by conduction and radiation. At higher altitude, where the density of air is less than that at sea level, the dissipation of heat through convection is considerably affected, whereas there is no effect on the conduction and radiation of heat. The deration of electric motors and transformers depends upon many factors. Hence, for a particular set of environmental conditions, the best way of knowing the deration is to refer the problem to the motor or transformer manufacturers because only they know the values of various parameters needed for the solution. In cases where the problem is required to be solved without referring to the manufacturer, for whatever reason, the deration on account of altitude should be obtained from Figure 11.3 which illustrates the deration factors for various altitudes. The graph in Figure 11.3 is based on cooling air temperature of 40°C or less. Occasionally mines are located at places where the temperature of cooling air itself is more than 40°C. In such cases the deration in the deration factor, of 1% for each 1°C difference between the air temperature and 40°C datum, may be taken and the new deration factor thus obtained may be applied in addition to the deration factor on account of altitude. The need for having both the deration factors, viz. one on account of altitude and the other on account of temperature, arises only under unlikely conditions because at altitudes beyond 1000 m, air temperature hardly ever rises beyond 40°C. The following example, therefore, is only for the purpose of illustration and understanding. EXAMPLE
A 500 kW electric motor, designed for use up to an altitude of 1000 m, is intended to be used at a place having 1900 m altitude where environmental temperatures occasionally rise to about 48°C. Find the power output of this motor at such a place. SOLUTION
From the graph in Figure 11.3 the deration factor for 1900 m altitude works out to 0.955. Since the temperature rises up to 48°C, this deration factor should be reduced further to an extent of (100 − 8)/100 = 0.92.
Deration Factor
1.00 0.95 0.90 0.85 0.80
0
1000
2000 Altitude in m
3000
4000
Figure 11.3 Deration factor for electric motors and transformers.
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Effect of altitude and severe weather
335
This gives the corrected deration factor to be 0.955 * 0.92 = 0.8786 and the derated power of the motor works out to 500 * 0.8786 = 439.3 kW. Whatever has been said above for electric motors also applies to electric transformers because the cooling of transformers also is in the same pattern as that of an electric motor.
11.3.3
Hydraulic system
As has been seen earlier, the hydraulic system of a blasthole drill consists of many components such as hydraulic pumps, hydraulic motors, hydraulic cylinders, hose pipes, valves etc. At higher altitude the barometric pressure is less than that at sea level. If the hydraulic system of the drill is a closed loop system then there is no effect as far as the power transmission capability of the system is concerned, but the system is more susceptible to bubble formation. This can be taken care of by providing a somewhat larger bleeder tank, if necessary. If the hydraulic system is open type then the positive suction pressure, which is generally present at the hydraulic pumps, reduces. This makes it necessary to provide a pressurized hydraulic tank or place the tank at higher level with respect to the hydraulic pumps. It may be noted that up to 1500 m altitude the effects mentioned are not very critical, and hence in many cases no remedial measures may be necessary up to this altitude. Air also takes away the heat generated by the hydraulic system through the heat exchanger. Since the density of the air is less at higher altitude, the dissipation of the heat is also slower. But side by side the air temperature is also less. As these two counter-balance each other, no remedial measures are required unless the altitude of operation exceeds 1500 m. The remedial measures are in terms of providing an hydraulic oil cooler of larger size. The best course of action to be taken by a user is to consult the manufacturer prior to the purchase of the blasthole drill since only they are aware of various factors considered while designing the heat exchanger of the hydraulic system. One has to take into account the increase in viscosity of the hydraulic fluid at higher altitude due to the decreased temperatures. For this the solution is very simple. Just change the hydraulic fluid with a thinner fluid that flows easily. Usually this information is readily available in the service manuals of the drill. In absence of such information the manufacturer should be consulted. Oil companies make such information available through their product literature. In fact, most of the drill manufacturers just repeat this information to their customers.
11.3.4
Air compressors
Low atmospheric pressures and low temperatures experienced at high altitudes affect the compressors on rotary blasthole drills to a very great extent. A compressor delivers less compressed air as the altitude becomes higher. Actually the power required for the compressor also decreases as per altitude. However, the deration of engine power output is at a higher rate than the reduced power rate for compression. Hence it is necessary to adopt an engine with higher power.
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As stated earlier the compressors used on rotary blasthole drills are positive displacement type, i.e. screw or vane compressors. Single stage compressors that develop sufficiently high pressure for rotary drilling are usually chosen for rotary blasthole drills. Table 11.5 gives values for discharge correction factors and percentage reduction in power requirements for compressing air to 689 kPa (100 psi) at different altitudes for oil-flooded single stage screw compressors as well as oil-free two stage screw compressors. The table is in imperial units but it contains factors for ACFM, SCFM, power reduction etc. Hence, the values of these factors can easily be interpolated either by Newton’s Method for Interpolation (preferred) or the Linear Interpolation Method. Here again, the best guides for tackling the situations are the compressor manufacturers who have all the design data for their compressors. In practice, when a user requires a blasthole drill to operate at a certain high altitude and informs the manufacturer accordingly, the manufacturer usually proposes some changes in the configuration of the blasthole drill. As far as the compressor is concerned these changes are: 1 2
Using a larger size compressor instead of the one that is required to achieve the necessary flushing parameters at sea level. Providing an engine or an electric motor with a larger power output as a compressor drive. An elaborate example in this regard is given elsewhere in this book.
11.3.5
Blasthole flushing
For flushing the blasthole, compressed air moves from the compressor discharge port to the drill bit through metal pipes on the blasthole drill and then the drill string components up to the bit where it is discharged in the blasthole. The pressure loss experienced by the compressed air while it flows through the accessories does depend upon the compressor discharge pressure, but does not quite depend upon the atmospheric parameters. Therefore, the pressure of the compressed air just before the drill bit is the same so long as the compressor discharge pressure is unaltered. This is true for any atmospheric pressure. Now, consider that pressures of the compressed air prevailing in the flushing path are as given in Table 11.6. The above table gives the pressure values for otherwise identical conditions of blasthole drill, drill string components, formation, drill bit, blasthole depth etc. As the pressure P0 is set at the compressor output valve, it does not change. Therefore, p0 = P0. Since the operation at sea level and at altitude does not change the components in the flow path on the blasthole drill or the components in the drill string, the pressure loss will remain the same. Therefore, (P0 − P1) = (p0 − p1). But since P0 = p0, p1 = P1, we also have p3 < P3. The pressure loss experienced in the annulus will also remain same, as the flow path is unchanged i.e. P2 − P3 = p2 − p3. Therefore, p2 < P2. The above means P1 − P2, i.e. pressure loss in the nozzles of the bit, increases because P1 − P2 < p1 − p2.
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Table 11.5 Effect of altitude on discharge capacity and power requirement of screw compressors.
Atmospheric Altitude ft
Pressure psia
Temperature deg F
Relative humidity %
0 1000 2000 3000 4000 5000 6000 7000 8000 9000 10000
14.7 14.18 13.66 13.17 12.69 12.23 11.78 11.34 10.90 10.49 10.10
68 64 61 57 54 50 47 43 39 32 32
36 36 36 36 36 36 36 36 36 36 36
Oil flooded single stage rotary screw compressors
Oil free two stage rotary screw compressor
Correction factor
Correction factor
Inlet pressure psia
ACFM
SCFM
0.3389 0.2995 0.2641 0.2325 0.2042 0.1791 0.1567 0.1368 0.1191 0.1035 0.0898
1.0000 0.9984 0.9969 0.9954 0.9940 0.9926 0.9912 0.9899 0.9887 0.9874 0.9862
1.0000 0.9718 0.9438 0.9164 0.8899 0.8636 0.8383 0.8133 0.7878 0.7638 0.7408
Total
Reduction in power %
ACFM
SCFM
Total
Reduction in power %
1.0000 0.9703 0.9409 0.9122 0.8845 0.8572 0.8309 0.8124 0.7789 0.7542 0.7305
0.0 1.8 3.5 5.2 6.9 8.5 10.1 11.6 13.1 14.6 16.1
1.0000 0.9998 0.9995 0.9991 0.9988 0.9985 0.9982 0.9979 0.9976 0.9974 0.9971
1.0000 0.9718 0.9438 0.9164 0.8899 0.8636 0.8383 0.8133 0.7878 0.7638 0.7408
1.0000 0.9716 0.9433 0.9156 0.8888 0.8623 0.8368 0.8116 0.7860 0.7618 0.7386
0.0 1.1 2.9 4.7 6.3 8.0 9.5 11.1 12.6 14.0 15.4
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Table 11.6 Pressure locations in the compressed air flow path. Location of air pressure
At sea level
At altitude
Pressure at the Compressor Outlet Before the Tricone Bit After the Tricone Bit i.e. At Blasthole Bottom After the Blasthole i.e. Atmospheric Pressure
P0 P1 P2 P3
p0 p1 p2 p3
The increase in pressure loss is due to the fact that air is ejected through the flow nozzles at a higher speed. This actually means that at higher altitude the blasthole flushing improves. Besides the above, when the compressed air passes through the flow path, it cools to a lower temperature at higher altitude than at sea level because the atmospheric air temperature is lower at higher altitudes. To whatever little extent it may be, this low temperature compressed air must be improving the heat dissipation at the bit bearing and thus resulting in the increased bit life.
11.3.6
Drill lubrication
Temperature and pressure reduce at high altitude. In blasthole drills the pumps used for either oil injection into compressed air for bit lubrication, or for supplying lubricant to different parts, do not have much suction head. Therefore, lower atmospheric pressure at high altitudes does not affect the pumping process to a significant extent. However, in colder temperatures at high altitude lubricating oils become thicker and do not flow easily. Many multi-viscosity synthetic lubricants that suit the temperatures year-round are available from lubricant manufacturers.
11.4
EFFECT OF SEVERE WEATHER
Severe weather conditions are often experienced at a blasthole drilling worksite even when it is not located at very high altitudes. Typically the conditions are: stormy winds, heavy rainfall, heavy snowfall and extremely cold or hot weather. There are no rigorous definitions for distinguishing the conditions as normal or severe, but some rough data presented in Table 11.7 will serve the purpose. The following elaboration on how the high altitude or severe weather affects the drill and drilling process are intended for imparting preliminary knowledge. If a blasthole drill is expected to be operated at a certain location, the user must find out the altitude and weather conditions likely to be experienced at the worksite throughout the year. Such information should be provided to the drill manufacturer right when the drill is being selected for a purchase. A manufacturer is best equipped to suggest the best alternatives to combat the weather conditions.
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Table 11.7 Severe weather criteria. Aspect
Normal conditions
Severe conditions
Winds
Velocity 50 km/h
Rainfall
100 mm/day
Cold Weather
Temperature >0°C
Temperature 180 MPa
200 250 311
600 1200 2050
150 300 625
50 125 270
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Table 26.4 Blasthole diameters based on shovel bucket capacities. Bucket capacity of the shovel (m3)
Hole diameter range in mm
4.5 7.5 9.17 11.5 15.3 20 35 50
76–127 127–215 171–250 200–270 229–311 250–349 270–381 311–445
Blasthole diameter is also influenced by the bench height. This will be elaborated later in the appropriate section. 26.4.4.2
Desired fragmentation
Closely spaced small diameter blastholes certainly give smaller fragment size but large diameter blastholes do not give proportionally larger fragment size. What the distribution of particle size will be after a blast can be predicted with reasonable accuracy by many mathematical models developed for the purpose. One such model is the Kuz-Ram model. Table 26.5 presents the prediction of fragment size distribution of blasts for 203 and 311 mm diameter blastholes, where all other factors remain the same except those mentioned in the table. The data in the table prove both the points mentioned in the previous paragraph. Actually such mathematical models – even though they indicate a relationship between fragment size distribution and blasthole diameter – are not useful for determining blasthole diameters, but can be used for verification that the fragment size will be good enough for the loading and hauling equipment when a particular diameter is selected. 26.4.4.3
Properties of rock specimen
In the absence of a more rigorous approach as explained in a previous chapter, the powder factors to be chosen for some of the commonly occurring rocks are given in Table 26.6. If blastholes are charged with more than the optimum amount of explosive, the rock mass gets fragmented to small size. With this, the need for crushing decreases but the overall cost increases. 26.4.4.4
Presence of geological structures
Properties of the rock mass also need to be taken into account while fixing the diameter of blastholes. The number and placement of joints in the rock mass may
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Table 26.5 Fragment size distribution as per Kuz-Ram model. Percent passing for blasthole with following details Fragment size (m)
Hole dia. = 203 mm Burden = 5.075 m Spacing = 6.343 m
Hole dia. = 311 mm Burden = 7.775 m Spacing = 9.718 m
0.00 0.05 0.10 0.15 0.20 0.25 0.30 0.35 0.40 0.45 0.50 0.55 0.60 0.65 0.70 0.75 0.80 0.85 0.90 0.95 1.00 1.05 1.10
0.0% 5.2% 17.6% 33.6% 50.2% 65.1% 77.1% 85.9% 91.9% 95.6% 97.7% 98.9% 99.5% 99.8% 99.9% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0%
0.0% 4.0% 13.7% 26.9% 41.4% 55.5% 67.9% 78.0% 85.7% 91.1% 94.7% 97.0% 98.4% 99.2% 99.6% 99.8% 99.9% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0%
Table 26.6 Powder factors for different rocks.
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Rock category
UCS range MPa
Very Soft Soft
about 50 50–100
Medium
100–200
Hard
200–300
Very Hard and Hard plus Tough
More than 300
Rock name Coal, Potash, Soft Shale, Marl Sandstone, Hard Shale, Limestone, Slate, Conglomerate Schist, Hornfels, Hard Limestone, Marble, Serpentinite, Dolomite Andesite, Dolerite, Hard Iron Ore, Basalt, Granite, Taconite, Skern, Quartzite, Hard Basalt
Powder factor in kg/m3 0.15–0.25 0.25–0.40
0.4–0.60 0.60–0.70 0.70–1.00
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necessitate reduction of blasthole diameter, fixed on the basis of desired production rate. A rock mass in a mining bench is rarely monolithic. It is usually inter-spread with joints. A monolithic block of the rock exists only within the boundaries of such joints. The process of formation of cracks and subsequent fragmentation of rock mass starts from the blasthole and continues undisturbed till it encounters some discontinuity in the form of joint, fault, fold, unconformity etc. Beyond such a discontinuity the formation of cracks and subsequent fragmentation is reduced considerably. Suppose the structure and spacing of joints in a bench is like the one shown in Figure 26.9 A, and if large diameter blastholes with larger spacing as shown in the figure are drilled and blasted, then two of the blocks do not have any blasthole within them. This will result in poor fragmentation within those two blocks. However, if smaller blastholes with smaller spacing, as shown in Figure 26.9B, are chosen then each of the blocks has at least one blasthole within it and will thus result in good fragmentation in all the blocks. Reduction of diameter of blastholes beyond a certain point can also be counter-productive. This happens because as the diameter of the blastholes is reduced, the number of blastholes required to be drilled increases disproportionately. This increases the cost of drilling. Further, the cost of explosives, cost of accessories, and cost of charging the blastholes increase for a larger number of blastholes. In very small diameter blastholes it is not possible to use inexpensive explosives like ANFO.
Bench Face Spacing
Block Without Blasthole
Burden
Block Without Blasthole
A - Rectangular Pattern Comprising of Fewer Large Blastholes Bench Face
Spacing
Burden
B - Rectangular Pattern Comprising of Many Small Blastholes
Figure 26.9 Dependency of blasthole diameter on joint structure.
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26.4.5
609
Blasthole depth
Really, this should be called length of blasthole. However, in surface mines, as most of the blastholes are in a vertical direction, the term is colloquially called depth. The length of the blasthole includes subdrilling. It depends upon height of bench, subdrilling and angle of inclination of the blasthole. For the purpose of material removal after blasting, one of the three alternatives are used. 1 2 3
Combination of Shovel and Dumper Walking Dragline. Combination of Wheel Loader and Dumper
The manner in which these combinations work differs, the most appropriate method of drilling and blasting also differs for each case. 26.4.5.1
Combination of shovel and dumper
Shovel-dumper combination is chosen in most of the metal mines, where the ore body is irregularly placed as shown in Figure 1.3 or on coal mines where the ore body is not in a horizontal or nearly horizontal layer. As shown in Figure 26.10, a shovel starts scraping the blasted rock into its bucket at the level of the ground on which it stands and moves the bucket forward while raising it till the maximum attainable height H is reached, or the bucket is completely filled with blasted rock. It then slews into an angle of about 120° to 135°, with the bucket filled with the broken rock and held at a height of H1, which is higher than the height of the dumper body. When the bucket comes into position above the dumper body, the contents of the bucket are unloaded in dumper body by opening the bottom plate of the bucket. After dumping the broken rock, the shovel turns back again and starts scraping the broken rock face. A complete cycle of excavation and dumping of blasted rock takes average time of about 22 to 26 seconds. The maximum attainable height of scraping, called cutting height, for a shovel depends upon the machine model. For most of the shovels used in large surface mines the cutting height lies between 10 m to 18 m.
H
Scrape Surface H1
Figure 26.10 Working of an electric shovel.
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The height of the bench and so the depth of the blasthole is decided on the basis of maximum cutting height of the shovel, after giving due consideration to the subdrilling needed and the inclination of the blasthole. A formula, as given below, was proposed for fixing bench heights in metal mining that uses shovel-dumper combination. H = 10 + 0.57 * (Cc − 6) where H = Bench height in m Cc = Capacity of shovel bucket in m3 However, this formula should be used with great caution because it holds good only for a few small and medium size shovels. Table 26.7 gives some details of the bucket capacities and heights of cut for some well known rope shovels. 26.4.5.2
Walking dragline
Walking draglines are used in coal mines. They do not need any hauling equipment because with their long booms they can scrape blasted overburden and throw it a long distance away. The working of a walking dragline is shown in Figure 26.11. While removing the overburden lying above a horizontal or near-horizontal coal bed, a dragline stands on the blasted overburden and throws its bucket at a long distance on the top of the exposed coal bed. It then starts pulling the bucket towards itself. In this action the bucket scrapes the overburden and gets filled. The filled bucket is then lifted well above the ground and the dragline turns into an angle of about 90° and dumps the material into a heap on the side. As the draglines are far bigger than shovels, the depth to which they can excavate is much greater. Boom length, bucket capacity, depth of excavation for some well known walking draglines is given in Table 26.8. In some coal mines the coal beds are at a shallow depth and the thickness of the overburden layer is small. In such cases the bench height and depth of blastholes is decided on the basis of the thickness of the layer and inclination of the blastholes rather than the depth of excavation of the dragline.
Table 26.7 Heights of cut for some electric mining shovels.
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Make
Model
Bu. capacity m3
Bu. capacity MT
Height of cut (m)
Bucyrus Intl. Bucyrus Intl. Bucyrus Intl. Bucyrus Intl. P&H Mining P&H Mining P&H Mining
295HD 295HR 395 495 2300XPC 2800XPC 4100XPC
21.3 25.5 35.7 61.2 25.5 35.7 61.2
38 45 63.5 109 45.4 63.5 108.9
18.4 13.5 16.6 16.8
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Design of a surface blast
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Lifting Rope Slewing Axis of Dragline Scrape Surface Pulling Rope H Coal Bed
Blasted Overburden
Figure 26.11 Working of a walking dragline.
Table 26.8 Depth of excavation for some walking draglines.
Make
Model
Bucyrus Intl.
680 W
Bucyrus Intl.
W2000
Bucyrus Intl.
8050
Bucyrus Intl.
8200
Bucyrus Intl.
8750
P&H Mining
9010C
P&H Mining
9020C
P&H Mining
9030C
26.4.5.3
Boom length (m)
Bucket capacity (m3)
58 (min) 90 (max) 75 (min) 101 (max) 99 (min) 99 (max) 84 (min) 122 (max) 109 (min) 132 (max) 100 (min) 107.7 (max) 88.4 (min) 123.4 (max) 99.1 (min) 129.5 (max)
24 12 34 24 61 45 84 51 116 76 57 38 89 54 117 90
Excavation depth (m)
50.5 74.5
56.4 68.3 54.8 85.7 56.5 76.7
Combination of wheel loader and dumper
In coal mines the layers of coal are soft and distinctly separate from the overburden. Therefore, wheel loaders are sometimes used as equipment for loading coal into dumpers in coal mines.
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A wheel loader works differently than a shovel or a dragline. With its bucket in the lowermost horizontal position, as shown on the right side in Figure 26.12, it moves forward. With this movement the coal or other loose material gets filled into its bucket. When sufficient volume of loose material is filled into the bucket it turns the bucket in upright position and simultaneously starts lifting the bucket in a backward movement on a curve. By the time it goes sufficiently on the back side, the bucket is in its highest position. The wheel loader again starts moving on a curve to approach the dumper placed on its side. When it has moved sufficiently towards the dumper and its bucket is in position just above the body of the dumper, the wheel loader tilts its bucket so as to unload the material in the dumper body. This completes one cycle and the dumper is again ready to start a second cycle. Unlike the shovel, a wheel loader does not scrape material. It has relatively low height and its bucket raises only to sufficiently high level to load a dumper of matching size. A wheel loader finds it a little more difficult to load material from a steeply sloping heap but is at ease while loading material from a low height, well spread heap. It is, therefore, well suited for handling coal. Bench height and, therefore blasthole depth, is limited in the case of coal beds where a wheel loader has to work. The bench heights selected for blastholes of different diameters varies considerably. Usually followed trends in this regard are as expressed through Table 26.9. If blastholes in a bench are vertical with no subdrilling, their blasting gives rise to a large stump as shown in left hand drawing in Figure 26.13. That is the reason why subdrilling to some depth below the intended ground surface created after blasting becomes essential. Subdrilling value i.e. length of subdrilling, is usually kept to 8 to 12 times the blasthole diameter.
26.4.6
Blasthole inclination
Often blastholes are drilled at some angle with the vertical rather than the usual vertically downward direction. Such inclined blastholes offer many advantages but many disadvantages as well. Various parameters associated with vertical and inclined blastholes are shown in Figure 26.13. Advantages of a blast engineered with inclined blastholes are: 1
Length of explosive column in the blasthole increases and at the same time true burden, shown by B1 in the figure, decreases. With this either better
Figure 26.12 Working of a wheel loader.
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Table 26.9 Bench heights chosen for blastholes of different diameters. Bench heights in m (ft) Blasthole dia. mm
ins
152.40 158.75 171.45 187.33 200.03 228.60 250.83 259.88 279.40 311.15 349.25 381.00 406.40
6 6.25 6.75 7.375 7.875 9 9.875 10.625 11 12.25 13.75 15 16
9.14 10.06 (30) (33)
10.97 11.89 13.11 14.02 14.94 15.85 17.07 17.98 18.90 20.12 21.03 21.95 22.86 24.08 24.99 (36) (39) (43) (46) (49) (52) (56) (59) (62) (66) (69) (72) (75) (79) (82)
Preferred
Not Preferred
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B
B
H1 = H * sec α B1 = B * cos α
B H1 H Zone of Poor Fragmentation SD
α
B1
SD1
Figure 26.13 Parameters associated with vertical and inclined blastholes.
2 3
4
5
6
7 8
9
fragmentation results or the burden B can be increased to get the same level of fragmentation. Volume of zone of poor fragmentation near the bench top, shown by square hatching in the figure, decreases. This means better overall fragmentation. Length of subdrilling can be decreased because formation of hump at the newly formed bench floor is decreased. In certain cases subdrilling can be totally eliminated. Yield of blast, i.e. the volume of blasted rock per meter length of the blasthole, is higher in case of inclined blastholes as compared to vertical blastholes. For this reason cost of drilling and blasting is decreased. Possibility of backbreaks is much less in a blast with inclined blastholes. Backbreaks caused by vertical blastholes like those shown in Figure 26.14 are very dangerous. In many cases they make it difficult to place the blasthole drill for drilling holes in the first row. New bench face gets a profile in such a way that the burden for the inclined blastholes in the next round is more or less uniform. The slope of the bench face is also more stable. This is elucidated in Figure 26.15. Probability of misfires is reduced. Large quantum of explosive energy, which always radiates in a plane perpendicular to the blasthole alignment, is directed towards the free surface. In addition to better fragmentation this also results in less energy for creating ground vibrations. Reduced over-crushing of the rock mass, hence less wastage in the case of minerals like coal. Disadvantages linked with inclined blastholes are:
1
2 3
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Aligning the blasthole drill for inclined blastholes is rather difficult and often results in alignment error by incorrectly choosing the spot for starting the drilling operation. Collaring i.e. start of drilling an inclined blasthole, is somewhat more difficult. Chances of hole deviation while drilling inclined blastholes are far higher. Drilling operations have to be done very carefully.
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Figure 26.14 Dangerous backbreak. Non Uniform Burden With Vertical Blastholes
Uniform Burden With Inclined Blastholes
Figure 26.15 Uniform burden with inclined blastholes.
4
5 6 7 9
More wear and tear is experienced by all the drill string components. This is particularly true in the case of large diameter rotary drills that have to exert very high feed force and torque on the drill string. Since the blasthole drill has to resist heavy horizontal force, the fatigue life of the drill is also reduced. Throw and flyrock hazards of blasting inclined holes is somewhat higher than that of vertical holes. Lesser penetration rate due to reduced efficiency of flushing. Larger quantum of rock mass is thrown at a greater distance. Due to this the muck pile is of lesser height.
All the above disadvantages become more and more severe as the blasthole inclination increases. In mines equipped with wheel loaders and dumper combination, inclined blastholes are practiced in many cases. The angle of inclination is usually of the order of
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15° to 20° but in some cases it is as high as 30°. Properties of the rock mass, particularly the orientation of joints and other types of discontinuities, play an important role in the choice of angle of inclination. The muck pile obtained from a blast of vertical blastholes is more suitable for shovel and dumper combination. Almost all metal mines practice blasts of vertical blastholes. Long inclined blastholes tend to deviate more than vertical blastholes. Therefore, blastholes in dragline mining are usually vertical.
26.4.7
Burden
If burden is excessively large, then only cracks will be developed in the rock mass. There will be no separation and hence no fragmentation will occur. The energy released by the detonation of explosive will be utilized in causing heavy vibrations in the rock mass. As against this, if the burden is very small, the gases will escape towards the bench face with very high velocity and the fragments of the rock near the face will be thrown in the air violently to cause flyrock problems that can prove disastrous or fatal in many instances. Several formulae have been proposed by mining technologists for the calculation of burden. They all relate burden value in terms of some other parameters. It can be stated that evaluation of burden by using these formulae gives a value lying somewhere between 25 * D to 40 * D, where D is the diameter of the blasthole. The most appropriate value of burden certainly depends upon the hardness of the rock mass. Figure 26.16 shows the variation of burden for different diameters of blastholes and different rocks. Density of rock as well as type of explosive used for blasting have significant influence on the value of burden, because in the case of heavier rocks a higher part of
Soft Rock
9 8
Medium Hard Rock Hard Rock
7
Burden in m
6 5 4 3 2 1
50
100 150 200 250 Blasthole Diameter in mm
300
350
Figure 26.16 Burden as a function of blasthole diameter.
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Design of a surface blast
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the explosive energy is spent in lifting and throwing the rock mass. In order to avoid this, burden is reduced while blasting with low energy explosives or when the bench has high density rock mass. Table 26.10 gives guidelines for calculating the value of burden for rocks of different densities in case of ANFO and Slurry Dynamite which are low and high energy explosives respectively. Many researchers have tried to establish a correlation between burden spacing and other blasthole parameters. Details are given in a later subsection.
26.4.8
Blasthole spacing
Spacing is usually calculated from the value of burden. When blastholes are drilled for a mine bench, spacing of value 1.1 * B to 1.5 * B, where B is the value of burden, is chosen. Even within this range the value of 1.1 * B has been found to be more appropriate for large diameter blastholes and 1.5 * B is more suitable for small diameter blastholes. Figure 26.17 gives guidelines for appropriate choice of spacing in terms of burden. Spacing value of less than 1.0 * B should never be used except when using controlled blasting techniques such as smooth blasting or cushion blasting. When blastholes in a row are very closely spaced, a crack from one blasthole progresses to the other blasthole very rapidly. With this, the tendency of forming a huge block of rock mass between the first row and bench face increases. Large craters are also formed at the bottom of the blasthole. This gives rise to stumps at the toe of the previous bench face. If the blastholes have large spacing in the same row, the chances of formation of humps in the newly formed bench face increase. Similarly the likelihood of formation of stumps at the toe of the previous bench increases. Table 26.10 Dependence of burden on rock density and type of explosive. Values of burden in terms of blasthole diameter D for rocks of different densities Type of explosive
Low 2200 kg/m3
Medium 2700 kg/m3
High 3200 kg/m3
ANFO Slurry Dynamite
28 * D 33 * D
25 * D 30 * D
23 * D 27 * D
Spacing value in Terms of Burden
1.5 1.4 1.3 1.2 1.1 150
200 250 300 350 Blasthole Diameter in mm
400
450
Figure 26.17 Spacing as a function of burden.
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Rotary drilling and blasting in large surface mines
Both these phenomena are illustrated in Figure 26.18. Determining the spacing-to-burden ratio is greatly affected by the geological conditions of the rock mass. Conditions when holes in the same row have to be blasted at different instances do arise. Often value of spacing is arrived at after onsite experience.
26.4.9
Subdrilling
Subdrilling means the length of blasthole drilled below the level of surface of the bench floor. This is well illustrated in Figure 26.1. In bench blasting it is always desirable that after the bench blast and subsequent removal of the fragments, the new floor formed will be at the same level as that of the original bench face, and there will be no stumps i.e. protrusions of unfragmented rock out of the newly formed surface. The need for subdrilling is to ensure these objectives are easily achieved. A blasthole is always initiated from its bottom. Therefore, the detonation zone travels from bottom to top. For this reason the compression waves do not travel in a plane perpendicular to the blasthole but at an upward angle. If a blasthole drilled without subdrilling is detonated from the bottom, there is always a zone as shown in the left hand sketch in Figure 26.19 where compression waves do not reach and tension waves are not reflected. Thus, the rock mass in this zone remains unfragmented, giving rise to stumps. Whereas if a blasthole drilled with subdrilling is detonated from the bottom, there is no zone of unfragmented rock as shown in right hand sketch in Figure 26.19. Thus, no stumps remain when the fragmented material is removed Stumps at the Toe of Old Bench Face
Stumps at the Toe Block
Block
Heavily Crushed Zone
Humps at the New Bench Face
Figure 26.18 Effects of spacing as a function of burden. Blasthole Direction of propagation of compression waves
Subdrilling Zone where compression waves do not reach
Figure 26.19 Need for subdrilling.
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Design of a surface blast
619
Subdrilling length has to be optimum. If the subdrilling is too large it will result in: 1 2 3 4
Increase in drilling and blasting costs An increased vibration level Excessive fragmentation of the bench, affecting slope stability Increased risk of cutoffs and overbreak
Optimum length of subdrilling depends upon the strength of the rock. Table 26.11 presents the subdrilling length to be used for blasting in intact rocks of different strength. In cases where the rock is highly fractured, subdrilling length can be reduced by about 25% as shown. In some coal mines there is a horizontal plane of weakness between the overburden layer and coal layer. In such cases, to ensure that only the overburden layer is fragmented and removed without fragmenting the coal layer, no subdrilling is practiced. In fact, in such circumstances it may be worthwhile to have the bottom of the blasthole at some distance above the top of the coal bed. If the blastholes are inclined, the subdrilling length can be reduced as shown in Figure 26.20.
Table 26.11 Recommended subdrilling length in terms of burden. Strength classes of the rock
Subdrilling in intact rock condition
Subdrilling in highly fractured conditions
Soft Rock with Easy Toe Medium Rock with Normal Toe Hard Rock with Difficult Toe
0.1 B to 0.2 B 0.3 B 0.4 B to 0.5 B
0.07 B to 0.15 B 0.25 B 0.3 B to 0.4 B
Inclination Angle in °
40
30
20
10
0 0.2 B
0.3 B 0.4 B Subdrilling Length
Figure 26.20 Reduction of subdrilling with blasthole inclination.
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26.4.10
Stemming height
While charging the blasthole, some part of the blasthole is intentionally filled with inert material like rock chips formed in drilling. This is called stemming. Almost invariably some length of a blasthole at its top is used for stemming. The purpose of stemming is to confine gases generated in detonation of the explosive and utilize the energy generated by the blast in the most effective manner. If stemming length is very small it leads to a premature escape of the gases leading to an airblast and a danger of flyrock, the hurling of rock fragments in a blast. On the other hand, if stemming length is excessively long, the portion around a blasthole near bench top is not subjected to sufficient cracking by the blast and large boulders are formed. It also leads to reduced loosening of the rock and subsequent difficulty in its removal. Apart from stemming length, the other parameters associated with stemming column are the type and size of material to be used for stemming. Studies have shown that coarse angular material, such as crushed rock, is the most effective stemming product. The optimal stemming length varies between 18D to 30D, where D is the diameter of the blast hole. Crushed rock of size between 0.04 D and 0.06D is found to be most suitable as stemming material. It has been found to effectively lower the stemming length by about 41%. Use of stemming plugs in the blasthole gives lower mean fragment size. For some specific objectives, stemming is also used in between two or more zones of explosive columns.
26.4.11
Size of the blast
The shape of the blast refers to the geometrical form of the area on the bench top from which all the blastholes are detonated together with delays in their detonation timings. In surface mines the shape of the area is almost invariably rectangular. Such area has length and width. The length is measured along the bench crest and the width is measured perpendicular to the bench face. As the first, second, third rows, and so on, are blasted with delay timings, the fragmented rock mass from the burden of the first row is thrown on the bench floor at a long distance. The fragmented rock mass from the burden of the second row is also thrown horizontally but a larger portion of this sits on the top of the rock mass of the first burden but near to the bench. For the blast of the third row the same phenomenon is repeated and the height of the heap increases as shown in Figure 26.21. The advantage of a large blast area is that it allows more effective utilization of loading and hauling equipment. Disadvantages of a large blast area are: 1 2 3
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The intensity and duration of vibrations created in the ground by the blast increases. The face of the bench formed after removing the blasted material from bench floor becomes crooked like the one shown in Figure 26.15. Fly rock hazard increases as it becomes difficult to introduce appropriate delays in the blasting of rear rows.
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Design of a surface blast
Fragments from Third Burden
621
Fragments from Fourth Burden
Fragments from Second Burden Fragments from First Burden
Figure 26.21 Formation of a heap of fragments after the blast.
If the equipment used for material removal can handle tall heaps of blasted fragments, then even 7 or 8 rows can be taken for a blast. In the case of coal beds the number of rows reduces to 3.
26.5
CALCULATION OF BURDEN
The most important parameter in blast design is the burden. This is because all the rockmass contained in the burden needs to be fragmented by the energy evolved in the blast. Due to the impossibility of mathematical derivation, many researchers have tried to arrive at empirical relations between many parameters of the blasthole and the environment and the burden. The following elaboration is based upon this. Some of these relations are explained here below.
Fraenkel (1952) B = (Rv * L0.3 * l0.3 * D0.8)/50 where B = Burden in m L = Length of blasthole in m l = Length of charge in m D = Diameter of blasthole in mm Rv = Resistance factor. It varies between 1 and 6. For rocks with high UCS it is 1.5 and for rocks with low UCS it is 5
Pearse (1955) B = Kv * 10–3 * D * (PD/RT)0.5 where B = Maximum burden in m Kv = Rock property constant that varies between 0.7 to 1.0 D = Diameter of blasthole in mm
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PD = Detonation pressure in kg/cm2 RT = Tensile strength of rock in kg/cm2
Hino (1959) This method of calculating the burden, proposed by Hino, takes into consideration many variables but somewhat intricately. The first equation is: B = (D/4) * (PD/RTd)1/n where B = Burden in m D = Diameter of blasthole in cm PD = Detonation pressure kg/cm2 RTd = Dynamic tensile strength in kg/cm2 The value of the characteristic constant n is to be determined from the data obtained in the crater-forming test for the explosive as under. n = log(PD/RTd)/log(2Do/(d/2)) where Do is the optimum depth of the center of gravity of the explosive charge in cm. It is to be determined graphically by using the equation Dp = ΔΣV10.333 where d = Diameter of the charge Dp = Depth of the CG of charge Δ = Relationship ratio of depth Dp/Dc Dc = Critical depth of CG of charge Σ = Volumetric constant of the charge V1 = Volume of charge used in the test CG means center of gravity.
Langefors and Kihlstrom (1963) Equation proposed by Langefors and Kihlstrom for calculating burden is as under. Bmax = (D/33) * ((ρe * PRP)/(Co * f * (S/B)))0.5 where Bmax = Maximum burden for good fragmentation in m D = Diameter at bottom of blasthole in m ρe = Density of explosive in the blasthole in kg/L
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623
PRP = Relative Weight strength of the explosive f = Degree of fixation of the blasthole. For vertical blastholes f = 1, for blastholes inclined at 3:1 f = 0.9 and for blastholes inclined at 2:1 f = 0.85 S/B = Spacing to burden ratio The value of Co is to be determined as under. Let the quantity of explosive needed to fragment 1 m3 of rock mass be called C. For hard rocks C is usually taken as 0.4. This value is to be modified as For B = 1.4 to 15 m Co = C + 0.75 For B = 4B For sequenced blast of single row blastholes S = (H + 7B)/8 when H < 4B S = 1.4B when H > 4B In all the above equations B = Burden calculated as above in ft, S = Spacing in ft H = Bench height in ft Stemming distance to be kept as under, For massive rock T = B For Stratified rock T = 0.7 * B In the year 1983, Konya and Walters refined their approach and came up with the following equations. B = ((2ρe/ρr) + 1.5) * D Stemming distance T = 0.7B Subdrilling distance J = 0.3B
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26.6
625
RELATIONSHIP OF BLASTHOLE PARAMETERS
For every blast there is such a relationship between the geometrical parameters of the blasthole, that proves to be the best for the blast from the viewpoint of fragmentation as well as safety. Ash adapted the simplest approach to establish the relationship. The following elaborations are largely based on his paper published in 1963. All the parameters used for establishing the relationship are shown in Figure 26.22.
Spacing to burden relationship As has been seen earlier in this chapter, spacing is considered to be proportional to the burden and can be expressed as, S = Ks * B where Ks is the proportionality constant.
Burden to blasthole diameter relationship Let us presume that for a certain burden and spacing, one unit of blasthole depth gives satisfactory blasting. Therefore, a volume V of rock mass fragmented by the energy output of the explosive in one unit depth of blasthole works out as:
H = Bench Height D = Blasthole Diameter B = Burden S = Spacing J = Subdrilling T = Stemming Height L = Blasthole Length
S
Crest
S T
H B
L
B
Toe D
J
Figure 26.22 Symbolic nomenclature of a blasthole.
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Rotary drilling and blasting in large surface mines
V = B * S * 1 = Ks * B2 We can equate the explosive energy output from one unit length of blasthole as Eo = (π/4) * D2 * Eu * ρu Since in a blast the explosive is same, and the charging density of the explosive is constant, we have Eo ∝ D2 Similarly, Eo ∝ V ∝ D2 i.e. B2 ∝ D2 or B ∝ D This gives: B = Kb * D Naturally, as the diameter of the blasthole increases the burden also increases, in the burden-to-blasthole diameter ratio Kb.
Subdrilling to burden relationship As has been seen earlier, for all the explosives and particularly ANFO, the velocity of detonation at the point of detonation is somewhat low but increases as the detonation progresses. Further, the energy output is also proportional to the velocity of detonation. The purpose of keeping some subdrilling length is to ensure that the velocity of detonation in the explosive reaches its maximum when the detonation reaches the floor level of the bench. In case of ANFO it has been found that the length required for the velocity of detonation to reach its maximum is 6 times the diameter of the blasthole. It is also true that the detonator and primer together are never at the bottom of the blasthole but at a height of about 2D. Therefore, J = 8D and since D ∝ B, we have J ∝ B. In other words, J = Kj * B where Kj is the subdrilling-to-burden ratio.
Stemming to burden relationship In bench blasting practice, the stemming length is also considered to be related to the burden by a stemming-to-burden ratio Kt. In other words, T = Kt * B
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Based upon some tangible logic involved in the blasting process, Langefors and Kihlstrom have concluded that Kt = 0.7
Bench height to burden relationship Bench height is always more than burden. The ratio of bench height to burden is symbolized by Kh and the equation relating bench height and burden is H = Kh * B On the basis of a study of the outcome of many blasts, Ash established that Kh should be more than 1.6.
Length of blasthole When the blastholes are vertical the length of blasthole is the sum of bench height H and subdrilling length J. The equation below can be established in this regard: L=H+J When blastholes are inclined with the vertical at an angle β, the relationship takes the form: L = H/cosβ + (1 − β/100) * J
26.7
DESIGN OF CONVENTIONAL SURFACE BLAST
Various parameters of a conventional bench blast with blastholes are shown in Figure 26.25. A method to affix the values of these parameters is as under. This method is based on the crater theory put forth by Livingstone. One of the basic assumptions for the blasthole is that Length/Diameter is less than 50. 1 Ascertain the diameter of blasthole D, on the basis of hourly production requirements by using Table 26.3. 2 On the basis of the diameter of the blasthole and the type of explosive used for blasting, find the burden B, and spacing distances S, by using Table 26.12. 3 By using Table 26.13 determine the subdrilling length J, in terms of blasthole diameter. 4 Arrive at the value for stemming length T, on the basis of blasthole diameter from Table 26.14. 5 Bench height H, can be determined on the basis of Table 26.15. As shown by Table 26.9, great variation can be allowed in bench height. 6 Length of blasthole L = H + J 7 Specific volume in m3/m Vb = π * (D2/4) * 10−6
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Rotary drilling and blasting in large surface mines
Table 26.12 Burden and spacing for different explosives in different types of rock mass. Type of rock mass and compressive strength in MPa Type of explosive
Parameter
Soft UCS < 70
Medium 70 < UCS < 180
Hard UCS > 180 MPa
ANFO
Burden B Stemming S Burden B Stemming S
28D 33D 38D 45D
23D 27D 32D 37D
21D 24D 30D 34D
Watergels/ Emulsions
Table 26.13 Subdrilling length for blastholes of different diameter. Blasthole diameter Fragment weight in kg
180–250 mm
250–445 mm
Subdrilling-J
7 D to 8D
5 D to 6D
Table 26.14 Stemming length for different types of rock masses. Type of rock mass and compressive strength in MPa Parameter
Soft UCS < 70
Medium 70 < UCS < 180
Hard UCS > 180 MPa
Stemming-T
40D
32D
25D
Table 26.15 Bench height for different types of rock masses. Type of rock mass and compressive strength in MPa Parameter
Soft UCS < 70
Medium 70 < UCS < 180
Hard UCS > 180 MPa
Bench Height-H
52D
44D
37D
8 9 10 11
Breakage volume is Vr = B * S * H Yield of breakage is Yb = Vr/L Bottom charge length Lb = 8D to 10D. Column charge length is Lc = L − T − Lb
From the above basic parameters and properties of explosives like density of bottom charge ρb and density of column charge ρc, other parameters related to the blast can be determined as below.
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1 2 3 4 5 6
629
Concentration of column charge Cc = ρc * Vb Concentration of bottom charge Cb = ρb * Vb Total bottom charge Qb = Lb * Cb Total column charge Qc = Lc * Cc Total quantity of explosive Qt = Qb + Qc Powder Factor Vr = Qr/Vr
Example 1 In a large copper mine, blastholes of diameter 311 mm are being drilled to cope with the desired production rate. The formation has UCS of 150 MPa. ANFO with a charge density of 0.85 kg/L is used as the explosive in the column charge, and emulsion with charge density of 1.35 kg/L is used as explosive in the bottom charge. Determine various parameters of the blast.
Solution 1 Since D = 311 mm and formation UCS is 150 MPa, Burden B = 23D = 23 * 0.311 = 7.153 m Spacing S = 27D = 27 * 0.311 = 8.397 m Subdrilling J = 6D = 6 * 0.311 = 1.866 m Stemming length T = 32 * D = 32 * 0.311 = 9.952 m Bench height H = 44 * D = 44 * 0.311 = 13.684 m Length of blasthole L = H + J = 15.55 m Specific volume Vb = π * 3112/4 * 10–6 = 0.07597 m3/m Breakage volume Vr = 7.153 * 8.397 * 13.684 = 821.91 m3 Yield of breakage Yb = 821.91/15.55 = 52.856 m3/m Bottom charge length Lb = 8 * 0.311 = 2.488 m Column charge length Lc = 15.55–9.952–2.488 = 3.11 m Density of bottom charge is 1.35 kg/L = 1350 kg/m3 Density of column charge is 0.85 kg/L = 850 kg/m3
Stemming Length = 15D
B
Variable Length of Column Charge
Intermediate Stemming = 1 m Bottom Charge = 55D
Figure 26.23 Rip rap production blasting.
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Concentration of bottom charge Cb = 0.07597 * 1350 = 102.56 kg/m Concentration of column charge Cc = 0.07597 * 850 = 64.57 kg/m Total bottom charge Qb = 2.488 * 102.56 = 255.17 kg Total column charge Qc = 3.11 * 64.57 = 200.81 kg Total quantity of explosive Qt = 255.17 + 200.81 = 455.98 kg Powder factor = 455.98/821.91 = 0.555 kg/m3
26.8
DESIGN OF OTHER TYPES OF BLASTS
Apart from conventional surface blasts, described in the previous section, there are other types of blasts in large mines. These are designed on somewhat different criteria. This section deals with such blasts.
26.8.1
Rip rap production blasting
As said earlier, the objective of rip rap blasting is to get large pieces of rock. The following should be noted in the design of a blast in large quarries meant for the purpose. Relevant blasthole alignment is shown in Figure 26.23. 1 The rock mass should be as monolithic as possible. A heavily fractured rock mass is almost useless for the purpose as it gives smaller rock pieces. 2 The rock should be as dense as possible. 3 The diameter of blastholes is around 100 mm. 4 Bench height should be between 15 to 20 m. 5 Blastholes should be inclined at an angle 5° or 10° with the vertical. 6 Burden should be about 35 * D to 40 * D. 7 Spacing should be 1.5 * B to 2.0 * B. 8 Subdrilling should be 10 * D 9 Powder factor to be used in case of bottom charge should be: For rocks with UCS > 100 MPa PF > 650 g/m3 For rocks with UCS < 100 MPa PF < 500 g/m3 10 Charge density in the plane of cut should be: For rocks with UCS > 100 MPa CD > 500 g/m2 For rocks with UCS < 100 MPa CD < 250 g/m2 11 The whole row of the blastholes should have instantaneous charge. Fragment distribution obtained in homogeneous rocks by using above blast design parameters has been found to be as given in Table 26.16.
26.8.2
Snake hole blasting
Initially snake hole blasting was thought of as an excellent means of achieving very good fragmentation. However, soon it was noticed that the disadvantages experienced were far too great to be offset by the advantages.
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Table 26.16 Fragment size distribution achieved in rip rap blasting. Percentage in the fragmented rock Fragment weight in kg
Rock with UCS < 100 MPa
Rock with UCS > 100 MPa
>3000 kg 1000–3000 kg 50–1000 kg Finer Than 50 kg
30 20 25 25
50 25 25 10
B
β
H S Snake Holes 0.5+B
S/2
L
Figure 26.24 Snake hole blasting.
The main disadvantages were as under: 1 2 3 4 5 6
Very specialized drills are required for drilling horizontal blastholes. Drilling horizontal blastholes without significant deviation is rather difficult. Penetration rate of drilling horizontal blastholes is very low. Horizontal drilling activity hampers the movement of other pieces of equipment to a large degree. Working at two levels (on bench top and bench floor) is quite cumbersome. Flyrock havoc caused by blast in the horizontal blastholes is excessively high and cannot be easily controlled.
Today, snake hole blasting is practiced in very few instances. Naturally, there has neither been much accumulated experience nor significant research to arrive at design criteria. Snake blastholes are shown in Figure 26.24. So far the practice followed in the field has been: 1 2
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The horizontal blastholes are of diameter 89 to 110 mm, so drilling by small size top hammer or DTH hammer drill is possible. Vertical blastholes are usually at an angle of 5° to 15°.
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Rotary drilling and blasting in large surface mines
Vertical blastholes are drilled to such a depth that the distance between the bottom of the vertical blastholes and the horizontal blastholes, which in other words is the burden Bh of the horizontal holes, is (0.5 + B) m, where B is the burden for vertical blastholes. Spacing of the vertical blastholes is as per the normal criteria of S = Ks * B, where Ks equals 1.0 to 1.5. The spacing Sh of horizontal blastholes is 0.5 * S, where S is the spacing of vertical blastholes. Horizontal blastholes have length L (i.e. depth) equal to n * B/cos β, where β is the angle of inclination of vertical blastholes.
For the following aspects, the field practice is not known but author’s recommendations are: a b c
Stemming length for vertical blastholes should be 0.8 * B to 1.0 * B. Stemming length of horizontal blastholes should be 1.0 * Bh. Explosive used in vertical blastholes can be ANFO but in the horizontal blastholes explosive with higher strength- such as ALANFO- should be used.
26.8.3
Cast blasting
In the early 1980s an idea of using explosive energy for throwing the fragmented rockmass to the adjacent waste heap simultaneously through the same blast was born. Over the last decades the techniques of such use have been refined greatly. For a successful cast blast the following parameters of the bench, pit, wall etc. should be practiced. The parameters are shown in Figure 26.25. Bench height should be more than 12 m. Pit width should be 1.25 times the height of the bench face. Presplitting is essential because it enables the burden for the subsequent blast round to be the same for all the blastholes in the first row. Presplitting blastholes should be 251 mm dia., with spacing of 3 m in soft or tough rocks, and 5 to 6 m in hard and brittle rocks. The type of formation matters in choosing the quantity of explosive. Since soft rocks fracture quickly and the gases formed in the explosion escape early, the throw in the soft rock masses is low. To compensate for this the quantity of explosive Near vertical free face formed
Waste Heap
Bench Height Pit Width
Figure 26.25 Parameters in cast blasting.
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required in blasting soft rock is rather high. Hard rock masses require less quantity of explosive. Burden also greatly depends upon geological factors of the rock mass. It is usually determined from a nomograph on the basis of total charge weight in a blasthole and rock blastability. The explosive selected must be able to give sufficient ejection velocity so the broken fragments are thrown in the appropriate direction to the appropriate distance. The ejection velocity can be calculated from the following empirical equation: Ve = 1.14 * (B/(0.07853 * d2 * ρe * PAP))−1.17 where Ve = Ejection velocity in m/s B = Burden in m d = Charge diameter in cm ρe = Density of explosive g/cm2 PAP = Absolute weight strength in cal/g Minimum ejection velocity is about 15 m/s. Spacing should be about 1.3 to 1.6 times burden. The blasthole layout pattern should be rectangular staggered. Stemming should be 20 to 25 times blasthole diameter. Subdrilling should be low i.e. 4 to 6 times blasthole diameter. This is essential so as to ensure that the coal layer is not broken, mixed and thrown with the overburden. Powder factor varies between 0.3 to 0.8 kg/m. All the blastholes in a row should be fired simultaneously without any delay so long as only one face is involved. If two faces are involved then the firing sequence should be as shown in Figure 26.26. In order to ensure that material fragmented and thrown from the first row blast does not infringe with the material fragmented and thrown by the second row blast,
All the blastholes on the same dotted line are to be blasted simultaneously.
Crest
Bench Top
Toe
Minor Face
Bench Floor Blast Initiation
Major Face
Figure 26.26 Firing sequence for cast blasting with two faces.
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Range DP in m
Powder Factor CE in kg/m3 1.250
35
Strain Energy Factor FE 4.4 4.2
1.125 30
4.0 1.000 3.8
25 0.875
3.6
20
0.750 3.4 0.625
15
3.2 0.500
10
3.0 0.375 2.8 0.250 2.6
Figure 26.27 Nomograph 1.
the delay timing between rows should be worked out on the basis of 20 to 25 ms/m of burden. A company, named D’Appolonia Consulting Engineers, has developed nomographs to evaluate some of the variables used in cast blasting. The nomographs are shown in Figure 26.27 to Figure 26.30 and their use is explained through Example 2.
Example 2 Cast blasting method is to be adopted in an existing coal mine. The following are the details of the existing method. Blasthole diameter −250 mm Strain energy factor FE −3.4 Bench height −13.5 m Throw distance DP −16 m Density of explosive −0.85 kg/L Determine other parameters of the cast blast.
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Design of a surface blast Blasthole Diameter D in mm 300 Blasthole Loading Factor qt kg/m Total Charge Weight Per Blasthole Qt in kg 250 105 1575 225 75 1125 200 900 675 175 45 450 360 30 150 Explosive 270 225 Density in g/cc 125 15 12 100
9 6
75
4.5 3
180 135 90 67.5 45 36 27 18 13.5 9
635
Blasthole Loading Length in m
1.40
15 13.5 12 10.5 9
1.30 1.20 1.15 1.10 1.05 1.00 0.95 0.90 0.85
7.5
0.80
4.5
6
3
Figure 26.28 Nomograph 2.
Solution 2 The stepwise solution by using D’Appolonia Method is as under. Step 1 – In nomograph 1, draw a straight line from FE = 3.4 to DP 16 m. This line intersects the powder factor (CE) line to give a powder factor value of 0.67 kg/m3. Step 2 – Using nomograph 2, draw a straight line from D = 250 to ρc = 0.85 kg/m3. This line intersects the blasthole loading factor qt as 40 kg/m. Step 3 – Calculate C1 and C2 by assuming K1 and K2 to be equal to 1. This actually means that burden and spacing are both equal to each other. C1 = (10.66 * q1)/CE * K2 = 10.66 * 40/(0.67 * 1) = 636.42 C2 = 0.3 * K1 * 636.42/H = 0.3 * 1 * 636.42/13.5 = 14.14 Step 4 – By using the above values of C1 and C2 find out the value of C3 from nomograph 3.
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C3
C2
13000
1600
C'3 12000
80
75
20 11000
1400
1200 70
10000 8000
1000
6000 5000
7000
4000 600
6
40
4.5 50
1.5 60
400
2000
3000 70 200
30 1000 20 10
30
3000
50
40
9
3
800
5000
Burden B in m 10.5
7.5
9000
65 60
C'2 10
C1
1000 80
0
Figure 26.29 Nomograph 3.
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Design of a surface blast
637
Total Charge Weight per Blasthole Qt in kg 1360 900 675 540 450
Optimum Burden in m
Blastability Factor FV 2.76 2.70
10.5 9.0
2.60
7.5 FE
FV
2.50
2.0
1.9
2.40
2.2
2.04
2.4
2.18
2.6
2.32
2.8
2.46
3.0
2.60
3.2
2.73
31.5
3.4
2.70
22.5 18
3.6
2.57
3.8
2.43
13.5
4.0
2.30
4.2
2.17
4.4
2.03
315 225
6.0
180 135
4.5
90 67.5 45
9
3.0
1.50
2.30 2.20 2.10 2.00 1.90
Figure 26.30 Nomograph 4.
Step 5 – Let C3 = C'3 and C2 = C'3. Use these values to find out the burden B from nomograph 3. With this we get value of burden B as 5.8 m. Step 6 – The length of charge in the blasthole works out to l = H − K1 * B = 13.5−1 * 5.8 = 7.7 m Step 7 – Using nomograph 2 again, determine Qt from the values of qt and l. In this case since qt = 40 kg/m and l = 7.7 m. Total charge weight Qt works out to 320 kg. Step 8 – Since FE = 3.4, from the table in nomograph 4 the value of FV can be found out as 2.7. Now, using nomograph 4 with values of Qt and FV to be 320 kg and 2.7 respectively the optimum burden Bo can be found out to be 7.2 m. Step 9 – If the values of B and Bo are nearly equal the use of nomographs is over and stemming length T as well as spacing S can be determined by the equations S = K2 * B and T = K1 * B. Step 10 – In this case B and Bo are not nearly equal. For such cases the calculations are required to be repeated from step 3 by using different values of K1 and K2 and recalculating C1 and C2.
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To avoid blind guesswork about K1 and K2, a rule of thumb has been suggested by D’Appolonia. The rule is to use the relation K2 = K13 for estimation. It can be seen that with a value of K1 = 0.86 and K2 = 0.636, C1 and C2 work out to 1043.6 and 19.95 respectively. For these values B and Qt works out to 7.3 m and 301.3 kg. With this value Bo also works out to 7.3 m. Thus, Spacing S = 0.636 * 7.3 = 4.6428 m. Similarly T = 0.86 * 7.3 = 6.28. Since the procedure described above is based on nomographs backed by sound mathematics, the whole procedure can be converted into a computer program. With such a program it is possible to get quick and more precise results.
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Appendix 1
Properties of atmospheric air at high altitudes
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Table A1.1 Properties of atmospheric air at high altitudes. Atmospheric pressure
Altitude
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m
ft
kPa
psi
−400 −200
−1312.34 −656.17
0 200 400 600 800 1000 1200 1400 1600 1800 2000 2200 2400 2600 2800 3000 3200 3400 3600 3800 4000 4200 4400 4600 4800 5000
0.00 656.17 1312.34 1968.50 2624.67 3280.84 3937.01 4593.18 5249.34 5905.51 6561.68 7217.85 7874.02 8530.18 9186.35 9842.52 10498.69 11154.86 11811.02 12467.19 13123.36 13779.53 14435.70 15091.86 15748.03 16404.20
106.223 103.751 101.325 98.945 96.611 94.322 92.076 89.875 87.716 85.599 83.524 81.489 79.495 77.541 75.626 73.749 71.910 70.109 68.344 66.615 64.922 63.264 61.640 60.050 58.494 56.971 55.479 54.020
15.406 15.048 14.696 14.351 14.012 13.680 13.355 13.035 12.722 12.415 12.114 11.819 11.530 11.246 10.969 10.696 10.430 10.168 9.912 9.662 9.416 9.176 8.940 8.710 8.484 8.263 8.047 7.835
Atmospheric temperature
Gravitational acceleration
°C
°F
m/s2
63.68 61.34 59.00 56.66 54.32 51.98 49.64 47.30 44.96 42.62 40.28 37.94 35.60 33.26 30.92 28.58 26.24 23.90 21.56 19.22 16.88 14.54 12.20 9.86 7.52 5.18 2.84 0.50
Decreases Below Surface of Earth 9.80665 32.1740 9.80603 32.1720 9.80542 32.1700 9.80480 32.1680 9.80419 32.1660 9.80357 32.1639 9.80295 32.1619 9.80234 32.1599 9.80172 32.1579 9.80111 32.1559 9.80049 32.1538 9.79988 32.1518 9.79926 32.1498 9.79865 32.1478 9.79803 32.1458 9.79742 32.1438 9.79680 32.1417 9.79619 32.1397 9.79557 32.1377 9.79496 32.1357 9.79434 32.1337 9.79373 32.1316 9.79311 32.1296 9.79250 32.1276 9.79188 32.1256 9.79127 32.1236
17.60 16.30 15.00 13.70 12.40 11.10 9.80 8.50 7.20 5.90 4.60 3.30 2.00 0.70 −0.60 −1.90 −3.20 −4.50 −5.80 −7.10 −8.40 −9.70 −11.00 −12.30 −13.60 −14.90 −16.20 −17.50
Density ft/sec2
kg/m3 1.2727 1.2487 1.2250 1.2017 1.1786 1.1560 1.1336 1.1116 1.0900 1.0686 1.0476 1.0269 1.0065 0.9864 0.9666 0.9472 0.9280 0.9091 0.8905 0.8723 0.8543 0.8366 0.8191 0.8020 0.7851 0.7685 0.7522 0.7361
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lb/ft3
Barometer mm
Water column m
0.0795 0.0780 0.0765 0.0750 0.0736 0.0722 0.0708 0.0694 0.0680 0.0667 0.0654 0.0641 0.0628 0.0616 0.0603 0.0591 0.0579 0.0568 0.0556 0.0545 0.0533 0.0522 0.0511 0.0501 0.0490 0.0480 0.0470 0.0460
796.741 778.195 760.000 742.151 724.643 707.471 690.630 674.115 657.921 642.044 626.478 611.219 596.263 581.604 567.239 553.162 539.370 525.857 512.620 499.654 486.954 474.518 462.340 450.416 438.742 427.315 416.129 405.182
10.831 10.579 10.332 10.089 9.851 9.618 9.389 9.164 8.944 8.728 8.517 8.309 8.106 7.907 7.711 7.520 7.333 7.149 6.969 6.793 6.620 6.451 6.285 6.123 5.965 5.809 5.657 5.508
Dynamic viscosity kg/(m . s)
1.789 * 10−5 1.783 * 10−5 1.777 * 10−5 1.771 * 10−5 1.764 * 10−5 1.758 * 10−5 1.752 * 10−5 1.745 * 10−5 1.739 * 10−5 1.732 * 10−5 1.726 * 10−5 1.720 * 10−5 1.713 * 10−5 1.707 * 10−5 1.700 * 10−5 1.694 * 10−5 1.687 * 10−5 1.681 * 10−5 1.674 * 10−5 1.668 * 10−5 1.661 * 10−5 1.655 * 10−5 1.648 * 10−5 1.642 * 10−5 1.635 * 10−5 1.628 * 10−5
Thermal conductivity W/(m . K)
0.0253 0.0252 0.0252 0.0251 0.0250 0.0249 0.0248 0.0247 0.0245 0.0244 0.0243 0.0242 0.0241 0.0240 0.0239 0.0238 0.0237 0.0236 0.0235 0.0234 0.0233 0.0232 0.0231 0.0230 0.0229 0.0228
Speed of sound m/s
ft/sec
341.826 341.061 340.294 339.526 338.755 337.983 337.210 336.434 335.657 334.878 334.097 333.314 332.529 331.743 330.955 330.164 329.372 328.578 327.782 326.985 326.185 325.383 324.579 323.773 322.966 322.156 321.344 320.530
1121.476 1118.966 1116.450 1113.929 1111.402 1108.869 1106.331 1103.786 1101.236 1098.680 1096.118 1093.550 1090.976 1088.396 1085.809 1083.217 1080.618 1078.013 1075.402 1072.784 1070.160 1067.529 1064.892 1062.249 1059.599 1056.942 1054.278 1051.608
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Appendix 2
Air properties at various temperatures
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Table A2.1 Properties of air at various temperatures.
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Temperature °C
Density kg/m3
−150
2.866
−100 −50 −40 −30 −20 −10 0 5 10 15 20 25 30 35 40 45 50 60 70 80 90 100 120 140 160 180 200 250 300 350 400
2.038 1.582 1.514 1.451 1.394 1.341 1.292 1.269 1.246 1.225 1.204 1.184 1.164 1.145 1.127 1.109 1.092 1.059 1.028 0.9994 0.9718 0.9458 0.8977 0.8542 0.8148 0.7788 0.7459 0.6746 0.6158 0.5664 0.5243
Specific Heat capacity Cp KJ/(kg . K)
Thermal conductivity W/(m . K)
Thermal diffusivity m2/s
983
0.01171
4.158 * 10−6
966 999 1002 1004 1005 1006 1006 1006 1006 1007 1007 1007 1007 1007 1007 1007 1007 1007 1007 1008 1008 1009 1011 1013 1016 1019 1023 1033 1044 1056 1069
0.01582 0.01979 0.02057 0.02134 0.02211 0.02288 0.02364 0.02401 0.02439 0.02476 0.02514 0.02551 0.02588 0.02625 0.02662 0.02699 0.02735 0.02808 0.02881 0.02953 0.03024 0.03095 0.03235 0.03374 0.03511 0.03646 0.03779 0.04104 0.04418 0.04721 0.05015
8.036 * 10−6 1.252 * 10−5 1.356 * 10−5 1.465 * 10−5 1.578 * 10−5 1.696 * 10−5 1.818 * 10−5 1.880 * 10−5 1.944 * 10−5 2.009 * 10−5 2.074 * 10−5 2.141 * 10−5 2.208 * 10−5 2.277 * 10−5 2.346 * 10−5 2.416 * 10−5 2.487 * 10−5 2.632 * 10−5 2.780 * 10−5 2.931 * 10−5 3.086 * 10−5 3.243 * 10−5 3.565 * 10−5 3.898 * 10−5 4.241 * 10−5 4.593 * 10−5 4.954 * 10−5 5.890 * 10−5 6.871 * 10−5 7.892 * 10−5 8.951 * 10−5
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Dynamic viscosity μ kg/(m . s)
Kinematics viscosity ν m2/s
8.636 * 10−6 −5
1.189 * 10 1.474 * 10−5 1.527 * 10−5 1.579 * 10−5 1.630 * 10−5 1.680 * 10−5 1.729 * 10−5 1.754 * 10−5 1.778 * 10−5 1.802 * 10−5 1.825 * 10−5 1.849 * 10−5 1.872 * 10−5 1.895 * 10−5 1.918 * 10−5 1.941 * 10−5 1.963 * 10−5 2.008 * 10−5 2.052 * 10−5 2.096 * 10−5 2.139 * 10−5 2.181 * 10−5 2.264 * 10−5 2.345 * 10−5 2.420 * 10−5 2.504 * 10−5 2.577 * 10−5 2.760 * 10−5 2.934 * 10−5 3.101 * 10−5 3.261 * 10−5
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Prandtl number Pr
Expansion coefficient 1/K × 103
3.013 * 10−6
0.7246
8.21
−6
0.7263 0.7440 0.7436 0.7425 0.7408 0.7387 0.7362 0.7350 0.7336 0.7323 0.7309 0.7296 0.7282 0.7268 0.7255 0.7241 0.7228 0.7202 0.7177 0.7154 0.7132 0.7111 0.7073 0.7041 0.7014 0.6992 0.6974 0.6946 0.6935 0.6937 0.6948
5.82 4.51
5.837 * 10 9.319 * 10−6 1.008 * 10−5 1.087 * 10−5 1.169 * 10−5 1.252 * 10−5 1.338 * 10−5 1.382 * 10−5 1.426 * 10−5 1.470 * 10−5 1.516 * 10−5 1.562 * 10−5 1.608 * 10−5 1.655 * 10−5 1.702 * 10−5 1.750 * 10−5 1.798 * 10−5 1.896 * 10−5 1.995 * 10−5 2.097 * 10−5 2.201 * 10−5 2.306 * 10−5 2.522 * 10−5 2.745 * 10−5 2.975 * 10−5 3.212 * 10−5 3.455 * 10−5 4.091 * 10−5 4.765 * 10−5 5.475 * 10−5 6.219 * 10−5
3.67
3.43
3.20
3.00 2.83 2.68 2.55 2.43 2.32 2.21 2.11 1.91 1.75 1.61 1.49
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Appendix 3
Bailing velocities in blasthole annulus
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Table A3.1 Velocity of air flowing up between the blasthole and drill pipe. Hole diameter D
Drill pipe diameter d
Annular area
Value 0.833D
Value 0.792D
Annulus/ hole area
Volume of compressed air (m3/min) required for attaining the bailing velocities (m/min) given here below
Inch
mm
Inch
Sq. In
sq. m
in mm
in mm
%
1000
6 6 6 6.25 6.25 6.25 6.75 6.75 6.75 6.75 7.375 7.375 7.375 7.875 7.875 7.875 7.875 9 9 9 9 9.875 9.875 9.875
152.40 152.40 152.40 158.75 158.75 158.75 171.45 171.45 171.45 171.45 187.32 187.32 187.32 200.02 200.02 200.02 200.02 228.60 228.60 228.60 228.60 250.82 250.82 250.82
12034 10134 7981 11686 9532 7126 14980 12826 10419 7759 14892 12232 7767 18756 16096 11631 8847 18467 16215 14409 10609 24583 22778 18978
126.95 126.95 126.95 132.24 132.24 132.24 142.82 142.82 142.82 142.82 156.04 156.04 156.04 166.62 166.62 166.62 166.62 190.42 190.42 190.42 190.42 208.94 208.94 208.94
120.70 120.70 120.70 125.73 125.73 125.73 135.79 135.79 135.79 135.79 148.36 148.36 148.36 158.42 158.42 158.42 158.42 181.05 181.05 181.05 181.05 198.65 198.65 198.65
65.97 55.56 43.75 59.04 48.16 36.00 64.88 55.56 45.13 33.61 54.04 44.38 28.18 59.69 51.22 37.01 28.15 44.99 39.51 35.11 25.85 49.75 46.10 38.41
12.03 10.13 7.98 11.69 9.53 7.13 14.98 12.83 10.42 7.76 14.89 12.23 7.77 18.76 16.10 11.63 8.85 18.47 16.21 14.41 10.61 24.58 22.78 18.98
3.5 4 4.5 4 4.5 5 4 4.5 5 5.5 5 5.5 6.25 5 5.5 6.25 6.675 6.675 7 7.25 7.75 7 7.25 7.75
mm 88.90 101.60 114.30 101.60 114.30 127.00 101.60 114.30 127.00 139.70 127.00 139.70 158.75 127.00 139.70 158.75 169.54 169.54 177.80 184.15 196.85 177.80 184.15 196.85
18.653 15.708 12.370 18.113 14.775 11.045 23.218 19.880 16.150 12.026 23.083 18.960 12.039 29.072 24.949 18.027 13.713 28.623 25.133 22.335 16.444 38.104 35.306 29.416
1300 15.64 13.17 10.37 15.19 12.39 9.26 19.47 16.67 13.54 10.09 19.36 15.90 10.10 24.38 20.92 15.12 11.50 24.01 21.08 18.73 13.79 31.96 29.61 24.67
1600 19.25 16.21 12.77 18.70 15.25 11.40 23.97 20.52 16.67 12.41 23.83 19.57 12.43 30.01 25.75 18.61 14.16 29.55 25.94 23.06 16.97 39.33 36.44 30.36
1900 22.87 19.25 15.16 22.20 18.11 13.54 28.46 24.37 19.80 14.74 28.30 23.24 14.76 35.64 30.58 22.10 16.81 35.09 30.81 27.38 20.16 46.71 43.28 36.06
2200 26.48 22.30 17.56 25.71 20.97 15.68 32.95 28.22 22.92 17.07 32.76 26.91 17.09 41.26 35.41 25.59 19.46 40.63 35.67 31.70 23.34 54.08 50.11 41.75
2500 30.09 25.34 19.95 29.21 23.83 17.81 37.45 32.07 26.05 19.40 37.23 30.58 19.42 46.89 40.24 29.08 22.12 46.17 40.54 36.02 26.52 61.46 56.95 47.44
2700 32.49 27.36 21.55 31.55 25.74 19.24 40.44 34.63 28.13 20.95 40.21 33.03 20.97 50.64 43.46 31.40 23.89 49.86 43.78 38.91 28.64 66.37 61.50 51.24
3000 36.10 30.40 23.94 35.06 28.60 21.38 44.94 38.48 31.26 23.28 44.68 36.70 23.30 56.27 48.29 34.89 26.54 55.40 48.64 43.23 31.83 73.75 68.33 56.93
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Book.indb 649
10.625 10.625 10.625 11 11 11 11 12.25 12.25 12.25 12.25 13.75 13.75 13.75 15 15 15 15 16 16 16 16 16 17.5 17.5 17.5 17.5 17.5 17.5
269.88 269.88 269.88 279.40 279.40 279.40 279.40 311.15 311.15 311.15 311.15 349.25 349.25 349.25 381.00 381.00 381.00 381.00 406.40 406.40 406.40 406.40 406.40 444.50 444.50 444.50 444.50 444.50 444.50
7.75 8.625 9 7.75 8.625 9 9.25 9 9.25 10 10.75 10 10.75 12 10.75 12 12.25 13 12 12.25 13.375 13.625 13.75 12.25 13 13.625 13.75 15 16
196.85 219.07 228.60 196.85 219.07 228.60 234.95 228.60 234.95 254.00 273.05 254.00 273.05 304.80 273.05 304.80 311.15 330.20 304.80 311.15 339.72 346.07 349.25 311.15 330.20 346.07 349.25 381.00 406.40
41.491 30.238 25.047 47.860 36.607 31.416 27.833 54.242 50.658 39.319 27.096 69.950 57.727 35.392 85.952 63.617 58.856 43.982 87.965 83.203 60.562 55.260 52.573 122.669 107.796 94.726 92.039 63.814 39.466
26768 19508 16159 30877 23617 20268 17956 34994 32683 25367 17481 45129 37243 22834 55453 41043 37971 28376 56751 53679 39072 35652 33918 79141 69546 61114 59380 41170 25462
224.81 224.81 224.81 232.74 232.74 232.74 232.74 259.19 259.19 259.19 259.19 290.93 290.93 290.93 317.37 317.37 317.37 317.37 338.53 338.53 338.53 338.53 338.53 370.27 370.27 370.27 370.27 370.27 370.27
213.74 213.74 213.74 221.28 221.28 221.28 221.28 246.43 246.43 246.43 246.43 276.61 276.61 276.61 301.75 301.75 301.75 301.75 321.87 321.87 321.87 321.87 321.87 352.04 352.04 352.04 352.04 352.04 352.04
46.80 34.10 28.25 50.36 38.52 33.06 29.29 46.02 42.98 33.36 22.99 47.11 38.88 23.83 48.64 36.00 33.31 24.89 43.75 41.38 30.12 27.48 26.15 51.00 44.82 39.38 38.27 26.53 16.41
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Gray colored rows indicate the most usually used combinations of blasthole and drill pipe diameters.
26.77 19.51 16.16 30.88 23.62 20.27 17.96 34.99 32.68 25.37 17.48 45.13 37.24 22.83 55.45 41.04 37.97 28.38 56.75 53.68 39.07 35.65 33.92 79.14 69.55 61.11 59.38 41.17 25.46
34.80 25.36 21.01 40.14 30.70 26.35 23.34 45.49 42.49 32.98 22.73 58.67 48.42 29.68 72.09 53.36 49.36 36.89 73.78 69.78 50.79 46.35 44.09 102.88 90.41 79.45 77.19 53.52 33.10
42.83 31.21 25.85 49.40 37.79 32.43 28.73 55.99 52.29 40.59 27.97 72.21 59.59 36.53 88.72 65.67 60.75 45.40 90.80 85.89 62.52 57.04 54.27 126.63 111.27 97.78 95.01 65.87 40.74
50.86 37.07 30.70 58.67 44.87 38.51 34.12 66.49 62.10 48.20 33.21 85.74 70.76 43.38 105.36 77.98 72.15 53.91 107.83 101.99 74.24 67.74 64.44 150.37 132.14 116.12 112.82 78.22 48.38
58.89 42.92 35.55 67.93 51.96 44.59 39.50 76.99 71.90 55.81 38.46 99.28 81.93 50.23 122.00 90.30 83.54 62.43 124.85 118.09 85.96 78.43 74.62 174.11 153.00 134.45 130.64 90.57 56.02
66.92 48.77 40.40 77.19 59.04 50.67 44.89 87.49 81.71 63.42 43.70 112.82 93.11 57.08 138.63 102.61 94.93 70.94 141.88 134.20 97.68 89.13 84.79 197.85 173.86 152.78 148.45 102.92 63.66
72.27 52.67 43.63 83.37 63.77 54.72 48.48 94.49 88.24 68.49 47.20 121.85 100.56 61.65 149.72 110.82 102.52 76.61 153.23 144.93 105.49 96.26 91.58 213.68 187.77 165.01 160.33 111.16 68.75
80.31 58.52 48.48 92.63 70.85 60.80 53.87 104.98 98.05 76.10 52.44 135.39 111.73 68.50 166.36 123.13 113.91 85.13 170.25 161.04 117.22 106.95 101.75 237.42 208.64 183.34 178.14 123.51 76.39
Appendix 4
Air pressure loss in steel pipes
All the necessary information for calculating pressure drop in seamless carbon steel pipes, pipe fittings etc. has been given in chapter 8 of this book. This appendix presents Tables A4.1a and A4.1b, that give pressure drop values for dry compressed air flowing through steel pipes in straight condition. The following conditions have been presumed for calculating pressure drop values given in the tables. Compressed Air Gauge Pressure = 1034 kPa Length of the pipe = 100 m Atmospheric Air Pressure (Abs) = 101325 Pa Atmospheric Temperature = 15°C Dynamic Viscosity of Air = 1.802 kg/(m . s) Absolute Roughness of Pipe = 0.00015 m Gas Constant for Air = 287.1 J/(kg . K) The discharge ratings given in first column on the left, are so chosen that they very closely match with the discharge ratings of compressors mounted on the rotary blasthole drills. Pressure drop values given in Table A4.1a pertain to schedule 40 and schedule 80 pipes as per ANSI Standard. Similarly, the pressure drop values given in Table A4.1b pertain to steel pipes made to DIN Standard. As most of the manufacturers of rotary blasthole drills are located in the United States of America, the original values of compressor pressure and discharge ratings as well as the diameters of the steel tubes are in imperial units but the tables show metric units. For ease in finding equivalent values the compressor discharge ratings have been given in cfm units in column no. 2 of Table A4.1a. Similarly, the nominal diameters of the steel tubes have been given in row no. 4 of Table A4.1a. Pressure drop values given in the tables have been calculated by using Darcy Weisbach equation. The first example in chapter 8 gives all the steps used in the calculations, except for the friction factor. In those examples the friction factor has been found by the use of Moody’s Chart, whereas for calculation of the values in Table A4.1a and A4.1b Haaland’s formula has been used. For actual calculations a spreadsheet program has been used. A complete spreadsheet formulation has been presented in Figure A4.1.
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652
Appendix 4
The main reason for use of Haaland’s formula is that it gives pressure drop values slightly higher than actual. The inaccuracy, therefore, will be on the safer side. Where pressure drop value works out to more than 2% of the compressed air gauge pressure, the pipe is considered unsuitable and instead of giving the pressure drop value the word “Unsuit” is shown. It is to be specifically noted that the pressure drop is inversely proportional to the compression ratio. To facilitate calculation of the pressure drop for gauge pressures other than 1034 kPa and/or atmospheric pressure other than 101.325 kPa, values given in Table A4.2 can be used. It may be noted that steel pipes of all the sizes mentioned in Table A4.1a and A4.1b are available in pressure ratings well above 1500 kPa. EXAMPLE
Find the pressure drop likely to occur in a galvanized schedule 80 pipe of nominal size 4" and length 57 m, when the compressed air flow has a volume of 25.485 m3/min and the gauge pressure of 448000 Pa. The altitude of the place is 3000 m. SOLUTION
Referring to the Table A4.1a we find that pressure loss for Nominal 4" pipe of schedule 80 for compressed air flow of 25.485 m3/min is 2.7720 kPa. This pressure loss is for gauge pressure of 1034 kPa and atmospheric pressure of 101.325 kPa that prevails at an altitude of 0 i.e. sea level. This must be adjusted to compressed air gauge pressure of 448000 Pa i.e. 448 kPa and atmospheric pressure of 70.109 kPa. From Table A4.2 we find that the pressure drop factor for 448 kPa and 70.109 kPa is 2.1913. Hence actual pressure drop for compressed air of gauge pressure of 448 kPa at an altitude of 3000 m will be 2.7720 * 2.1913, which equals to 6.0743 kPa.
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Table A4.1a Pressure loss for compressed air pressure of 1034 kPa in schedule 40 and schedule 80 seamless carbon steel pipes.
Flow volume m3/min
cfm
23.361 25.485 26.901 29.733 32.564 33.98 35.396 45.307 53.802 56.634 73.624 75.04 79.287 84.951 103.356 107.604
825 900 950 1050 1150 1200 1250 1600 1900 2000 2600 2650 2800 3000 3650 3800
Pressure loss in kpa for following nominal sizes of schedule 40 seamless carbon steel pipes
Pressure loss in kpa following nominal sizes of schedule 80 seamless carbon steel pipes
3"
3"
3.5"
4"
5"
6"
8"
10"
Actual internal diameter in mm 77.93 7.3817 8.7710 9.7637 11.9085 14.2652 15.5237 16.8354 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit
3.5"
4"
5"
6"
8"
10"
122.25
146.33
193.68 247.65
0.2814 0.3333 0.3703 0.4502 0.5379 0.5846 0.6333 1.0285 1.4429 1.5966 2.6816 2.7847 3.1054 3.5603 5.2528 5.6901
0.0673 0.0795 0.0882 0.1070 0.1275 0.1385 0.1499 0.2422 0.3387 0.3745 0.6265 0.6504 0.7248 0.8303 1.2224 1.3236
Actual internal diameter in mm
90.12
102.26
128.19
154.05
202.72
254.51 73.66
84.45
97.18
3.45750 4.10624 4.56968 5.57081 6.67055 7.25773 7.86967 12.84569 18.07587 20.01747 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit
1.7915 2.1266 2.3659 2.8827 3.4502 3.7532 4.0689 6.6351 9.3314 10.3321 17.4054 18.0777 20.1705 Unsuit Unsuit Unsuit
0.5560 0.6592 0.7329 0.8919 1.0663 1.1594 1.2564 2.0440 2.8706 3.1773 5.3436 5.5494 6.1900 7.0987 10.4810 11.3549
0.2162 0.2559 0.2843 0.3455 0.4127 0.4484 0.4857 0.7882 1.1052 1.2228 2.0525 2.1313 2.3765 2.7243 4.0181 4.3523
0.0534 0.0631 0.0700 0.0848 0.1010 0.1097 0.1187 0.1916 0.2679 0.2961 0.4950 0.5139 0.5726 0.6558 0.9650 1.0448
0.0170 0.0200 0.0222 0.0268 0.0319 0.0346 0.0374 0.0600 0.0836 0.0923 0.1536 0.1594 0.1775 0.2030 0.2980 0.3224
4.5629 5.4201 6.0325 7.3556 8.8090 9.5852 10.3940 16.9722 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit
2.3348 0.7104 2.7720 0.8425 3.0843 0.9367 3.7589 1.1403 4.4998 1.3637 4.8954 1.4829 5.3076 1.6071 8.6587 2.6160 12.1802 3.6752 13.4873 4.0682 Unsuit 6.8446 Unsuit 7.1084 Unsuit 7.9295 Unsuit 9.0944 Unsuit 13.4305 Unsuit 14.5510
9.9116 11.7788 13.1131 15.9963 19.1646 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit
0.0195 0.0230 0.0254 0.0308 0.0366 0.0397 0.0429 0.0689 0.0961 0.1061 0.1767 0.1834 0.2042 0.2336 0.3430 0.3712
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Book.indb 654
Table A4.1b Pressure loss for compressed air pressure of 1034 kPa in DN seamless carbon steel pipes. Flow Volume 3
m /min
Pressure loss in kpa for following diameters of seamless carbon steel pipes
cfm
80 mm
90 mm
100 mm
110 mm
120 mm
130 mm
140 mm 150 mm 160 mm 180 mm 190 mm 200 mm 225 mm 250 mm
23.361
825
6.43670
3.48158
2.01213
1.22690
0.78196
0.51725 0.35315 0.24778 0.17803 0.09760 0.07414 0.05716 0.03153 0.01857
25.485
900
7.64753
4.13487
2.38866
1.45584
0.92744
0.61319 0.41845 0.29345 0.21074 0.11542 0.08764 0.06753 0.03721 0.02189
26.901
950
8.51266
4.60155
2.65760
1.61932
1.03130
0.68167 0.46505 0.32603 0.23407 0.12812 0.09725 0.07492 0.04125 0.02425
29.733
1050 10.38188
5.60970
3.23844
1.97233
1.25552
0.82945 0.56558 0.39631 0.28437 0.15549 0.11797 0.09083 0.04995 0.02933
32.564
1150 12.43563
6.71714
3.87634
2.35992
1.50164
0.99163 0.67587 0.47338 0.33952 0.18547 0.14065 0.10825 0.05946 0.03488
33.98
1200 13.53234
7.30843
4.21688
2.56680
1.63298
1.07816 0.73471 0.51449 0.36892 0.20146 0.15274 0.11753 0.06452 0.03784
35.396
1250 14.67536
7.92466
4.57175
2.78237
1.76982
1.16831 0.79599 0.55730 0.39955 0.21810 0.16532 0.12719 0.06979 0.04091
45.307
1600 Unsuit
12.93557
7.45659
4.53419
2.88152
1.90036 1.29348 0.90469 0.64793 0.35292 0.26724 0.20537 0.11239 0.06571
53.802
1900 Unsuit
18.20245 10.48778
6.37422
4.04873
2.66865 1.81535 1.26892 0.90822 0.49406 0.37387 0.28712 0.15687 0.09155
56.634
2000 Unsuit
20.15767 11.61288
7.05708
4.48182
2.95366 2.00890 1.40398 1.00472 0.54636 0.41337 0.31739 0.17333 0.10111
73.624
2600 Unsuit
Unsuit
19.56562 11.88277
7.54167
4.96681 3.37570 2.35744 1.68572 0.91520 0.69184 0.53075 0.28921 0.16834
75.04
2650 Unsuit
Unsuit
20.32152 12.34138
7.83242
5.15806 3.50553 2.44799 1.75038 0.95020 0.71826 0.55099 0.30019 0.17470
79.287
2800 Unsuit
Unsuit
Unsuit
13.76894
8.73741
5.75334 3.90959 2.72978 1.95159 1.05911 0.80046 0.61395 0.33436 0.19451
84.951
3000 Unsuit
Unsuit
Unsuit
15.79433 10.02130
6.59779 4.48273 3.12946 2.23695 1.21354 0.91700 0.70320 0.38278 0.22256
103.356 3650 Unsuit
Unsuit
Unsuit
Unsuit
14.80050
9.74072 6.61555 4.61651 3.29849 1.78779 1.35029 1.03496 0.56264 0.32671
107.604 3800 Unsuit
Unsuit
Unsuit
Unsuit
16.03547 10.55278 7.16656 5.00064 3.57267 1.93607 1.46215 1.12059 0.60905 0.35356
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Appendix 4
655
Figure A4.1 Screen image of the worksheet (Calculation of compressed air pressure drop when the air flows through a steel pipe). Table A4.2 Pressure drop factor.
Book.indb 655
Altitude in m
Inverse of compression ratio for different gauge pressures of the compressed air Atmospheric pressures pressure in kPa 1034 862 758 689 448 414 379 350
0
101325
1.0000
1.1785
1.3212
1.4365
2.0668
2.2031
2.3637
2.5155
500
95465
1.0052
1.1858
1.3303
1.4473
2.0890
2.2285
2.3929
2.5486
1000
89875
1.0102
1.1927
1.3390
1.4576
2.1108
2.2532
2.4214
2.5810
1500
84552
1.0150
1.1994
1.3475
1.4677
2.1319
2.2772
2.4492
2.6126
2000
79495
1.0196
1.2059
1.3556
1.4773
2.1523
2.3006
2.4762
2.6434
2500
74670
1.0240
1.2121
1.3635
1.4867
2.1722
2.3233
2.5025
2.6734
3000
70109
1.0283
1.2180
1.3710
1.4956
2.1913
2.3452
2.5279
2.7025
3500
65768
1.0323
1.2237
1.3782
1.5042
2.2098
2.3664
2.5526
2.7307
4000
61640
1.0362
1.2292
1.3852
1.5125
2.2277
2.3869
2.5765
2.7581
4500
57722
1.0399
1.2344
1.3918
1.5204
2.2450
2.4068
2.5997
2.7846
5000
54020
1.0435
1.2394
1.3981
1.5280
2.2615
2.4258
2.6219
2.8101
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Appendix 5
Air pressure loss in hose pipes
Table 5.1a and 5.1b in this appendix give pressure loss in compressed air flow in hose pipes in their stretched condition. The method of calculation of pressure loss is the same as that for steel pipes except that Haaland’s equation for friction factor is used. All other conditions, except the absolute roughness of the hose pipe and pipe diameters, are the same. Since modern hose pipes are made from plastic and their inner surface is very smooth, the absolute roughness of these pipes is considered to be 0.000002 m. Hose pipes of all the sizes given in Table 5.1a and 5.1b are available in pressure ratings well above 1500 kPa. Pressure drop factors given in appendix 4 can also be applied to the pressure drop values given in this appendix.
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Table A5.1a Pressure loss for compressed air pressure of 1034 kPa in ASTM hose pipes. Flow volume
Pressure loss in kpa for following diameters of seamless carbon steel pipes
m3/min
73.03
cfm
76.20
88.90
101.6
114.30
127.00
146.33
152.40
168.28
193.68
202.72
203.20
219.08
254.00
23.361
825
6.04360
4.91769
2.33048
1.22199
0.69202
0.41636 0.21042 0.17304 0.10741 0.05465 0.04390 0.04340 0.03024 0.01487
25.485
900
7.08974
5.76793
2.73184
1.43182
0.81057
0.48755 0.24631 0.20253 0.12569 0.06393 0.05134 0.05076 0.03537 0.01739
26.901
950
7.83005
6.36951
3.01566
1.58014
0.89433
0.53783 0.27165 0.22335 0.13859 0.07048 0.05660 0.05596 0.03898 0.01916
29.733
1050
9.41271
7.65533
3.62191
1.89679
1.07311
0.64512 0.32571 0.26776 0.16610 0.08444 0.06780 0.06704 0.04669 0.02294
32.564
1150
11.12968
9.04993
4.27889
2.23974
1.26663
0.76122 0.38417 0.31579 0.19585 0.09953 0.07991 0.07901 0.05502 0.02702
33.98
1200
12.03868
9.78812
4.62645
2.42108
1.36893
0.82257 0.41506 0.34116 0.21156 0.10749 0.08630 0.08533 0.05941 0.02918
35.396
1250
12.98099 10.55327
4.98656
2.60892
1.47487
0.88610 0.44703 0.36742 0.22782 0.11574 0.09292 0.09187 0.06396 0.03141
45.307
1600
20.49907 16.65524
7.85416
4.10310
2.31685
1.39064 0.70076 0.57580 0.35676 0.18107 0.14533 0.14368 0.09999 0.04906
53.802
1900
Unsuit
Unsuit
10.78784
5.62929
3.17582
1.90488 0.95908 0.78787 0.48791 0.24747 0.19857 0.19632 0.13657 0.06696
56.634
2000
Unsuit
Unsuit
11.86208
6.18765
3.48987
2.09278 1.05340 0.86529 0.53577 0.27168 0.21799 0.21552 0.14991 0.07349
73.624
2600
Unsuit
Unsuit
19.30251 10.04933
5.65933
3.38966 1.70374 1.39897 0.86546 0.43837 0.35162 0.34763 0.24167 0.11836
75.04
2650
Unsuit
Unsuit
19.99900 10.41038
5.86199
3.51071 1.76439 1.44873 0.89618 0.45390 0.36406 0.35993 0.25022 0.12253
79.287
2800
Unsuit
Unsuit
Unsuit
11.52914
6.48973
3.88561 1.95217 1.60278 0.99128 0.50194 0.40257 0.39799 0.27665 0.13545
84.951
3000
Unsuit
Unsuit
Unsuit
13.10449
7.37326
4.41306 2.21625 1.81939 1.12497 0.56946 0.45667 0.45148 0.31378 0.15358
103.356 3650
Unsuit
Unsuit
Unsuit
18.87526 10.60621
6.34136 3.18067 2.61026 1.61278 0.81561 0.65389 0.64644 0.44908 0.21963
107.604 3800
Unsuit
Unsuit
Unsuit
20.34769 11.43034
6.83254 3.42611 2.81148 1.73683 0.87817 0.70400 0.69598 0.48345 0.23640
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Table A5.1b Pressure loss for compressed air pressure of 1034 kPa in DN hose pipes. Flow volume
Pressure loss in kpa for following diameters of seamless carbon steel pipes
m3/min
80
cfm
90
100
110
120
130
140
150
160
180
190
200
225
250
23.361
825
3.88412
2.19586
1.31939
0.83269 0.54724 0.37205 0.26035 0.18677 0.13691 0.07771 0.05993 0.04684 0.02661 0.01605
25.485
900
4.55480
2.57393
1.54602
0.97544 0.64089 0.43564 0.30479 0.21861 0.16023 0.09092 0.07011 0.05479 0.03111 0.01876
26.901
950
5.02924
2.84126
1.70622
1.07632 0.70707 0.48055 0.33617 0.24110 0.17669 0.10024 0.07729 0.06040 0.03430 0.02068
29.733
1050
6.04310
3.41227
2.04826
1.29164 0.84827 0.57637 0.40312 0.28905 0.21179 0.12012 0.09260 0.07236 0.04107 0.02476
32.564
1150
7.14242
4.03103
2.41873
1.52476 1.00109 0.68005 0.47553 0.34091 0.24975 0.14161 0.10915 0.08528 0.04840 0.02917
33.98
1200
7.72421
4.35835
2.61464
1.64801 1.08187 0.73483 0.51379 0.36831 0.26980 0.15295 0.11789 0.09210 0.05226 0.03149
35.396
1250
8.32716
4.69749
2.81757
1.77565 1.16551 0.79156 0.55340 0.39667 0.29056 0.16470 0.12694 0.09916 0.05626 0.03390
45.307
1600 13.13329
7.39778
4.43198
2.79035 1.83006 1.24203 0.86781 0.62170 0.45518 0.25780 0.19862 0.15511 0.08794 0.05295
53.802
1900 18.05705
10.15987
6.08129
3.82594 2.50772 1.70106 1.18800 0.85075 0.62265 0.35244 0.27147 0.21195 0.12010 0.07229
56.634
2000 19.86140
11.17118
6.68474
4.20463 2.75541 1.86877 1.30495 0.93438 0.68379 0.38698 0.29805 0.23268 0.13183 0.07933
73.624
2600 Unsuit
18.17481
10.85901
6.82164 4.46574 3.02608 2.11148 1.51089 1.10504 0.62476 0.48098 0.37534 0.21248 0.12778
75.04
2650 Unsuit
18.83033
11.24934
7.06617 4.62545 3.13409 2.18672 1.56464 1.14431 0.64691 0.49802 0.38863 0.21998 0.13229
79.287
2800 Unsuit
Unsuit
12.45887
7.82371 5.12012 3.46857 2.41968 1.73108 1.26587 0.71548 0.55075 0.42974 0.24321 0.14623
84.951
3000 Unsuit
Unsuit
14.16215
8.89006 5.81621 3.93912 2.74734 1.96511 1.43676 0.81184 0.62486 0.48751 0.27584 0.16582
103.356 3650 Unsuit
Unsuit
20.40254
12.79336 8.36225 5.65910 3.94433 2.81968 2.06054 1.16333 0.89507 0.69811 0.39472 0.23715
107.604 3800 Unsuit
Unsuit
Unsuit
13.78866 9.01104 6.09714 4.24903 3.03713 2.21921 1.25268 0.96375 0.75162 0.42492 0.25526
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Appendix 6
Air pressure loss in pipe fittings
Pressure loss occurs when a flow enters a flow pipe from a large reservoir. Resistance coefficients depend upon the geometry of the entrance as shown hereunder,
V
When the entrance starts inside the reservoir, as shown in the drawing on the left, the resistance coefficient KL is equal to 0.8 when L ≈ D
D L
Sharp edged entrance. r = 0.
r V
V
D
D
For rounded entrance the resistance coefficient depends upon the roundness as shown in following graph.
Resistance Coefficient
0.5 0.4 0.3 0.2 0.1 0.0 0.0
0.05
0.10
0.15
0.20
0.25
r/D
Figure A6.1 Resistance coefficient at entrance.
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L D
D
D
V
V
V r
Pressure loss also occurs when flow enters a large reservoir from flow pipe. However the resistance coefficients does not depend upon the geometry of the exit. For fully developed laminar flow the resistance coefficient is 2, whereas, for fully developed turbulent flow the resistance coefficient is 1.
Figure A6.2 Resistance coefficient at exit.
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Gradual Contraction
D
θ
V
Gradual Expansion
d
d
V
θ
D
Gradual Expansion 0.8
0.7
Resistance Coefficient
0.6 0.5
0.4 30°
0.3
0.2 20° 0.1
0.0 1.0
10° 5° 2.0
3.0
4.0
D/d Gradual Contraction 0.4
Resistance Coefficient
150° 0.3 120° 100° 0.2
90 75°
0.1
0.0 1.0
50°–60° 15°–40°
2.0
3.0
D/d
Figure A6.3 Resistance coefficient at gradual contraction and expansion.
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Pressure loss occurs when a the diameter of flow path suddenly increases or decreases.
V
d
D
D
In case of sudden expansion of the flow path, as shown in figure on the left, the resistance coefficient is calculated as KL = (1 – d2/D2)
V
For sudden contraction the pressure loss coefficient can be found from the following graph
d
0.6 0.5 0.4 KL
0.3 0.2 0.1 0.0 0.0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 0.9 1.0 d2/D2
Figure A6.4 Resistance coefficient for sudden expansion and contraction.
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Appendix 6
665
Tee with straight flow Flanged
: KL = 0.2
Threaded
: KL = 0.9
Tee with side flow Flanged
: KL = 1.0
Threaded
: KL = 2.0
Threaded Coupling : KL = 0.08
90° Mitre Bend (Without Vanes) : KL = 1.1
90° Mitre Bend (With Vanes) : KL = 0.2
Figure A6.5 Loss coefficient at direction control fittings.
For a 90° short radius bend Flanged
: KL = 0.3
Threaded : KL = 0.9
For a 45° sharp bend : KL = 0.4
For a 180° short radius bend Flanged
: KL = 0.2
Threaded : KL = 1.5
Figure A6.6 Resistance coefficient at direction control fittings.
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666
Appendix 6
Globe Valve, Fully Open : KL = 10
Angle Valve, Fully Open : KL = 2
Swing Check Valve : KL = 2
Gate Valve Fully Open : KL = 0.15 1/4 Closed : KL = 0.26 1/2 Closed : KL = 2.1 3/4 Closed : KL = 17
Figure A6.7 Resistance coefficient at flow control valves.
Table A6.1 Resistance coefficients for some pipe fittings.
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Sr. no.
Fitting type and status (if applicable)
Leq or KL
1 2 3 4 5 6 7 8
Ball Check Valve Butterfly Valve Foot Valve, Poppet Disc Type Foot Valve, Swing Disc Type Butterfly Valve Ball Valve, Fully Open Ball Valve, 1/3 Closed Ball Valve, 2/3 Closed
Leq = 150 Leq = 45 Leq = 420 Leq = 75 Leq = 45 KL = 0.05 KL = 5.5 KL = 200
11/22/2010 2:43:46 PM
Appendix 7
A note on tungsten carbide and other hard metals
INTRODUCTION When it comes to cutting tools, tungsten carbide is the most widely used material. Whether it is metal cutting in manufacturing plants or rock cutting in the mining industry, the tools almost invariably make use of tungsten carbide inserts at the cutting points. Use of tungsten carbide inserted bits has become so widespread that almost all the manufacturers who earlier used to make steel tooth bits have turned towards manufacturers of tungsten carbide inserted bits. This note is intended towards giving some useful details of tungsten carbide inserts that could not find an appropriate place in the main book.
PLACE OF TUNGSTEN CARBIDE IN SUPERHARD MATERIALS There are a few materials that find a place above tungsten carbide in hardness ranking. Some such materials are mentioned in Table A.7.1. However, their use in cutting tools has remained at a very low level for one of many reasons such as: 1 2 3 4
Their hardness is lost as their temperature increases by the heat generated in cutting. Their density is comparable to the materials on which they are used as cutting tools so their separation for reuse is difficult. They are relatively rare to find or difficult to make and hence their use is not viable economically. Tools made by using them cannot be satisfactorily shaped in the powder metallurgical processes.
Some of the ‘harder than tungsten carbide’ materials are used in making grinding tools or some are used as an additive in the tungsten carbide while forming a tungsten carbide tool itself. Materials having hardness comparable to tungsten carbide and used in cutting tools, as stated in previous paragraph, are listed Table A.7.2.
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668
Appendix 7
GENERAL INFORMATION ON TUNGSTEN CARBIDE Tungsten is found in earth’s crust in the form of Wolframite (Fe,Mn)WO4 and Scheelite CaWO4. With almost 80% of global production share, China is the largest producer of tungsten. Annual global production of tungsten is estimated at 28000 tons. Two grades of tungsten carbide viz. W2C and WC were discovered in the laboratory of the School of Pharmacy at the University of Paris, by H. Moissan and P. Williams in the years 1896 and 1898.
Table A7.1 Materials ‘harder than tungsten carbide’. Density g/cm3
Hardness GPa
Bulk modulus GPa
Shear modulus GPa
3.52 3.48
70 45–50 40–45
443 400 496
535 409 332
2.51 4.5 2.66
30–33 32–35
244 Not Known
263 Not Known
AlMgB14 + Add1
2.67
35–40
Not Known
Not Known
AlMgB14 + Add2
2.7
40–46
Not Known
Not Known
Name
Formula
Diamond Boron Nitride Carbon Nitride Boron Carbide Titanium Boride Aluminum Magnesium Boride Aluminum Magnesium Boride Aluminum Magnesium Boride
C BN C3 N4 B4C TiB2 AlMgB14
Table A7.2 Materials of hardness comparable to tungsten carbide.
Book.indb 668
Hardness
Crystal structure
Melting point (°C)
Theoretical density, g/cm3
Modulus of elasticity, GPa
Thermal expansion, mm/m⋅K
Mono Tungsten Carbide
2,200
Hexagonal
2,800
15.63
696
5.2
W2C
Di Tungsten Carbide
3,000
Hexagonal
2,777
17.3
–
–
TiC
Titanium Carbide
3,000
Cubic
3,100
4.94
451
7.7
VC
Vanadium Carbide
2,900
Cubic
2,700
5.71
422
7.2
HfC
Hafnium Carbide
2,600
Cubic
3,900
12.76
352
6.6
ZrC
Zircon Carbide
2,700
Cubic
3,400
6.56
348
6.7
NbC
Niobium Carbide
2,000
Cubic
3,600
7.8
338
6.7
Carbide formula
Carbide name
WC
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Appendix 7
669
Tungsten carbide is shaped by the powder metallurgical process. Cobalt is the main additive. It acts as a binder between two crystals of tungsten carbide. Some other additives are also used in the process to give the desirable properties to the product.
TUNGSTEN CARBIDE AND ROCK DRILLING As far as rock drilling is concerned, almost 95% of the inserts used in bits and other components of the drill string are made from mono tungsten carbide, i.e. WC. The only other viable alternative to tungsten carbide is titanium carbide i.e. TiC. The first instance of use of sintered tungsten carbide inserts for use in rock drilling goes back to 1914 but these experiments were not made public. The experiments must have shown promise but due to the infancy of rock drilling techniques themselves and attention being given towards the first world war situation, it was probably not taken a note of. The invention of cemented carbide tool materials was first disclosed in 1923 in Karl Schroter/s patent application. The two inventions claimed were: 1
2
A unique hard metal alloy composition, namely the combination of the very hard tungsten carbide, WC, with small amounts of iron group metals viz. Fe, Ni and Co. The manufacture of the hard metal alloys by the application of the process of powder metallurgy, namely pressing and sintering of mixed powders of tungsten carbide and binder metal.
Again the patent holder and drilling industry were probably unaware of the promises of the use of tungsten carbide inserts. The first use of carbide inserted bits for rock drilling was made by Demag in Germany in the year 1936. After the end of Second World War these wonderful tools were noticed by other countries in the world. With a gestation period of about a decade, carbide inserted bits became a common tool in rotary percussion drilling in the late 1950’s. Tricone bits with tungsten carbide inserts appeared in the 1960’s.
MANUFACTURE OF TUNGSTEN CARBIDE INSERTS A tungsten carbide insert is formed by a powder metallurgical process wherein sintering is one of the major operation. The complete process is in steps as under, 1 2 3 4 5 6
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Making Fine Metal Powder Mixing Different Powders Pressing the Powder in the Die Presintering Sintering Grinding
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670
Appendix 7
Making fine metal powder The objective of this activity is to make a powder of tungsten carbide and other additives so that they have appropriate particle size, shape, particle size distribution and purity. Methods used for making powders are: reduction of oxides or ores, electrolytic deposition, atomizing, mill grinding or thermal deposition. As a general rule, finer size particles give better strength of the inserts. Special techniques have to be used for obtaining ultra-fine powders of sub micron size.
Blending and mixing different powders For tungsten carbide inserts the main component is tungsten carbide. Cobalt is the most widely used binder. Occasionally other carbides such as TiC, TaC and NbC are also mixed with the tungsten carbide and cobalt to give the desired properties. When tungsten carbide inserts to be formed are larger size special type of wax is also added. Wax is necessary as a lubricant for achieving uniform material density within the insert. This operation is called blending. After blending the powders are required to be mixed so as to get uniform distribution of all the ingredients. Mixing is carried out in special mills having specific properties of the mixing vessel.
Pressing the powder mixture in a die The mixture of the powders is pressed into a die to get the desired shape of the final product with adequate and uniform density. In this operation blended and mixed powder is put into a die either under gravity or by arranging pre-formed tablets of the mixture. The dies are pressed by hydraulic presses. Pressure varies between 150 to 500 MPa depending upon the size and shape of the insert. The final product of the pressing operation is the raw carbide insert of size much larger than the final sintered product.
Pre-sintering Pre sintering is an operation where the insert is heated to low temperatures. This operation is particularly essential in case of tungsten carbide inserts because they contain wax. Pre-sintering removes the wax from the raw insert in the form of gases that are pumped out of the furnace. Pre-sintering does not impart any hardness to raw inserts. Sometimes machining operations are carried out on pre-sintered products as the final sintered product is too hard to be machined.
Sintering The sintering process involves heating the pre-sintered insert to high temperatures so that the binder melts and flows throughout the tungsten carbide matrix. In this operation porosity is removed and a uniform binding is obtained around the tungsten carbide. The volume of the insert considerably reduces in this operation – sometimes by
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Appendix 7
671
as much as 35%. When the sintered insert is subjected to controlled cooling it attains a hardness to the desired level. The micro-structure of a coarse-grained tungsten carbide insert is shown in Figure A7.1. The dark gray particles in the figure are of tungsten carbide and the light gray represents the cobalt binder which firmly holds the tungsten carbide particles together.
Coating Many different types of coatings are applied to the tungsten carbide insert for attaining the desired properties. For example, a polycrystalline diamond coating is given to inserts for extreme wear resistance, or a titanium carbide coating is given to the insert for reducing friction between the insert cutting tool and the metal to be cut. Some coatings are spray-applied on the raw insert after the pre-sintering stage but before final sintering. Certain coatings, particularly polycrystalline diamond coatings (in the raw graphite form), are sprayed after sintering and the insert is again sintered so that the graphite turns into diamond.
Grinding The sintered insert shape is required to be within very precise prescribed limits, particularly in the case of tungsten carbide inserts in button form, because they are shrink-filled into the bit. Precision shaping is carried out by grinding. Grinding wheels of boron nitride or diamond are used in such operations, And use of a coolant is essential.
Tungsten Carbide Grains
Cobalt Binder
Figure A7.1 Micro structure of tungsten carbide.
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672
Appendix 7
PROPERTIES OF TUNGSTEN CARBIDE INSERTS From the viewpoint of rock drilling the important properties of a tungsten carbide insert are hardness, toughness and wear resistance. The tungsten carbide phase in the insert is responsible for hardness and wear resistance. The cobalt phase imparts toughness to the insert. As can be seen from Table A7.3 the hardness of the insert and wear resistance, which is closely related to the hardness, decrease very rapidly with increasing cobalt contents but toughness – represented by the transverse rupture strength – increases. Particle size also plays a very important role in increasing the hardness. Table A7.4 sufficiently indicates the effect of using very fine particles of tungsten carbide and cobalt powder. In each of the two compositions of the mixture used for inserts, the hardness as well as the compressive strength of the insert have increased by using fine particles. Wear resistance has also increased very significantly by using finer particles.
TUNGSTEN CARBIDE INSERTS FOR DRILLING BITS Tungsten carbide inserted bits are used for both rotary and rotary percussion drilling. In rotary drilling the drilling bit and consequently the tungsten carbide inserts are subjected to very heavy but relatively steady feed force. This causes the formation in contact with the insert to fragment. On the other hand, in rotary percussive drilling the insert is subjected to impact loads that lead to formation fracturing. By combining proportions in the mixture of tungsten carbide and cobalt, particle size, additives etc. several grades of tungsten carbide inserts have been made by manufacturers all over the world. To avoid complications, the industry has developed a code for indicating the suitability of inserts for applications involving steady force or impact. From this code the code designations meant for drilling applications are given in Table A7.5. Many reputed manufacturers give an industry code for each type of insert manufactured by them. Table A7.6 gives such information for tungsten carbide inserts manufactured by Kennametal. Generally, tungsten carbide inserts used for drilling have cobalt contents between 6 to 18%. Some manufacturers make inserts by using techniques which enable them to vary the proportion of cobalt from the central portion of the insert to the outer side
Table A7.3 Effect of proportion of tungsten carbide and cobalt on the insert properties.
W
Co
Other
Hardness at room temp. HV
80 90
20 10
0 0
1,050 1,625
Cemented carbide
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Modulus of elasticity GPa
Transverse rupture strength MPa
Coefficient of thermal expansion 10−6/K
Thermal conductivity W/m⋅K
Density g/cm3
490 580
2,850 2,280
6.4 5.5
100 110
13.55 14.5
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Appendix 7
673
of the insert. This gives high hardness on the surface portion and high toughness at the center so that the insert becomes a dual property insert. Coating of Polycrystalline Diamonds and Tantalum Carbide on the tungsten carbide inserts is also used by some manufacturers.
Table A7.4 Effect of grain size of ingredients on the insert properties. Nominal composition
Hardness
Density
Transverse strength,
Compressive Modulus strength, of elasticity
WC
Cobalt Grain size
HRA
(g/cm3)
MPa
MPa
GPa
Relative wear resistance
97 94 94 94 90 90
3 6 6 6 10 10
92.5–93.2 92.5–93.1 91.7–92.2 90.5–91.5 90.7–91.3 87.4–88.2
15.3 15 15 15 14.6 14.5
1,590 1,790 2,000 2,210 3,100 2,760
5,860 5,930 5,450 5,170 5,170 4,000
641 614 648 641 620 552
100 100 58 25 22 7
Medium Fine Medium Coarse Fine Coarse
Table A7.5 Industry classification of sliding wear shocks and perpendicular impact shocks. Code designation
Suitability for wear application
Code designation
Suitability for impact application
C-9 C-10 C-11
No Sliding Wear shock Light Sliding Wear Shock Heavy Sliding Wear Shock
C-12 C-13 C-14
Light Perpendicular Impacts Medium Perpendicular Impacts Heavy Perpendicular Impacts
Table A7.6 Properties and industry codes for Kennametal tungsten carbide inserts. Trans. rupture strength
Density
Comp. Hardness strength
Designation WC
Grain Cobalt size
103xpsi
g/cm3
HRA
K96 K3833 K94 K3109 K92 K91 K3520 K90 KF306
94.5 89 88.5 87.9 84.3 80.5 80 75.2 94
5.5 11 11.5 12.1 15.7 19.5 20 24.8 6
300 430 410 430 475 440 430 430 390
14.85 14.4 14.2 14.2 13.8 13.4 13.45 13.8 14.95
KF310
90
10
500
KF315
85
15
550
Composition by% weight
Book.indb 673
Fine Fine Fine Fine Fine Fine Fine Fine Extra Fine Extra Fine Extra Fine
103xpsi
Relative impact resistance
Industry code
92.1 89.4 89.8 88 88.4 86.8 84.2 84.8 93.3
765 705 720 635 670 605 530 505 850
29 100 57 100 82 89 100 71 –
C-9©-10 C-11©-13 C-10©-12 C-11©-14 C-11©-13 C-11©-14 C-14 C-14 C-9
14.5
92.2
790
–
C-10©-12
13.9
90.4
660
–
C-11©-13
11/22/2010 2:43:54 PM
Appendix 8
Scales of hardness – Measurements and their conversion
Hardness of substances is determined by different types of tests. Minerals and metals are the two main types of substances on which such tests are carried out. Scales used for attributing hardness value to a mineral are: Moh’s Scale, New Moh’s Scale and Knopp’s Scale of Hardness. Of these three scales the Moh’s Scale is the oldest and has been used most widely in the scientific literature in the past. New Moh’s Scale has a few more steps of hardness. Knopp’s Scale is more rational in the relative hardness and therefore its use is ever increasing. Hardness levels of some of the common minerals are as given in the following table. The figures in the parenthesis after the name of some of the minerals indicate the crystallographic character of the crystals of minerals. It may be noted that some reference books give different values of hardness and corresponding minerals particularly in case of New Moh’s Scale of Hardness. Hardness of metals is determined by several tests. The hardness values attributed by these tests are different. Fortunately, due to systematic testing done in the past, the values of metal hardness can be correlated with each other. The table on the next page gives equivalent hardness values for most of the hardness scales. Tables on the next page give equivalent values of hardness by different hardness scales. The meanings of symbols used for explanation are as under. C A D B A1 A2 A3 K1 K2 S1 S2 S3
Book.indb 675
-
Rockwell C Scale with 150 kg load on the diamond brale, Rockwell A Scale with 60 kg load on the diamond brale, Rockwell D Scale with 100 kg load on the diamond brale, Rockwell B Scale with 100 kg load on the 1/16” ball, 15 kg load on 15 N penetrator, 30 kg load on 30 N penetrator, 45 kg load on 45 N penetrator, Diameter of the ball impression in mm, Hardness number, Vicker’s or Firth hardness number, Scleroscope hardness, Tensile Strength in kPa.
11/22/2010 2:43:54 PM
Book.indb 676
Table A8.1 Moh’s and New Moh’s scales of hardness. Moh’s scale of hardness
New Moh’s scale of hardness
Knopp’s scale of hardness
Hardness level
Corresponding mineral
Chemical formula
Hardness Corresponding level mineral
Chemical formula
1
Talc
Mg3Si4O10(OH)2
1
Talc (001)
Mg3Si4O10(OH)2 2.5
Talc (001)
Mg3Si4O10(OH)2
2
Halite (100)
NaCl
25
Halite (100)
NaCl
60
Gypsum
CaSO4⋅2H2O
80
Galena (100)
PbS
125
Calcite
CaCO3
2
Gypsum
CaSO4⋅2H2O
3
Calcite
CaCO3
4
Fluorite
CaF2
3
Galena (100)
PbS
Hardness level
Corresponding mineral
Chemical formula
4
Fluorite (100)
CaF2
175
Fluorite (100)
CaF2
5
Scheelite (111)
CaWO4
370
Scheelite (111)
CaWO4
525
Apatite
Ca5(PO4)3(OH,Cl,F)
600
Magnetite (111)
Fe3O4
730
Orthoclase
KAlSi3O8
5
Apatite
Ca5(PO4)3(OH,Cl,F)
6
Orthoclase
KAlSi3O8
7
Quartz
SiO2
7
Quartz(1011)
SiO2
1060
Quartz (1011)
SiO2
8
Topaz
Al2SiO4(OH,F)2
8
Topaz (001)
Al2SiO4(OH,F)2
1520
Topaz (001)
Al2SiO4(OH,F)2
9
Corundum
Al2O3
9
Corundum (1120)
Al2O3
2120
Corundum (1120)
Al2O3
10
Titanium Carbide
TiC
3000
Titanium Carbide
TiC
11
Boron
Pure Boron i.e. B
4000
Boron
Pure Boron i.e. B
6
10
Diamond
Pure Carbon i.e. C
Magnetite (111)
Fe3O4
11/22/2010 2:43:54 PM
12
Boron Carbide
B4C
4910
Boron Carbide
B4C
13
Boron Carbide
B6.5C
5750
Boron Carbide
B6.5C
14
Diamond Carbando
Pure Carbon i.e. C
8000
Diamond Carbando
Pure Carbon i.e. C
15
Industrial Diamond
Pure Carbon i.e. C
10000
Industrial Diamond
Pure Carbon i.e. C
Table A8.2 Equivalent values of hardness on different scales. Rockwell
Superficial Rockwell
Diamond brale C
A
D
80 79 78 77 76 75 74 73 72 71 70 69 68 67 66 65 64 63 62 61 60 59 58 57 56 55 54 53 52 51 50 49 48 47 46 45 44 43 42 41
92 92 91 91 90 90 89 89 88 87 87 86 86 85 85 84 84 83 83 82 81 81 80 80 79 79 78 77 77 76 76 75 75 74 73 73 73 72 72 71
87 86 85 84 83 83 82 81 80 80 79 78 77 76 76 75 74 73 73 72 71 70 69 69 68 67 66 65 65 64 63 62 61 61 60 59 59 58 57 56
1/16" “N” brale penetrator ball B A1 A2 A3 97 96 96 95 95 94 94 93 93 92 92 91 91 90 90 89 89 88 88 87 87 86 86 86 85 85 84 84 83 83 82 82 81
92 92 91 91 90 89 89 88 87 87 86 85 85 84 83 82 81 80 79 79 78 77 76 75 74 73 72 71 70 69 69 68 67 66 65 64 63 62 61 60
87 87 86 85 84 83 82 81 80 79 78 77 79 75 73 72 74 70 69 68 67 66 65 63 62 61 60 59 57 56 55 54 53 51 50 49 48 47 46 44
Brinell 10 mm ball 3000 kg load K1
2.25 2.30 2.30 2.35 2.35 2.40 2.45 2.55 2.55 2.60 2.60 2.65 2.70 2.75 2.75 2.80 2.85 2.90 2.90 2.95 3.00 3.00 3.05 3.10 3.10
K2
S1
S2
S3
745 710 710 682 682 653 627 578 578 555 555 534 514 495 495 477 461 444 444 432 415 415 401 388 388
1,865 1,787 1,710 1,633 1,556 1,478 1,400 1,323 1,245 1,160 1,076 1,004 942 894 854 820 789 763 746 720 697 674 653 633 613 595 577 560 544 528 513 498 484 471 458 446 434 423 412 402
99 98 97 96 95 93 91 88 87 85 83 82 80 78 77 75 74 72 71 69 68 67 65 64 63 62 61 59 58 56 55
2248 2172 2096 2027 1979 1924 1855 1800 1751 1689 1641 1600 1551 1510 1455 1420 1393 1365 1317 (Continued)
Book.indb 677
11/22/2010 2:43:54 PM
Table A8.2 (Continued). Rockwell Diamond brale
Book.indb 678
C
A
D
40 39 38 37 36 35 34 33 32 31 30 29 28 27 26 25 24 23 22 21 20 18 16* 14* 12* 10* 8* 6* 4* 2* 0*
70 70 69 69 68 68 67 67 66 66 65 65 64 64 63 63 62 62 62 61 61
55 55 54 53 52 52 51 50 49 48 48 47 46 45 45 44 43 42 42 41 40
1/16" ball B
109 109 108 108 107 106 106 105 104 103 103 102 101 100 99 99 98 97 95 94 92 90 89 87 85 84 82 81 79 77 74 72 70 68 65
Superficial Rockwell
Brinell
“N” brale penetrator
10 mm ball 3000 kg load
A1
A2
A3
K1
K2
S1
S2
S3
80 80 79 79 78 78 77 77 76 76 75 75 74 73 73 72 72 71 71 70 69
60 59 58 57 56 55 54 53 52 51 50 50 49 48 47 46 45 44 43 42 42
43 42 41 40 39 37 36 38 34 33 32 30 29 28 27 26 24 23 22 21 20
3.15 3.2 3.25 3.3 3.35 3.35 3.4 3.45 3.5 3.55 3.6 3.65 3.7 3.75 3.8 3.80 3.85 3.90 3.95 4.00 4.05 4.10 4.15 4.25 4.35 4.40 4.50 4.60 4.65 4.80 4.80 4.90 5.00 5.10 5.20 5.30 5.40 5.50 5.60
375 363 352 341 331 331 321 311 302 293 285 277 269 262 255 255 248 241 235 229 223 217 212 203 192 187 179 170 166 156 156 149 143 137 131 126 121 116 112
392 382 372 363 354 345 336 327 318 310 302 294 286 279 272 266 260 254 248 243 238 230 222 213 204 195 187 180 173 166 160 156 150 143 137 132 127 122 117
54 53 51 50 49 48 46 45 44 43 42 41 40 39 38 37 37 36 35 35 34 33 32 31 29 28 27 26 25 25 25 23 22 21
1276 1248 1213 1179 1158 1124 1096 1062 1034 1007 979 951 924 903 867 855 841 814 800 779 765 738 703 676 634 621 600 572 545 531 510 503 483 462 448 427 414 400 386
20 19 18 15
11/22/2010 2:43:55 PM
Appendix 9
Manufacturers of rotary blasthole drilling equipment
Table A9.1 gives a list of items frequently needed in rotary blasthole drilling. Names of manufacturers or suppliers are listed in Table 9.2 where in column 5, the products in their range are listed. Table A9.1 Symbols used for showing the items manufactured by the manufacturers listed in Table A9.2. Item
Book.indb 679
Symbol
Item
Symbol
Item
Symbol
Air Measuring Kit
AMK
Drill Pipe Integrated
DPI
Recovery Tools
RT
Bit Breaker
BB
Drill Stem Wrench
DSR
Shock Absorber External
SAE
Blasthole Camera
BC
Feed Force Measuring Kit
FFK
Shock Absorber Internal
SAI
Blasthole Dewatering Pump
BDP
Drill GPS
GPS
Stabilizers Integral Blade
SIB
Blasthole Inclinometer Drill Mounted
BID
Hoisting Plugs
HP
Stabilizers Roller
SR
Blasthole Inclinometer Handheld
BIH
Lifting Bails
LB
Stabilizers Replaceable Sleeve
SRS
Blasthole Plugs
BP
Laser Measuring Instruments
LMI
Stabilizer Welded Blade
SWB
Blasthole Sampler
BS
Rotary Deck Bushing
RDB
Tricone Bit Carbide Insert
TBC
Crossover Sub
CS
Rotary Blasthole Drill Large (Bit Load 351–551 kN)
RDL
Tricone Bit Steel Tooth
TBS
Drag Bit
DB
Rotary Blasthole Drill Medium (Bit Load 225–351 kN)
RDM
Miscellaneous Items
MI
Drill Computer
DCO
Rotary Blasthole Drill Small (Bit Load 100–225 kN)
RDS
Drill Loggers
DL
Rotary Blasthole Drill Truck/ Rubber Tired
RDT
Drill Pipe Fabricated
DPF
Rotary Blasthole Drill Ultra Large (Bit Load 551kN or More)
RDU
11/22/2010 2:43:55 PM
Table A9.2 Manufacturers of rotary blasthole drilling equipment. Sr. No. Manufacturer name
Book.indb 680
Manufacturer’s address
1
Aquila Drilling System
2
Atlas Copco Drilling Solutions 2100, North First Street, Garland Texas 75040, USA Inc.
100 North East Adams Street, Peoria. Illinois, 61629 USA
3
B. F. Carr & Associates
P.O. Box 220250, St Louis, Missouri63122–250, USA
4
BDS Inc.
P. O. Box 1160, Malta, Montana, 59538 USA
5
Bit Brokers International
P. O. Box 100, 5568 Logan Road, Logan, Illinois, USA
6
Blue Demon
P. O. Box 724, Lebanon, Georgia, 30146 USA
7
Bucyrus International Inc.
1100 Milwaukee Avenue, South Milwaukee, Wisconsin, 53172–0500, USA
8
Colcrete Eurodrill Co. Ltd.
Tower Business Park, Derby Road, Clay Cross, Derbyshire, S45 9 AG United Kingdom.
9
D. K. Thomas Equipment Corporation
P. O. Box 200, 1453 Route 9, Spofford, New Hampshire, 03462 USA
10
DataMetrics Corporation International
1717 Diplomacy Row, Orlando, Florida, 32809 USA
11
Diversified Drilling & Industrial Equipment F.Z.C.
Al Mamoura Building, Third Floor, Block (c), Beside of Ras Al Khaimah Immigration, Ras Al Khaimah, United Arab Emirates
12
Downhole Stabilization Inc.
P. O. Box 2467, 3515 Thomas Way, Bakersfield, California, 93308 USA
13
Driconeq Drilling Construction Equipment AB
P. O. Box 325, Svetsarevagen 4, Sunne 686 26, Sweden
14
Drill Pro International Inc.
108 E, 39th Street, Boise, Idaho, 83714 USA
15
Drill Supplies Inc.
1350 E. Glendale Avenue, Sparks, Nevada, 89431 USA
16
Dryrite Pumps
1854 Stevens Road, Eagle Point, Oregon, 97524 USA
17
DZBO
49100, Dnepropetrovsk, Udarnikov str., ap. 27
18
Eimco Elecon (I) Ltd.
Anand-Sojitra Road,Vallabh Vidyanagar, Gujarat 388120 India
19
Foremost Industries LP
1225 – 64th Avenue NE, Calgary, Alberta, T2E 8P9 Canada
20
Galleon Turbeco Flotek Company
11317 W. County Road. 128 Midland, Texas 79711, USA
21
Geophysical Instrument & Supply Co.
6323 Cambridge Street, Minneapolis, Minnesota, 55416 USA
22
Gill Rock Drill Co. Inc
903–905 Cornwall Road, Lebanon Road, Pennysylvania, 17042 USA
23
Globe Drill
8 Nambi Way, Kalgoorlie, Western Australia, 6430
24
Grant Prideco Drilling Products and Services
1450 Lake Robbins Dr. Suite 600, The Woodlands, Texas, 77380 USA
25
HAUSHERR System-Bohrtechnik GmbH & CO. KG
Heisenbergstr. 11, D-59423 Unna, West Germany
11/22/2010 2:43:56 PM
Tel. / Fax No.
Website
Products manufactured/sold
001 309 675 1000
www.cat.com/minestar
DCO, GPS
001 972 496 7400/
www.cmt.atlascopco.com
RDS, RDM, RDL, RDU, RDT, SR, CS, DSR, TBS, TBC, DPF, RDB, SAE
001 314 497 6075 /001 314 966 6985
www.stemplug.com
BP
001 406 654 1727 /001 406 654 2778
www.bds-usa.com
BDP
001 618 435 5811 /001 618 435 2388
www.bitbrokers.com
DB
001 770 591 2021 /001 770 591 6356
www.bluedemon.com
DB, TBC
001 414 768 4000 /001 414 768 4474
www.bucyrus.com
RDM, RDL, RDU, DPF
044 1246 868700 /044 1246 868701
www.colcrete-eurodrill.co.uk DPF
001 603 363 4706 /001 603 363 4249
www.driller.com
BD, BDG, BDP
001 407 251 4577 /407 251 4588
www.datametrics.com
GPS
00971 7 228 9223 / 00971 7 228 9233
No Website
SAI, SIB, SR, SRS, SWB
001 661 631 1044 /001 881 631 1195
No Website
SIB, SWB
046 565 121 10 /046 565 146 20
www.driconeq.com
CS, HP, LB, DPF, RB, SWB
001 208 336 4970 /001 208 345 1193
www.drillpro.net
RDB TBI, TBS, SIB, DPF
001 775 355 1199 /001 775 355 8775
www. americawestdrillingsupply. com
TBS, TBC, DB, CS, RT
001 541 830 1403 /001 541 830 0208
www.dryritepumps.com
BDP
038 056 7901157, 038 056 7901159
DPF, DPI, RT, CS
091 2692 227812 /091 2692 236506
www.emico.elecon.com
RDM
001 800 661 9190
www.foremost.ca
RDB, SAE
001 432 563 4172 / 001 432 563 1873
SR, SIB, SWB
001 952 929 8000 /001 952 926 5498
www.giscogeo.com
BC
001 717 272 3861 /001 717 272 4140
www.gillrockdrill.com
DPF, RT, CS,BP
061 8 9091 3008 / 061 8 9091 3009
www.globedrill.com.au
RDS
001 281 297 8500 /001 281 297 8525
www.grantprideco.com
DPF, DPI,
049 23 03–98624 /049 23 03–986241
www.hausherr.com
RDS, RDM, RDT, DPF
(Continued)
Book.indb 681
11/22/2010 2:43:56 PM
Table A9.2 Continued.
Book.indb 682
Sr. No. Manufacturer name
Manufacturer’s address
26
Heavy Engineering Corporation Ltd.
Dhurwa, Ranchi, Jharkhand, 834004 India
27
Henan Shenlong Petroleum Drilling Tools Co. Ltd
Technical Industrial Park, JiYuan, Henan, China
28
Ideal Supply
P. O. Box 15397, Asheville, North Carolina, 28813 USA
29
JSI Rocktools
11F, Jiulian Building, No. 213, North Baoshan Road, Guiyang 550004, Guizhou, China
30
Kay Rock Bit Company
P. O. Box 684865, Austin, Texas, 78768, USA
31
King Pin International
Hoffsj. Lovensky, 76B NO-0382, Oslo, Norway
32
Kingsland Drill International Inc.
7920 Ward Drive, Mercersburg, Pennysylvania, 17236 USA
33
Legra Pumps
9 Fursden Street, Glenella, Mackay, 4740 Queensland, Australia
34
Leica Geosystems
2700 E Executive Drive, Suite 100, Tucson, Arizona 85706, USA
35
LIM SA
Rue Jean Bartin,Villeurbanne 69100, Rhone, France
36
LMP Precision Engineering Co. Pvt. Ltd.
Plot No. 2P&3, GIDC, Antalia, Bilimora, Gujarat, 396325, India
37
Lone Star Drill Bits
P. O. Box 1219, Stafford, Texas,77497 USA
Marks Products Inc.
1243 Burnsville Road, Williamsville,VA 24487–2147, USA
M&P Flange and Pipe Protection
9426 Katy Fwy Bldg No. 11, Houston, Texas 77055, USA
38
Measurement Devices Ltd.
Acer House, Hackness Road, Northminster Business Park, Upper Poppelton,York,Yo26 6QR, United Kingdom
39
Mills Machine Company Inc.
P. O. Box 1514, Shawnee, Oklahoma, 74802 USA
40
Mining Technology International Inc
4400 North Big Spring, Bldg E, Suite 120, Midland, Texas 79705, USA
41
Modular Mining Systems Inc.
8289 E. Hemisphere Loop, Tucson, Arizona, 85706–5028, USA
42
MTI Group
3 Yeeda Way, Malaga, Western Australia 6090
43
NKMZ
NKMZ, Kramatorsk, Donetsk Region, 84305 Ukraine
44
OMZ
Ermolaevskiy per., 25, str. 1, Moscow, 123379
45
P&H Mining
4400 West National Avenue, Milwaukee, Wisconsin, 53214, U. S. A.
46
Pulsar Measuring Systems Ltd. 3, Forge Close, The Street, Chipperfield, Herts, WD4 9DL United Kingdom
47
R. Smith & Sons Inc.
2605 Chapman Street, Houston, Texas 77009,USA
48
Reamco Inc.
1149 Smede Highway, Broussard, Houston, Texas. 70518 USA
11/22/2010 2:43:56 PM
Tel. / Fax No.
Website
Products manufactured/sold
091 651 240 1249 /091 651 2401571
www.hecltd.com
RDM, DPF, CS
086 391 663 5039 /086 391 663 5036
www.petrodrillingtoos.com
DPF, DPI, SIB, CS
001 828 274 0144 /001 828 274 0140
www.idealsupply.com
MI
086 851 6780905 /086 851 6780960
www.jsintl.com.cn
DB, TBS, TBC
001 899 556 8665 /-
www.kayrockbitcompany. com
DB, DPF
047 2252 3638 /047 2252 3640
www.kingpin-international. com
RDB
- /001 717 328 4998
www.kingslanddrill.com
DP
061 7 4942 3922 /061 7 4942 3868
www.legra.com.au
BDP
001 520 529 8729 /-
www.leica-geosystems.com
GPS
033 472 146530
www.limgeo.com
DL, BID
091 2634 284714 /091 2634 281079
www.lmpprecision.com
RDS, RDM, DPF, CS,
- /001 877 817 0978
www.lonestarbits.com
DB, TBS, TBC
001 540–396–4740 /001540–396–4741 www.marksproducts.com
BC
001 800 8882605 /001 713 4636343
www.m-p.com
LB, MI
044 1904 791 139 /044 1904 791 532
www.mdl.co.uk
Ideal Supply
001 405 2734900 /001 405 273 4956
www.millsmachine.com
DB, SWB, CS, RT, LB, HP
001 432 684 4010 /001 432 684 4306
www.mti.com
DPF, DPI, SWB, SR, SAE, CS, LB, RDB
001 520 746 9127 /001 520 889 5790
www.modularmining.com
GPS
061 8 9209 2111 /061 8 209 2200
www.mtigroup.com.au
BP, MI
038 6264 37080 /038 6264 72249
www.nkmz.com
RDM, RDL, DPF
007 095 974 6021/007 095 796 9003
www.omz.ru
RDL,RDM, DPF
001 414 671 4400 /001 414 671 7236
www.phmining.com
RDL, RDU, DPF
044 1923 261412 /044 1923 261299
www.pulsarmeasuring.com
BIH
001713 2238495 /001 713 225–5015
www.r-smith-and-son.com
LB, MI
001 337 364 9244 /001 337 3649603
www.reamcoinc.com
SWB, SIB, SRS, DPF, DPI
(Continued)
Book.indb 683
11/22/2010 2:43:56 PM
Table A9.2 Continued. Sr. No. Manufacturer name
Book.indb 684
Manufacturer’s address
49
Reed Hycalog Canada
2700–144–4 th Avenue SW, Calgary, Alberta, T2P 3 N4 Canada
50
Revathi Equipment Limited
Pollachi Road, Coimbtore, Tamil Nadu 641021 India
51
Rotacan
400-B Kirkpatrick Street, North Bay, Ontario, P1B8G5 Canada
52
Rudgormash
53
Sandvik Mining and Construction
SE-811 81 Sandviken, Sweden
54
Schlumberger Ltd
300 Schlumberger Drive, Sugar Land, TX 77478, USA
55
SCHRAMM Inc.
800 E.Virginia Avenue, West Chester, Pennysylvania 19380, USA
56
Shanxi Oil Drilling Industry Central Manufacturing Co.
Jinzhong City, Shanxi Province, Chaoyang Shouyang County, China
57
Smith Drilling
11, Junction Road, Industries North, Germiston, Johannesburg, Gauteng, 2008 South Africa
58
Smith International Inc.
16740 Hardy Street, Houston, Texas, 77032 USA
59
Sphere Drilling Supplies
3112 -80th Avenue SE, Calgary, Albarta, T2C1 J3, Canada
60
Star Iron Works
P. O. Box 155, R. D. 3, Punxsutawney, Pennysylvania, 15767 USA
61
TeeKee
P. O. Box 313, 295 Prince Street, Despatch, Silverton 0127, South Africa.
62
Thread Protectors LLC
806 East Main St. Unit F, Broussard, La. 70518 USA
63
Thunderbird Mining Systems
2609 151st Place N. E., Redmond, WA, 98052–5522, USA
64
Torquato Drilling Accessories
215 Susquehanna Street, Old Forge, Pennsylvania 18519 USA
65
Total Engineering Pvt Ltd.
2 Stockdale Road, O’conner, 6163 Western Australia
66
Trimble Navigation Ltd
935 Stawart Drive, Sunnyvale, California, 94085 USA
67
TSK Bits
TB Tamachi Building, 5&6 Floor, No 11–1, Shiba 4-Chome, Minato-ku, Tokyo 108–0014 Japan
68
UDR Sampling System
332 Treasure Road, Welshpool, WA 6106, Australia
69
UralBMT
Verhnie Sergi, Near Ekaterinburg, Ural Mountain Region Russia
70
Varel International Inc.
P. O. Box 5401571343 Patton, Suite 106, Carrollton, Texas 75007 USA
71
Vulcan Industrial Engineering Co. Ltd.
A-2/440, G.I.D.C.,Vithal Udyognagar, Anand, Gujarat, 388121 India
72
Weatherford International Ltd.
515 Post Oak Blvd.,Suite 600, Houston, Texas 77027, USA
11/22/2010 2:43:56 PM
Book.indb 685
Tel. / Fax No.
Website
Products manufactured/sold
001 403 234 9999 /001 403 264 3300
www.reedhycalog.com
TBS, TBC, AMK
091 422 2610851 /91 422 2610427
www.revathi.co.in
RDS, RDM, RDL, DPF, CS
001 705 474 5858 /001 705 474 9595
www.rotacon.com
TBC
046 26 262000 /046 26 260222
www.sandvik.com
RDS, RDM, RDL, RDU, RDT, TBS, TBC, DPF, RDB
www.schlumberger.com
BB
001 610 696 2500 /001 610 696 6950
www.schramminc.com
RDS, DPF
086 354 4602448 /086 354 4602431
www.huanjie.com.cn
DPF, DPI, CS, SIB, SWB,
027 11 873 9830 /027 11 825 6830
www.smithdrilling.co.za
RDS, DPF, TBS, TBC, CS, SWB
001 281 443 3370 /001 281 233 5190
www.smithbits.com
TBS, TBC, AMK
001403 720 9333 /001 403 236 2010
www.sphere-drilling.com
RDB, DPF
001 814 427 2555 /001 800 927 0560
www.starironworks.com
DPF, SWB, CS,
027 12 803 7305 /027 12 803 8970
www.tee-kee.co.za
DPF, RDB, CS, SWB, RT, SAE
001 337 837–8881 /001 337 837–8885 www.threadproducts.com
HP, LB, MI
001 425 869 2727 /001 425 869 2735
www.thirdpac.com
GPS
001570 457 8565 /001570 457 8585
www.torquato.com
TBC
061 8 9314 5333 /061 8 9314 5336
www.totaldrillsupport.com
RDB, DPF
001 408 481 8000 /-
www.trimble.com
GPS
081 3 5730 1913 /081 3 5730 1915
www.tskbits.com
TBS, TBC
061 8 9351 9666 /061 8 9351 9555
www.udr.com.au
BS
001 630 690 5625 /001 630 560 5689
www.uralbmt.com
TBS, TBC
001 972 242 1160 /001 972 242 8770
www.varelintl.com
TBS, TBC, AMK, FFK
091 2692 236032 /091 2692 238723
www.viecl.com
DPF, DPI, CS, SR, SIB, SWB, RDB
001 713 693 4000
www.weatherford.com
SIB, SWB
11/22/2010 2:43:56 PM
Appendix 10
Details of rotary blasthole drills manufactured all over the world
This appendix gives important details of crawler-mounted rotary blasthole drills in Table A10.1 and carrier-mounted rotary blasthole drills in Table 10.2. The following points should be noted in respect of these details. 1
2
3 4
5 6
7
Book.indb 687
Most of the details presented have been checked. However, every manufacturer is in continuous quest for improvement in the product and therefore, details do change with every improvement in the blasthole drill. Details presented here may have changed even before you read this compilation. For most authentic and detailed specifications, the manufacturer of the drill should be contacted. Addresses of all the manufacturers have been given in a separate appendix. Most other specifications of rotary blasthole drills are fairly easy to understand and apprehend. However, since mast height (i.e. length), drill head, number of pipes in the pipe rack, wall thickness of the drill pipes etc. greatly affect the stability of the blasthole drill, one must be very careful in choosing the appropriate combinations of these items. Rather than merely relying on the specifications presented here or in the product brochure, the manufacturer must be consulted to ensure that what is desired from the drill is actually being fulfilled by the drill under consideration. In the following tables wherever some detail of the drill could not be obtained the specific place of the detail is marked with XXX. In the case of type of drill, the first letter indicates source of power and the second letter indicates the mode of power distribution. In this context E means electric, D means diesel and H means hydraulic Abbreviations used in the case of main power source are DE – Diesel Engine, SEL – single electric motor, MEL – many electric motors. In the tables the technical details have been given in International Units but exception has been made in case of weight of the blasthole drill, where more familiar unit viz. kg has been used instead of kN. Rotary drills having bit load capability less than 10000 kg are not included in this appendix because such drills are unsuitable for rotary drilling and are usually used for DTH drilling.
11/22/2010 2:43:57 PM
Table A10.1 Major details of crawler mounted rotary blasthole drills.
Max.
Min.
Max. Machine Weight
Rotary Speed Range
Max. Torque
Max. Bit Max. Feed Load Speed Nos.
Max. Discharge
Pressure
Main power source D = Diesel Engine, SE = Single Electric Motor, ME = Many Electric Motors
mm
mm
kg
RPM
Nm
kN
m/min
Nos.
m3/min
kPa
Type, kW
Blasthole Diameter
Largest compressors provided on the drill
Manufacturer
Model
Type E to E, E to H, D to E, D to H
Atlas Copco
PV 351
D/E to H
406
269
188240
0–87
25760
556
48.15
1
107.6
758
DE, 1230/SEL, 1043
Atlas Copco
DM -M3
D to H
311
251
104327
0–200
13806
400
43.9
1
73.6
758
DE, 709
Atlas Copco
PV275
D/E to H
269
171
83915
0–150
11795
334
38.71
1
73.6
758
DE, 708/SEL, 671
Atlas Copco
PV271
D/E to H
269
171
83915
0–150
11795
334
38.71
1
73.6
758
DE, 708/SEL, 671
Atlas Copco
PV235
D/E to H
251
152
58060
0–160
10575
289
58.8
1
53.8
758
DE, 597/SEL, 522
Atlas Copco
DML
D/E to H
270
149
49895
0–160
9762
267
44.5
1
53.8
758
DE, 596/SEL, 522
Atlas Copco
DML-SP
D to H
251
152
45360
Table 0–100
10169
245
60
1
53.8
758
DE, 596
Atlas Copco
DM50
D to H
254
152
41382
0–107
10575
222
69
1
34
758
DE, 418
Atlas Copco
DM45
D to H
254
149
38556
0–107
10575
200
44.5
1
29.7
758
DE, 380
Atlas Copco
DM30
D to H
171
127
28121
0–100
7321
133
30.48
1
25.5
758
DE, 317
Atlas Copco
DM25-SP
D to H
178
102
28121
Table 0–170
4746
111
21.9
1
25.5
758
DE, 317
Bucyrus International
59R
E to E
445
311
210920
0–120
21647
704
30.4
1
107.6
448
MEL, 896
Bucyrus International
49R
E to E
406
251
174663
0–125
16812
627
22.8
1
107.6
482
MEL, 896
Bucyrus International
39HR
E to H
349
229
149668
0–160
18845
543
47.8
1
85
310
SEL, 746
Bucyrus International
SKL – 20
D/E to H
381
229
131088
0–150
20880
489
37.5
1
102
689
DE, 899/SEL895
Bucyrus International
SKL – 16
D/E to H
381
229
131088
0–150
20880
489
37.5
1
102
689
DE, 899/SEL895
Bucyrus International
SKS – 16
D to H
311
229
84700
0–150
15185
411
33.5
1
68
689
DE, 784
Bucyrus International
SKS – 13
D to H
311
229
84700
0–150
15185
411
33.5
1
68
689
DE, 784
Bucyrus International
SKS – 10
D to H
311
229
84700
0–150
15185
411
33.5
1
68
689
DE, 784
Bucyrus International
SKF – 11
D to H
270
152
54583
0–220
12880
276
42.9
1
48.1
689
DE, 597
Bucyrus International
SKF – 10
D to H
270
152
54583
0–220
12880
276
42.9
1
48.1
689
DE, 597
Bucyrus International
SKF – 8
D to H
270
152
54583
0–220
12880
276
42.9
1
48.1
689
DE, 597
Bucyrus International
SKFX – 15
D to H
270
152
65026
0–220
12880
236
38.1
1
48.1
689
DE, 597
Bucyrus International
SKFX – 12
D to H
270
152
65026
0–220
12880
236
38.1
1
48.1
689
DE, 597
Changsha Keda CMCL
DR400
D to H
405
351
180000
0–150
25000
550
20
1
400
DE, 1500
Changsha Keda CMCL
DR370
D to H
351
269
136000
0–150
17500
490
18
1
84.9
700
DE, 1200
Changsha Keda CMCL
DR310
D to H
311
251
105000
0–150
14000
400
20
1
74.6
760
DE, 950
Changsha Keda CMCL
DR270
D to H
269
193
99000
0–150
6500
311
39
1
53.8
750
DE, 755
Emico Elecon
250 E
E to H
279
229
67000
0–97
14091
329
26.4
1
31.15
689
SEL, 373
Emico Elecon
160 D
D to H
203
127
34000
0–100
5000
142
20.8
1
17
689
DE,220
Globe Drill
KAL500
D to H
229
152
52000
0–125
5000
222
4.5
1
39
700
DE, 676
Hausherr
HBM160
D/E to H
190
130
40000
0–100
8000
160
23
1
30
700
DE, 470/SEL, XXX
Hausherr
HBM120
D/E to H
170
190
30000
0–134
6900
120
35
1
27
700
DE, 373/SEL, XXX
Heavy Eng. Corp.
HE5–250–4
E to E
270
250
90000
0–150
9316
294
1
1
28.31
689
MEL, 310
LMP Precision
SK100
E to H
279
200
65000
0–135
7912
311
3
1
34
689
DE, 373/SEL 373
LMP Precision
RBH-6–100 D/E
D/E to H
165
140
27000
0–135
4971
160
14
1
21.23
689
DE, 76 + DE, 224
NKMZ
SBSHS-250/270–32
E to E
270
243
90000
0–150
6000
400
13
1
40
689
MEL, 648
NKMZ
SBSHS-250H
E to E
250
229
78000
0–120
6500
350
12
1
32
689
MEL, 500
OMZ
SBSh-270-IZ
E to E
311
270
136000
0–120
13000
450
6
1
38
490
MEL, 648
OMZ
SBSh-270–34
E to E
270
229
141000
0–120
8400
350
6
1
38
490
MEL, 648
OMZ
SBSh-G-250
E to E
250
229
90000
0–150
5870
300
6
1
32
490
MEL, 500
P&H Mining
320XPC
E to E
445
311
165000
0–119
25760
667
24.4
1
109
448
MEL, 805
P&H Mining
250XP-ST
D to H
349
243
113400
0–200
16269
466
38.1
1
102
448
DE, 746
P&H Mining
250XP-DL
E to H
349
243
113500
0–200
16269
466
38.1
1
102
448
DE, 746
Revathi Equip’nt
C850E
E to H
349
311
130000
0–100
13050
445
15
1
45.3
689
SEM, 500
Revathi Equip’nt
C750
D/E to H
270
250
67000
0–95
11300
311
19
1
34
689
DE, 402/SEL, 373
Revathi Equip’nt
C650
D/E to H
200
160
40000
0–100
4520
133
12
1
32.5
2413
DE, 402/SEL, 373
Rudgormash
SBSH-250 MNA-32
E to E
270
160
85000
0–120
6500
350
1
1
32
689
Rudgormash
SBSH 160/200 40D
D to H
215
160
55000
0–120
XXX
235
XXX
1
25
780
DE, 485
Sandvik
1190 E
E to H
349
229
140614
0–97
16869
523
17
1
94.4
414
SEL, 746
Sandvik
D90 KS
D to H
349
229
104328
0–97
16869
523
21.6
1
85
552
DE, 839
Sandvik
DR 460
D to H
311
251
93440
0–175
10462
445
XXX
1
56.63
689
DE, 746
Sandvik
D75 KS
D to H
279
229
64864
0–94
14236
409
27
1
56.63
689
DE, 597
Sandvik
D50 KS
D to H
229
152
47627
0–126
9934
267
38
1
34.7
689
DE, 470
Sandvik
D45 KS
D to H
229
152
47627
0–126
9934
245
38
1
25.5
689
DE, 403
Sandvik
D55SP
D to H
254
172
79832
0–130
9934
232
35.4
1
56.63
689
DE, 597
Sandvik
D245S
D to H
203
127
35000
0–114
8282
209
21
1
34.7
689
DE, 470
Sandvik
D25 KS
D to H
171
127
33566
0–96
8282
142
32
1
32.8
689
DE, 470
Schramm
T450BH
D to H
171
114
21000
0–110
8518
110
49
1
29.7
2413
DE, 432
Appendix.indd 688
107
MEL, 500
11/22/2010 4:24:42 PM
For Largest Diameter Blasthole For maximum single pass combination
Nos.
Length each
Corresponding maximum hole depth
Tramming speed
Bailing velocity Indicator for for hard rock largest suitability hole (max. Bit with load/max. largest Hole dia.) comp.
No
m
m
km/h
kN/mm
m/min
1.77
1.370
2783
For maximum depth combination
Pipes in the Rack
Pipes in the Rack Corresponding maximum hole depth
Drill Pipe Diameter
m
m, mm
Drill pipe diameter
Single pass depth Nos.
mm
m
No, m
340
19.81
2
10.67
41.15
Same as for Max. Single Pass Combination
273
12.19
4
12.19
60.95
219
219
11.3
4
12.19
60.06
Same as for Max. Single Pass Combination
219
16.76
2
7.62
32
Same as for Max. Single Pass Combination
203
9.14
5
9.14
54.84
127
12.19
5
12.19
73.14
194
10.67
4
10.67
53.35
127
9.14
5
10.67
62.49
Kelly 178
18.29
0
0
18.29
Same as for Max. Single Pass Combination
178
9.14
2
9.14
27.42
114
9.14
5
9.14
178
9.14
2
9.14
27.42
114
9.14
5
9.14
127
7.9
5
7.6
45.9
Same as for Max. Single Pass Combination
Kelly 121
15.24
0
0
15.24
12.2 m 121 mm Kelly for Hole Depths Up to 12.2 m
340
19.81
2
9.14
38.09
Same as for Max. Single Pass Combination
1.45
340
21.3
2
9.14
38.09
Same as for Max. Single Pass Combination
273
16.5
1
30.2
220
16.5
5
340
20
2
9.14
38.28
273
20
5
340
16.5
1
13.41
29.91
273
16.5
235
16.46
2
7.62
31.7
235
16.5
235
13.4
1
12.19
25.59
235
235
10.3
1
10.67
20.96
178
11
1
10.67
178
10.1
1
178
8.6
178
Length each
Single pass depth
12.19
5
12.19
73.14
1.287
4223
1.77
1.240
3841
1.77
1.240
3841
1.77
1.151
3143
0.989
1943
0.975
2187
54.84
0.876
1319
54.84
0.788
1152
0.780
2476
0.625
1905
1.582
1662
1.45
1.545
2783
82.35
3.05
1.555
2289
9.14
65.7
2.09
1.283
4393
5
13.41
83.55
2.09
1.283
4393
4
7.62
46.98
2.01
1.322
2086
13.4
5
12.19
74.35
2.01
1.322
2086
235
10.3
5
10.67
63.65
2.01
1.322
2086
21.67
178
11
2
10.67
32.34
3.14
1.022
1486
10.67
20.77
178
10.1
4
10.67
52.78
3.14
1.022
1486
1
9.14
17.74
178
8.6
4
9.14
45.16
3.14
1.022
1486
15.85
2
7.62
31.09
178
15.8
2
7.62
31.04
2.09
0.874
178
12.8
1
10.67
23.47
178
12.8
4
10.67
55.48
2.09
0.874
1486
340
19.8
1
19.8
39.6
Same as for Max. Single Pass Combination
1.358
2813
273
19.8
1
19.8
39.6
Same as for Max. Single Pass Combination
1.396
2221
219
19.8
1
19.8
39.6
Same as for Max. Single Pass Combination
1.286
1948
178
19.8
1
19.8
39.6
Same as for Max. Single Pass Combination
1.156
1684
219
10.2
3
10.2
40.8
Same as for Max. Single Pass Combination
1.178
1327
127
9.14
13.7
13.7
1486
3
9.14
36.56
Same as for Max. Single Pass Combination
0.700
863
43.5
Same as for Max. Single Pass Combination
0.969
1693
152
11.5
4
8
140
14
1
14
28
140
8
6
8
56
0.842
2315
127
12
1
12
24
127
8
6
8
56
0.706
2692
200
8
3
8
32
Same as for Max. Single Pass Combination
1.090
1096
1.116
1449
219
10.67
3
10.67
42.68
Same as for Max. Single Pass Combination
127
10.67
2
6.1
22.87
127
6.1
5
6.1
36.6
0.971
2436
200
8
3
8
32
Same as for Max. Single Pass Combination
1.481
1548
200
11
2
11
33
Same as for Max. Single Pass Combination
1.400
1811
235
19
1
13
32
Same as for Max. Single Pass Combination
1.447
1166
235
17
1
17
34
Same as for Max. Single Pass Combination
1.296
2737
203
11
2
11
33
Same as for Max. Single Pass Combination
1.200
1914
381
21.3
1
21.3
42.6
391
273
19.81
2
19.8
59.41
Same as for Max. Single Pass Combination
273
12.2
5
12.2
73.2
235
1.335
2747
273
15.24
3
15.24
60.96
Same as for Max. Single Pass Combination
1.275
1220
200
10.67
3
10.67
42.68
Same as for Max. Single Pass Combination
1.152
1316
140
11
5
7.62
49.1
Same as for Max. Single Pass Combination
0.665
2028
200
8
3
8
32
Same as for Max. Single Pass Combination
1.296
1238
140
8.5
4
8.5
42.5
Same as for Max. Single Pass Combination
1.093
1196
273
19.81
0
0
19.81
273
12.2
6
12.2
85.4
1.499
2543
273
19.81
0
0
19.81
273
12.2
6
12.2
85.4
1.499
2289
244
16.76
2
7.62
32
244
12.8
3
12.8
51.2
1.430
1939
244
10.67
4
10.67
53.35
Same as for Max. Single Pass Combination
1.467
3939
178
9.14
4
9.14
45.7
Same as for Max. Single Pass Combination
1.166
2129
140
9.14
6
9.14
63.98
Same as for Max. Single Pass Combination
1.068
989
194
16.76
2
7.62
32
Same as for Max. Single Pass Combination
0.913
2682
140
8.53
4
9.14
45.09
Same as for Max. Single Pass Combination
1.030
2045
140
8.53
2
9.14
26.81
Same as for Max. Single Pass Combination
0.832
4332
102
7.62
5
7.62
45.72
Same as for Max. Single Pass Combination
0.643
2007
Appendix.indd 689
19.8 12.2
2 6
19.8 12.2
59.4 85.4
1.61
1.499
2625
1.9
1.335
2747
1.9
1.62
11/22/2010 4:04:42 PM
Table A10.2 Major details of truck mounted rotary blasthole drills.
Blasthole diameter
Manufacturer
Book.indb 690
Mode I
Largest compressors
Max.
Min.
Max. machine weight
Rotary speed range
Max. torque
Max. bit load
Max. feed speed
Nos.
Max. discharge
Pressure
Power of drill engine
mm
mm
kg
RPM
Nm
kN
m/min
Nos.
m3/min
kPa
kW 611
Atlas Copco
T4BH
254
191
26309
0-107
10575
133.4
14.63
1
35.4
2413
Reichdrill
T650
172
127
24948
0-108
6463
133.4
45.11
1
30.3
2585
433
Reichdrill
T750
311
222
45359
0-141
12924
311
61
1
48.1
689
522
Sandvik
DR320
172
115
24401
0-89
8204
143
41.4
1
28.3
2413
403
11/22/2010 2:43:59 PM
Largest drill pipe for largest blasthole Pipes in the rack
Book.indb 691
Indicator for hard rock suitability
Bailing velocity for largest hole with largest comp.
Carrier details
Drill pipe diameter
Single pass depth
Nos.
Length each
Maximum hole depth
mm
m
No
m
m
178
7.62
3
7.62
30.48
0.5252
1373
6×4 / 8×4
140
7.62
4
7.62
38.1
0.7755
3864
6×4
219
12.19
2
12.19
36.6
1.0000
1256
127
7.62
6
7.62
51.8
0.8313
2678
Overall drill dimensions (Mast Down)
Front axle capacity
Rear axle capacity
Length
Width
Height
kg
kg
m
m
m
Atlas Copco
9072
20865
10.7
2.4
4.12
Reichdrill
9072
20865
12.1
2.4
4.2
8×4
Reichdrill
18144
31751
18.5
4.3/Operating 5.6 5.6
6×4
Sterling LT9500
9072
20865
App 12.5
2.54
Type
Built by
m/min
4.1
11/22/2010 2:43:59 PM
Appendix 11
Details of drill pipes used for rotary blasthole drilling
Tables A11.1, A11.2 and A11.3 give dimension and weight details of rotary drill pipes commonly used inblasthole drilling. The weights presented are calculated from basic parameters such as dimension anddensity of steel etc. The first column of the tables indicates world-renowned manufacturers who make the drill pipes. Key for the manufacturers is as under 1 2 3 4 5
Book.indb 693
- Drillco International - Mining Technology International - Star Iron Works - Smith Drilling - Atlas Copco
11/22/2010 2:43:59 PM
Table A11.1 Fabricated drill pipes of diameter less than 250 mm.
Book.indb 694
Drill pipe O. diameter
Wall thickness
Cross sectional area
Moment of inertia
Producer
Inch
mm
In
mm
in2
mm2
in4
mm4
3 1, 2, 3 1, 2 1 1, 2, 3 1, 2, 3, 5 1, 2, 3 1, 3, 4, 5 5 3, 5 1, 2 5 1 4, 5 3 3 1, 2, 3 1, 3, 5 4 1, 2, 5 5 3 3, 5 1, 2, 4, 5 1 1 1, 2, 3, 5 1 5 5
3.5 4 4.5 4.5 5 5 5.5 5.5 6 6.25 6.25 6.25 6.5 6.625 6.75 7 7 7 7.5 7.625 7.625 7.625 8.625 8.625 8.625 9.25 9.25 9.25 9.25 9.625
88.9 101.6 114.3 114.3 127 127 139.7 139.7 152.4 158.75 158.75 158.75 165.1 168.28 171.45 177.8 177.8 177.8 190.5 193.68 193.68 193.68 219.08 219.08 219.08 234.95 234.95 234.95 234.95 244.48
0.75 0.5 0.375 0.5 0.5 0.75 0.5 0.75 0.75 0.5 0.75 1 0.75 0.864 0.75 0.5 0.75 1 1 0.75 0.875 1 0.906 1 1.5 0.75 1 1.5 1.5 1.5
19.05 12.7 9.525 12.7 12.7 19.05 12.7 19.05 19.05 12.7 19.05 25.4 19.05 21.946 19.05 12.7 19.05 25.4 25.4 19.05 22.225 25.4 23.012 25.4 38.1 19.05 25.4 38.1 38.1 38.1
6.48 5.498 4.86 6.283 7.069 10.014 7.854 11.192 12.37 9.032 12.959 16.493 13.548 15.637 14.137 10.21 14.726 18.85 20.42 16.199 18.555 20.813 21.97 23.955 33.576 20.028 25.918 36.521 36.521 38.288
4180.3 3546.9 3135.3 4053.7 4560.4 6460.5 5067.1 7220.6 7980.6 5827.1 8360.7 10640.9 8740.7 10088.5 9120.7 6587.2 9500.8 12161 13174.4 10450.8 11971 13427.7 14174.5 15454.6 21661.7 12921 16721.3 23561.9 23561.9 24702
6.6 8.6 10.4 12.8 18.1 23.3 24.8 32.4 43.5 37.6 49.9 58.9 56.9 66.3 64.6 54.2 72.9 87.2 110.4 96.8 107.5 116.8 165.9 177.1 222.5 182.3 223.7 284.5 284.5 326.7
2739124 3575547 4337824 5312241 7539295 9703778 10318006 13465764 18101205 15654509 20775203 24510372 23702043 27609704 26893219 22577023 30360222 36286690 45950882 40309899 44725060 48611153 69047485 73708936 92613522 75871823 93130220 118402950 118402950 135991287
11/22/2010 2:44:00 PM
Book.indb 695
Body weight
Drill pipe weight in kg for different lengths in m as under
lb/ft
kg/m
6.096
7.62
9.144
10.668
12.192
13.716
15.24
16.764
22.05 18.71 16.54 21.38 24.05 34.08 26.73 38.08 42.09 30.73 44.1 56.12 46.1 53.21 48.11 34.74 50.11 64.14 69.49 55.12 63.14 70.82 74.76 81.51 114.25 68.15 88.19 124.27 124.27 130.29
32.84 27.87 24.63 31.85 35.83 50.76 39.81 56.73 62.7 45.78 65.69 83.6 68.67 79.26 71.66 51.75 74.64 95.55 103.51 82.11 94.05 105.5 111.37 121.42 170.19 101.52 131.38 185.12 185.12 194.08
200 170 150 194 218 309 243 346 382 279 400 510 419 483 437 315 455 582 631 501 573 643 679 740 1037 619 801 1128 1128 1183
250 212 188 243 273 387 303 432 478 349 501 637 523 604 546 394 569 728 789 626 717 804 849 925 1297 774 1001 1411 1411 1479
300 255 225 291 328 464 364 519 573 419 601 764 628 725 655 473 683 874 946 751 860 965 1018 1110 1556 928 1201 1693 1693 1775
350 297 263 340 382 542 425 605 669 488 701 892 733 846 764 552 796 1019 1104 876 1003 1125 1188 1295 1816 1083 1402 1975 1975 2070
400 340 300 388 437 619 485 692 764 558 801 1019 837 966 874 631 910 1165 1262 1001 1147 1286 1358 1480 2075 1238 1602 2257 2257 2366
450 382 338 437 491 696 546 778 860 628 901 1147 942 1087 983 710 1024 1311 1420 1126 1290 1447 1528 1665 2334 1392 1802 2539 2539 2662
500 425 375 485 546 774 607 865 956 698 1001 1274 1047 1208 1092 789 1138 1456 1577 1251 1433 1608 1697 1850 2594 1547 2002 2821 2821 2958
551 467 413 534 601 851 667 951 1051 767 1101 1401 1151 1329 1201 868 1251 1602 1735 1376 1577 1769 1867 2035 2853 1702 2202 3103 3103 3254
11/22/2010 2:44:00 PM
696
Appendix 11
Table A11.2 Fabricated drill pipes of diameter more than 250 mm. Drill pipe O. diameter
Wall thickness
Cross sectional area
Moment of inertia
Producer
Inch
mm
In
mm
in2
mm2
in4
mm4
1, 2, 3, 4, 5 5 1, 2 1, 2 5 5 1, 2, 4 1, 2 4, 5 1, 2, 5 2 2 2 2 4 4
10.75 10.75 10.75 10.75 11.75 12.25 12.75 13.375 13.375 13.375 14 14 14.375 14.375 15 15
273.05 273.05 273.05 273.05 298.45 311.15 323.85 339.73 339.73 339.73 355.6 355.6 365.13 365.13 381 381
1 1.25 1.5 2 1.25 1 1 1 1.25 1.5 1.5 2 1.5 2 1.25 1.5
25.4 31.75 38.1 50.8 31.75 25.4 25.4 25.4 31.75 38.1 38.1 50.8 38.1 50.8 31.75 38.1
30.631 37.306 43.59 54.978 41.233 35.343 36.914 38.877 47.615 55.96 58.905 75.398 60.672 77.754 53.996 63.617
19761.6 24068.6 28122.2 35469.5 26602.1 22801.8 23815.2 25082 30719.1 36102.9 38003 48643.9 39143.1 50164 34836.1 41043.3
367.8 428.1 478.5 553.6 576.3 563.6 641.7 749.1 884.3 1002.1 1167.1 1394.9 1274.2 1527.3 1286.6 1467.2
153092137 178209075 199151563 230443981 239874485 234568629 267080564 311786071 368078982 417120238 485763550 580587047 530374401 635708568 535533882 610682934
Table A11.3 Integral drill pipes.
Book.indb 696
Drill pipe diameter d
Internal diameter
Cross sectional area
Moment of inertia
Producer
Inch
mm
In
mm
in2
mm2
in4
mm4
1 1 1 1 1 1 1 1 1 1 1 1 1
4 4.5 5 5.5 6.25 6.5 7 7.625 8.625 9.25 10.75 12.75 13.375
101.6 114.3 127 139.7 158.75 165.1 177.8 193.68 219.08 234.95 273.05 323.85 339.73
1.5 1.75 2.063 2.313 2.813 2.813 2.813 2.813 3.25 3.25 4.75 4.75 4.75
38.1 44.45 52.388 58.738 71.438 71.438 71.438 71.438 82.55 82.55 120.65 120.65 120.65
10.799 13.499 16.294 19.558 24.467 26.97 32.272 39.451 50.13 58.905 73.042 109.956 122.78
6967.2 8709 10512.2 12618.2 15785.1 17400.2 20820.5 25452.1 32342.2 38003 47123.8 70939 79212.6
12.3 19.7 29.8 43.5 71.8 84.6 114.8 162.9 266.2 353.9 630.6 1272.2 1545.9
5127078 8186645 12400083 18111984 29897870 35193430 47778077 67787495 110788796 147299752 262457893 529538457 643451068
11/22/2010 2:44:01 PM
Appendix 11
Body weight
Drill pipe weight in kg for different lengths in m as under
lb/ft
kg/m
12.192
13.716
15.24
16.764
18.288
19.812
20.117
21.336
104.23 126.94 148.33 187.08 140.31 120.26 125.61 132.29 162.02 190.42 200.44 256.56 206.45 264.58 183.74 216.47
155.26 189.1 220.95 278.68 209.01 179.15 187.11 197.06 241.35 283.65 298.58 382.18 307.54 394.13 273.7 322.47
1893 2306 2694 3398 2548 2184 2281 2403 2943 3458 3640 4660 3750 4805 3337 3932
2130 2594 3031 3822 2867 2457 2566 2703 3310 3891 4095 5242 4218 5406 3754 4423
2366 2882 3367 4247 3185 2730 2852 3003 3678 4323 4550 5824 4687 6007 4171 4914
2603 3170 3704 4672 3504 3003 3137 3304 4046 4755 5005 6407 5156 6607 4588 5406
2839 3458 4041 5096 3822 3276 3422 3604 4414 5187 5460 6989 5624 7208 5005 5897
3076 3746 4377 5521 4141 3549 3707 3904 4782 5620 5915 7572 6093 7809 5423 6389
3123 3804 4445 5606 4205 3604 3764 3964 4855 5706 6007 7688 6187 7929 5506 6487
3313 4035 4714 5946 4459 3822 3992 4204 5149 6052 6371 8154 6562 8409 5840 6880
Weight
Book.indb 697
697
Drill Pipe Weight in kg for Different Lengths in m as Under
lb/ft
kg/m
7.62
9.144
10.668
12.192
36.75 45.93 55.45 66.55 83.26 91.77 109.81 134.24 170.58 200.44 248.55 374.15 417.79
54.74 68.42 82.59 99.14 124.02 136.71 163.58 199.97 254.1 298.58 370.24 557.35 622.36
417 521 629 755 945 1042 1246 1524 1936 2275 2821 Not Made Not Made
501 626 755 907 1134 1250 1496 1829 2323 2730 3385 Not Made Not Made
584 730 881 1058 1323 1458 1745 2133 2711 3185 Not Made Not Made Not Made
667 834 1007 1209 1512 1667 1994 2438 3098 Not Made Not Made Not Made Not Made
11/22/2010 2:44:02 PM
Appendix 12
Details of stabilizers and other miscellaneous items
The details of various types of stabilizers given in this appendix are for the products made by the specific manufacturers mentioned in each case. There may be significant differences between such dimensions for similar products made by other manufacturers. Even for the same manufacturer the dimensions must be checked from the manufacturers before taking any important decision.
Shoulder to Shoulder Length
Drill Bit
Drill Pipe
Body Diameter
Blasthole Diameter
Figure A12.1 Sketch of Atlas Copco roller stabilizer.
Table A12.1 Details of Atlas Copco roller stabilizer. Blasthole diameter Body diameter
Book.indb 699
Sh. to Sh. length
Inches
mm
Inches
mm
Inches
mm
9.000 9.875 10.625 11.000 12.250 13.750 15 16
228.60 250.82 269.88 279.40 311.15 349.25 381 406.4
7 or 7.625 7.625 or 8.625 8.625 or 9.250 8.625 or 9.250 9.625 or 10.750 10.750 to 11.750 11.750 to 13.375 13.375
177.8 or 193.7 193.7 or 219.1 219.1 or 235 219.1 or 235 244.5 to 273.1 273.1 or 298.5 298.5 or 339.7 339.7
32 or 42 32 or 42 or 50 32 or 42 or 50 32 or 42 or 50 32 or 42 or 50 32 or 42 or 50 42 or 50 42 or 50
812.8 or 1066.8 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 1066.8 or 1270 1066.8 or 1270
11/22/2010 2:44:03 PM
700
Appendix 12
Shoulder to Shoulder Length
Drill Pipe
Body Diameter
Drill Bit
Blasthole Diameter
Figure A12.2 Sketch of MTI roller stabilizer. Table A12.2 Details of MTI roller stabilizer. Blasthole diameter
Body diameter
Inches
mm
Inches
6.250 6.750 7.875
158.75 171.45 200.025
9.000 9.875 10.625
228.6 250.825 269.875
11.000
279.4
12.250 13.750 15.000
311.15 349.25 381
5.125 5.625 6.250 or 7.000 7.625 8.625 8.625 or 9.25 9.250 or 9.625 10.750 12.250 13.375 or 14.375
Sh. to Sh. length
Weight
mm
Inches
lbs
130.18 142.88 158.75 or 177.8 193.68 219.08 219.08 or 234.95 234.95 or 244.48 273.05 311.15 339.72 or 365.12
26.500 26.500 26.500
673.1 673.1 673.1
26.500 30.750 30.750
673.1 781.05 781.05
30.750
781.05
30.750 41.000 41.000
781.05 1041.4 1041.4
130 165 200 or 220 290 400 400 or 460 460 or 510 650 1000 1300
kg 59 75 91 132 182 182 or 209 209 or 232 295 454 590
Shoulder to Shoulder Length
Body Diameter
Blasthole Diameter
Figure A12.3 Sketch of MTI welded straight blade stabilizer.
Book.indb 700
11/22/2010 2:44:03 PM
Appendix 12
701
Table A12.3 Details of MTI welded straight blade stabilizer. Blasthole Diameter
Body Diameter
Inches
mm
Inches
6.250 6.750 7.875
158.75 171.45 200.025
9.000 9.875 10.625
228.6 250.825 269.875
11.000
279.4
12.250 13.750 15.000
311.15 349.25 381
5.125 5.625 6.250 or 7.000 7.625 8.625 8.625 or 9.25 9.250 or 9.625 10.750 12.250 13.375 or 14.375
Sh. to Sh. Length
Weight
mm
Inches
lbs
130.18 142.88 158.75 or 177.8 193.68 219.08 219.08 or 234.95 234.95 or 244.48 273.05 311.15 339.72 or 365.12
26.500 26.500 26.500
673.1 673.1 673.1
145 185 250
66 84 113
28.500 30.750 30.750
673.1 781.05 781.05
350 470 520
159 213 236
30.750
781.05
580
263
30.750 41.000 41.000
781.05 1041.4 1041.4
790 1100 1500
358 499 680
kg
A C
D
B E
F H
Figure A12.4 Sketch of IDS welded straight blade stabilizer.
Table A12.4 Details of IDS welded straight blade stabilizer. H
Book.indb 701
A
B
C
D
E
F
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
in
Mm
6.250 6.750 7.875 9.000 9.875 10.625 11.000 12.250 13.750 15.000
158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381
60 60 61 61 61 64 64 64 66 66
1524 1524 1549.4 1549.4 1549.4 1625.6 1625.6 1625.6 1676.4 1676.4
28 28 28 28 28 28 28 28 28 28
711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2
18 18 18 18 18 18 18 18 18 18
457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2
14 14 15 15 15 18 18 18 20 20
355.6 355.6 381 381 381 457.2 457.2 457.2 508 508
12 12 12 12 12 12 12 12 14 14
304.8 304.8 304.8 304.8 304.8 304.8 304.8 304.8 355.6 355.6
1.5 1.5 2 2 2 2 2 2 2 2
38.1 38.1 50.8 50.8 50.8 50.8 50.8 50.8 50.8 50.8
11/22/2010 2:44:04 PM
702
Appendix 12
A C
D
B E
F
H
Figure A12.5 Sketch of IDS welded spiral blade stabilizer.
Table A12.5 Details of IDS welded spiral blade stabilizer. H
A
B
C
D
E
F
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
6.250 6.750 7.875 9.000 9.875 10.625 11.000 12.250 13.750 15.000
158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381
60 60 61 61 61 64 64 64 66 66
1524 1524 1549.4 1549.4 1549.4 1625.6 1625.6 1625.6 1676.4 1676.4
28 28 28 28 28 28 28 28 28 28
711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2
18 18 18 18 18 18 18 18 18 18
457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2
14 14 15 15 15 18 18 18 20 20
355.6 355.6 381 381 381 457.2 457.2 457.2 508 508
12 12 12 12 12 12 12 12 14 14
304.8 304.8 304.8 304.8 304.8 304.8 304.8 304.8 355.6 355.6
1.5 1.5 2 2 2 2 2 2 2 2
38.1 38.1 50.8 50.8 50.8 50.8 50.8 50.8 50.8 50.8
F D
G
E
C
B
A
H
Figure A12.6 Sketch of integral spiral blade stabilizer.
Book.indb 702
11/22/2010 2:44:06 PM
Book.indb 703
Table A12.6 Details of integral spiral blade stabilizer. A
B
C
D
E
H
F
G
Weight
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
lb
kg
6.25 6.75 7.875 9 9.875 10.625 11 12.25 13.75 15 17.5
158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381 444.5
4.75 4.75 6.75 8 8 8 8 9 10 10 10
120.65 120.65 171.45 203.2 203.2 203.2 203.2 228.6 254 254 254
2 2.25 2.25 2.25 2.8125 2.25 2.8125 2.8125 2.8125 2.8125 3
50.8 57.15 57.15 57.15 71.4375 57.15 71.4375 71.4375 71.4375 71.4375 76.2
28 28 28 28 28 30 30 29 31 31 31
711.2 711.2 711.2 711.2 711.2 762 762 736.6 787.4 787.4 787.4
18 17 18 18 18 20 20 19 20 20 20
457.2 431.8 457.2 457.2 457.2 508 508 482.6 508 508 508
2 2 2.5 2.5 2.5 3.25 3.25 3.25 4.5 4.5 4.5
50.8 50.8 63.5 63.5 63.5 82.55 82.55 82.55 114.3 114.3 114.3
62 62 64 63 63 69 69 69 73 73 74
1574.8 1574.8 1625.6 1600.2 1600.2 1752.6 1752.6 1752.6 1854.2 1854.2 1879.6
13 13.25 15 15.25 14 14.75 14 16 17.5 16 14.75
330.2 342.9 381 387.35 355.6 374.65 355.6 406.4 444.5 406.4 374.65
280 370 620 840 840 1000 1000 1550 1670 1670 2100
127.0 167.8 281.2 381.0 381.0 453.6 453.6 703.1 757.5 757.5 952.5
Table A12.7 Details of integral straight blade stabilizer. A
B
C
D
E
H
F
G
Weight
11/22/2010 2:44:08 PM
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
in
mm
lb
kg
6.25 6.75 7.875 9 9.875 10.625 11 12.25 13.75 15 17.5
158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381 444.5
4.75 4.75 6.75 8 8 8 8 9 10 10 10
120.65 120.65 171.45 203.2 203.2 203.2 203.2 228.6 254 254 254
2 2.25 2.25 2.25 2.8125 2.25 2.8125 2.8125 2.8125 2.8125 3
50.8 57.15 57.15 57.15 71.4375 57.15 71.4375 71.4375 71.4375 71.4375 76.2
28 28 28 28 28 30 30 29 31 31 31
711.2 711.2 711.2 711.2 711.2 762 762 736.6 787.4 787.4 787.4
18 17 18 18 18 20 20 19 20 20 20
457.2 431.8 457.2 457.2 457.2 508 508 482.6 508 508 508
2 2 2.5 2.5 2.5 3.25 3.25 3.25 4.5 4.5 4.5
50.8 50.8 63.5 63.5 63.5 82.55 82.55 82.55 114.3 114.3 114.3
62 62 64 63 63 69 69 69 73 73 74
1574.8 1574.8 1625.6 1600.2 1600.2 1752.6 1752.6 1752.6 1854.2 1854.2 1879.6
13 13.5 15 15.25 14 14.75 14 16 17.5 16 14.75
330.2 342.9 381 387.35 355.6 374.65 355.6 406.4 444.5 406.4 374.65
280 370 620 840 840 1000 1000 1550 1670 1670 2100
127.0 167.8 281.2 381.0 381.0 453.6 453.6 703.1 757.5 757.5 952.5
Book.indb 704
Sleeve Body Diameter
Sleeve Bore Blade Width Diameter Sleeve End Diameter Sleeve
Mandrel Upset Diameter
Fishing Neck Diameter
Mandrel
Crown Length Sleeve End Length
Sleeve Length
Fishing Neck Length
Total Overall Length Blasthole Diameter
Bore ID (Near
Bore ID (String)
Figure A12.8 Sketch of replaceable sleeve stabilizer.
11/22/2010 2:44:09 PM
Book.indb 705
Table A12.8 Details of sleeve of the DTI replaceable sleeve stabilizer. Blasthole diameter
Sleeve Body diameter
Sleeve ID
Sleeve size designation
in
mm
in
mm
in
475 625 775 775 963
6.25–6.75 8.375–9.875 8.875–12.25 9.625–17.5 12.25–17.5
158.75–171.45 212.725–250.825 225.425–311.15 244.475–444.5 311.14–444.5
5.75 7.5 9.25 10 11
146.05 190.5 234.95 254 279.4
4.75 6.25 7.75 7.75 9.625
Sleeve length
Blade width
Minimum crown length
Weight of sleeve
mm
in
mm
in
mm
in
mm
lb
kg
120.65 158.75 196.85 196.85 244.475
14 14 18 18 18
355.6 355.6 457.2 457.2 457.2
2 2.5 3 3 3.5
50.8 63.5 76.2 76.2 88.9
12.2 11.1 14.6 11.1 11.9
309.88 281.94 370.84 281.94 302.26
43 84 162 294 253
19.50 38.10 73.48 133.36 114.76
Table A12.8 A Details of mandrel of the DTI relatable sleeve stabilizer. Maximum fishing neck diameter
Mandrel size designation in
mm
Mandrel upset diameter
Sleeve end diameter
Fishing neck length
Total Sleeve end overall length length
Bore ID (near bit)
Bore ID (string)
in
in
mm
in
in
in
mm
in
mm 57.15
mm
mm
mm
in mm
Weight of mandrel lb
kg
47
4.75–5
120.65– 5.125– 127 5.75
130.175– 4.75 146.05
120.65
23 584.2 32 812.8 62 1574.8 1.5
38.1
2.25
302 136.98
62
6.5–7
165.1– 177.8
7.5
190.5
6.25
158.75
23 584.2 32 812.8 62 1574.8 2.25
57.15
2.8125 71.4375
500 226.80
77
8.25
209.55
9.25
234.95
7.75
196.85
27 685.8 37 939.8 71 1803.4 2.8125 71.4375 2.8125 71.4375
893 405.06
96
10
254
11
279.4
9.625 244.475 27 685.8 37 939.8 71 1803.4 2.8125 71.4375 2.8125 71.4375 1400 635.03
11/22/2010 2:44:10 PM
Table A12.9 Weights of thread protectors.
Book.indb 706
Weight of pin
Weight of box
Weight of pair
Connection type
lb
kg
lb
kg
lb
kg
3 1/2 API Regular 4 1/2 API Regular 5 1/2 API Regular 6 5/8 API Regular 7 5/8 API Regular 8 5/8 API Regular
7 17 21 28 35 55
3.18 7.71 9.53 12.70 15.88 24.95
8 14 18 23 31 46
3.63 6.35 8.16 10.43 14.06 20.87
15 31 39 51 66 101
6.80 14.06 17.69 23.13 29.94 45.81
11/22/2010 2:44:11 PM
Appendix 13
Details of shock absorbers
Clearance Diameter Pin Connection to Drive Head
Shoulder to Shoulder Length Rubber Element Box Connection to Drill Pipes
Figure A13.1 Atlas Copco Shock Absorber.
Table A13.1 Details of Atlas Copco shock absorbers.
Book.indb 707
Model designation
Max. Pull down load in kg
Clearance diameter in mm
Shoulder to shoulder length in mm
Assembly weight in kg
28 22 18 15
68038 38555 27215 15875
863.6 685.8 558.8 508
914.4 762 711.2 711.2
1044 578.5 379 295
11/22/2010 2:44:11 PM
708
Appendix 13
Clearance Diameter Pin Connection to Drive Head
Element Assembly Shoulder to Shoulder Length
Box Connection to Drill Pipes
Figure A13.2 MTI Shock Absorber. Table A13.2
Book.indb 708
Details of MTI shock absorbers.
Model designation
Max. pull down load in kg
Clearance diameter in mm
Shoulder to shoulder length in mm
Assembly weight in kg
130 100 75 50 40
45359–68038 36287–45359 31751–36287 22680–31751 15875–22680
660.4 609.6 457.2 457.2 457.2
927.1 863.6 812.8 838.2 762
953 850 644 508 363
11/22/2010 2:44:12 PM
Appendix 14
Properties of some rocks
Properties of rocks differ considerably from place to place. Further, there is no standard about the nomenclature of the rocks. For this reason, a range of values is often given in the following table rather than a specific value. It is also quite likely that the values for properties for a rock with a particular name as presented in different tables differ considerably. This can happen because the properties have been arrived at after testing samples obtained from different places. Therefore, the tables should not be amalgamated. When a very important decision is to be taken several rock samples from the worksite should be properly tested in a well-equipped laboratory and the values found by such testing should be used for taking the decision.
Book.indb 709
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710
Appendix 14
Table A14.1 Properties of certain rocks. Schmidt Cercher P wave hardness abrasivity velocity index index m/s
Rock type
Name of the rock
Dry density g/cm3
Igneous
Granite
2.53–2.62
1.02–2.87 54–69
4.5–5.3
Diorite Gabbro Rhyolite Andesite Basalt
2.80–3.00 2.72–3.00 2.40–2.60 2.50–2.80 2.21–2.77
0.10–0.50 1.00–3.57 0.40–4.00 0.20–8.00 67 0.22–22.1 61
4.2–5.0 3.7–4.6
Conglomerate Sandstone Shale Mudstone Dolomite Limestone Gneiss Schist
2.47–2.76 1.91–2.58 2.00–2.40 1.82–2.72 2.20–2.70 2.67–2.72 2.61–3.12 2.60–2.85
1.62–26.4 10–37 20.0–50.0 27 0.20–4.00 0.27–4.10 35–51 0.32–1.16 49 10.0–30.0 31
Phyllite Slate Marble Quartzite
2.18–3.30 2.71–2.78 2.51–2.86 2.61–2.67
1.84–3.64 0.65–0.81 0.40–0.65
Sedimentary
Metamorphic
Porosity %
2.7–3.8 2.0–3.5 1.5–3.8 1.5–4.2 0.6–1.8
1.0–2.5 3.5–5.3 2.2–4.5
2.3–4.2
S wave velocity m/s
4500–6500 3500– 3800 4500–6700 4500–7000 4500–6500 5000–7000 3660– 3700 1500–4600 2000–4600 5500 3500–6500 5000–7000 6100–6700 3460– 4000 3500–4500 5000–6000
4.3–5.9
Table A14.2 Properties of certain rocks.
Book.indb 710
Name of the rock
Porosity
Compressive strength MPa
Tensile strength MPa
Punch shear Strength MPa
Cohesion MPa
Angle of internal friction °
Young’s modulus GPa
Elastic
Density
Strength properties
Specific gravity
Physical properties
Quartz Granite Dolerite Sandstone Limestone (Grade I) Limestone (Grade II)
2.66 2.76 2.84 2.06 2.65 2.04
2.58 2.61 2.7 2.45 2.7 2.63
0.2 0.77 3.44 16.87 11.23 15.52
188.89 169.8 81.95 44.96 59.92 47.2
8.69 9 6.93 4.99 6.35 5.2
25.4 20.63 13.29 8.44 12.79 11.55
34.5 32 20 18 14 6
63 56 48 42 46 40
102 92 58 41.6 47.5 35
11/22/2010 2:44:12 PM
Appendix 14
Poisson’s ratio
Strain at failure MPa
Point load index
Fracture Mode I Toughness MPa
30–70
0.17
0.25
5–15
0.11–0.41
7–30 7–30 5–10 5–15 10–30
30–100 40–100 10–50 10–70 40–80
0.10–0.20 0.30 0.20–0.35 0.30 0.20–0.40 0.20 0.10–0.20 0.35
6–15
>0.41 >0.41
10–15 9–15
>0.41
10−11–10−9 10−12–10−11 10−13–10−10 10−14–10−12 10−11–10−8
30–230 20–170 5–100 10–100 20–120 30–250 100–250 70–150
3–10 4–25 2–10 5–30 6–15 6–25 7–20 4–10
10–90 15–50 5–30 5–70 30–70 20–70 30–80 5–60
0.10–0.15 0.14 0.10 0.15 0.15 0.30 0.24 0.15–0.25
10−14–10−12 10−14–10−11 10−14–10−13
5–150 50–180 50–200 150–300
6–20 7–20 7–20 5–20
10–85 20–90 30–70 50–90
0.26 0.20–0.30 0.35 0.15–0.30 0.40 0.17 0.20
Tensile Coefficient of UC strength strength permeability MPa MPa
Elastic modulus MPa
10−14–10−12
100–300
7–25
10−14–10−12 10−14–10−12 10−14–10−12 10−14–10−12 10−14–10−12
100–350 150–250 80–160 100–300 100–350
10−10–10−8 10−10–10−8
Book.indb 711
1–8
0.15 0.17
0.027–0.041 0.027–0.041
0.1–6
0.12
3–7 5–15 5–10
0.027–0.041 0.11–0.41 0.005–0.027
1–9 4–12 5–15
0.027–0.041 0.11–0.41 >0.41
Longitudinal wave velocity m/s
Shear wave velocity m/s
Shore hardness
Vickers hardness
Micro bit drilling hardness mm/min
Micro bit drilling abrasivity Wt. Loss x 10000
Cercher’s index
Quartz percent %
Protodyanokov index
Index
Poisson’s ratio
Dynamic
0.16 0.20
711
0.26 0.33 0.3 0.28 0.24 0.3
5225 4350 3270 2000 3200 3016
4059 2851 2430 850 1430 1280
82 76 46 41 35.27 26.3
710 630 330 285 240 180
115 100 65 85 37 22.5
0 0 0.38 0.85 0.76 1.4
6.8 6.1 5 5.6 4.8 3.4
100 35–40 0 40–50 15–20 10
20 20 10.4 3.4 8.24 6.3
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712
Appendix 14
Table A14.3 Typical densities of certain rocks. Rock name (type)
Density in kg/m3
Rock name (type)
Density in kg/m3
Andesite (I)
2.61
Dolomite (S)
2.30
Basalt (I)
2.99
Limestone (S)
2.11
Diabase (I)
2.91
Sandstone (S)
2.24
Diorite (I)
2.85
Shale (S)
2.10
Gabbro (I)
3.03
Av. Sedimentary Rocks
2.19
Granite (I)
2.64
Amphibolite (M)
2.96
Peridotite (I)
3.15
Gneiss (N)
2.80
Porphyry (I)
2.74
Granulite (M)
2.63
Pyroxenite (I)
3.17
Phyllite (M)
2.74
Av. Igneous Rocks
2.69
Quartzite (M)
2.60
Av. Acidic Igneous Rocks
2.61
Schist (M)
2.64
Av. Basic Igneous Rocks
2.79
Serpentine (M)
2.78
Clay (S)
1.70
Slate (M)
2.79
Coal (S)
1.35
Av. Metamorphic Rocks
2.76
Table A14.4 Compressive and tensile strengths of certain rocks. Compressive strength
Book.indb 712
Tensile strengt h
Rock name
lb/in2
MPa
lb/in2
MPa
Quartzite Quartzite Quartzite Argilite Diabase Basalt Basalt Basalt Gabbro Gabbro Granite Granite Granite Marble Limestone Limestone Dolomite Hornblende Schist
31650 22250 43700 31400 53300 9800 26500 40800 29600 25050 24350 22000 28950 18150 14200 17800 13800 29600
218.22 153.41 301.30 216.50 367.49 67.57 182.71 281.31 204.08 172.71 167.89 151.68 199.60 125.14 97.91 122.73 95.15 204.08
2510 2550 2950 2620 3550 730 1990 4020 2150 1810 1780 1300 1850 1010 820 910 600 1080
17.31 17.58 20.34 18.06 24.48 5.03 13.72 27.72 14.82 12.48 12.27 8.96 12.76 6.96 5.65 6.27 4.14 7.45
11/22/2010 2:44:13 PM
Table A14.5 Densities of certain minerals. Mineral name
Chemical formula
Density in kg/m3
Anhydrite Apatite Arsenopyrite Asbestos Augite Barite Bauxite Beryl Biotite Calcite Cassiterite Celestite Chalcocite Chalcopyrite Chlorite Chromatite Chromite Cinnabar Corundum Cuprite Diopside Dioptase Dolomite Emerald Fayalite Fluorite Forsterite Franklinite Galena Gibbsite Graphite Gypsum Halite Hematite Hornblende Hypersthene Illite Ilmenite Jacobsite Limonite Magnesite Magnetite Montmorillonite Olivine Opal Orthoclase Orthoferrosilite
CaSO4 Ca5(PO4)3(F,Cl,OH ) FeAsS Mg3Si2O5(OH)4 (Ca,Na)(Mg,Fe,Al,Ti)(Si,Al)2O6 BaSO4 AlO(OH) Be3 Al2Si6O18 K(Mg,Fe++)3[AlSi3O10(OH,F)2 CaCO3 SnO2 SrSO4 Cu2S CuFeS2 Na0.5(Al,Mg)6(Si,Al)8O18(OH)12•5(H2O) CaCrO4 Fe++Cr2O4 HgS Al2O3 Cu2O CaMgSi2O6 CuSiO2(OH)2 CaMg(CO3)2 Ni3(CO3)(OH)4•4(H2O) Fe++2SiO4 CaF2 Mg2SiO4 (Zn,Mn++,Fe++)(Fe+++,Mn+++)2O4 PbS Al(OH)3 C CaSO4.2H2O NaCl Fe2O3 Ca2[Mg4(Al,Fe+++)]Si7 AlO22(OH)2 (Mg,Fe++)2Si2O6 (K,H3O)(Al,Mg,Fe)2(Si,Al)4O10[(OH)2,(H2O)] Fe++TiO3 (Mn++,Fe++,Mg)(Fe+++,Mn+++)2O4 Fe+++O(OH) MgCO3 Fe++Fe+++2O4 (Na,Ca)0.3(Al,Mg)2Si4O10(OH)2•n(H2O) (Mg,Fe)2SiO4 SiO2•n(H2O) KAlSi3O8 (Fe++,Mg)2Si2O6
2.96–2.98 3.16–3.22 6.07 2.53 3.2–3.6 4.48 3–3.07 2.63–2.9 2.8–3.4 2.71 6.8–7 3.9–4 5.5–5.8 4.1–4.3 2.42 3.142 4.5–5.09 8.1 4–4.1 6.11 3.25–3.55 3.28–3.35 2.8–2.9 2.6 4.39 3.01–3.25 3.21–3.33 5.07–5.22 7.2–7.6 2.3–2.4 2.09–2.33 2.3 2.17 5.3 3–3.47 3.2–3.9 2.75 4.72 4.75 3.3–4.3 3.21 5.1–5.2 2–2.7 3.27–3.37 1.9–2.3 2.56 3.88–4.02
(Continued)
Book.indb 713
11/22/2010 2:44:14 PM
Table A14.5 (Continued)
Book.indb 714
Mineral name
Chemical formula
Density in kg/m3
Orthopyroxenes Pentlandite Plagioclase Pyrite Pyrrhotite Quartz Rutile Serpentinite Siderite Sphalerite Sylvite Talc Titanite Tourmaline Troilite Wolframite Zircon
(Fe,Mg)2SiO3 (Fe,Ni)9S8 (Na,Ca)(Si,Al)4O8 FeS2 Fe1-XS SiO2 TiO2 Mg3Si2O5(OH)4 Fe++CO3 (Zn,Fe)S KCl Mg3Si4O10(OH)2 CaTiSiO5 NaFe+++3 Al6(BO3)3Si6O21F FeS (Fe,Mn)WO4 ZrSiO4
3.59 4.6–5 2.61–2.76 5–5.02 4.58–4.65 2.6–2.65 4.25 2.53–2.65 3.96 3.9–4.2 1.99 2.7–2.8 3.45–3.6 3.31 4.58–4.65 7.75 4–4.8
11/22/2010 2:44:14 PM
Book.indb 715
Table A14.6 Properties of certain rocks. Compressive strength Rock name
Specific gravity
Amphibolite Andesite Argillite Basalt Chert, dolomite Conglomerate Diabase Diorite Diorite, augite Dolomite Gabbro Granite Granite, aplitic Granite, gneissic Granite, pre-Cambrian Granodiorite Greenstone Hematite ore Hornfels Limestone Limestone, chalky
3.07 2.81 2.81 2.94 2.67 2.67 2.94 3.01 2.74 2.60 3.00 2.66 2.65 2.66 2.80 2.74 3.02 5.07 3.19 2.68 1.89
lb/in2
MPa
61335 56535 19720 44950 29290 23925 46545 39730 48285 18995 44805 37700 51185 30305
423 183 136 310 202 165 321 274 333 131 309 260 353 209
36540 39005 88015 77285 22330 4205
252 269 607 533 154 29
Poisson’s ratio
0.14
0.29 0.25 0.18 0.33 0.2 0.26 0.02 0.27 0.24
0.28 0.02
Engineering classification of intact rock based on
Modulus of rigidity
Young’s modulus
lb/in2
MPa
lb/in2
GPa
UCS
Compressibility
6641000 3944000
45800 27200
4596500 3436500 4698000 5408500 6119000 4886500 2900000 6394500 3422000 4756000 1299200 7583500 4060000 6104500
31700 23700 32400 37300 42200 33700 20000 44100 23600 32800 8960 52300 28000 42100 40900 26500 5380
104.0 64.6 84.1 77.9 56.2 77.9 95.8 107.0 84.1 47.6 119.0 59.2 80.6 18.6 82.1 68.6 105.0 200.0 95.8 68.1 11.1
A,Very High B, High B, High A,Very High B, High B, High A,Very High A,Very High A,Very High B, High A,Very High A,Very High A,Very High B, High
5930500 3842500 780100
15080000 9367000 12194500 11295500 8149000 11295500 13891000 15515000 12194500 6902000 17255000 8584000 11687000 2697000 11904500 9947000 15225000 29,000,000 13891000 9874500 1609500
1 – Low 2 – Medium 1 – Low 2 – Medium 2 – Medium 2 – Medium 1 – Low 1 – Low 1 – Low 2 – Medium 1 – Low 2 – Medium 2 – Medium 3 – High 2 – Medium 2 – Medium 1 – Low 1 – Low 1 – Low 2 – Medium 3 – High
A,Very High A,Very High A,Very High A,Very High B, High D, Low
11/22/2010 2:44:14 PM
Appendix 15
Bibliography
Table A15.1 List of books, presentations, theses and reports covering drilling and blasting. Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year Type
1 Advanced Drilling Techniques
Maurer W. C.
698
Petroleum Publishing
1980 Book
2 Air and Gas Drilling
Lyons W. C.
656
McGraw Hill Companies Inc.
2000 Book
3 Air Drilling Handbook
Van Omer H.
167
International Society for Explosives Engineers
1987 Book
4 Applied Geomechanics in Drilling
Vaitenko V. S.
266
A. A. Balkema
1995 Book
5 Atlas Copco Manual
Editorial Committee
656
Atlas Copco
1982 Book
6 Bench Drilling Techniquesand Equipment Selection Manual
Ed – Drake R.
92
Ingersoll Rand Company
1990 Book
7 Blaster’s Handbook
Hopler R. B.
744
International Society for Explosives Engineers
1998 Book
8 Blasthole Bit Handbook
Anonymous
28
Hughes Tool Company
1975 Book
9 Blasthole Drilling in Open Pit Mining
Ed. – Ulf Linder, Diana Norwood et al.
208
Atlas Copco Drilling Solutions LLC
2009 Book
10 Blasthole Technology
Anonymous
58
Reed Tool Company
1972 Book
11 Blasting in Opencast Mines and Quarries
Anonymous
24
Indian Explosives Ltd.
1972 Book
12 Blasting Principles for Open Pit Mining (Vol. 1 & 2)
Hustrulid W.
1038
A. A. Balkema
1999 Book
13 Bucyrus Erie Answers Your Questions About Blasthole Drills
Nelmark J.
8
Bucyrus Erie Company
1981 Book
(Continued)
Book.indb 717
11/22/2010 2:44:15 PM
Table A15.1 (Continued) Name of the book/ Sr. article/product no. literature
Author/editor
Pages
Publisher
14 Coal Mining in India
Mathur S. P.
474
M. S. Enterprise
1999 Book
15 Compressed Air and Gas Data
Gibbs C. W.
800
Ingersoll Rand Company
1976 Book
16 Compressed Air and Gas Handbook
Rollins J. P.
903
Prentice Hall
1989 Book
17 Construction Planning, Equipment and Methods
Schexngyder C. J., Peurifoy W. B., Ledbetter W. B.
688
McGraw Hill Companies Inc.
2002 Book
18 Damage and Fracture in Heterogeneous Materials
Mishnaevsky L. L. Jr.,
230
A. A. Balkema
1998 Book
19 Design and Application Data for Varel Drill Bits
Anonymous
24
Varel Manufacturing Co.
1978 Book
20 Diamond Drill Handbook
Cumming
547
J. K. Smit & Sons Diamond Products Ltd.
1975 Book
21 Diamond Drilling Handbook
Heinz W. F.
538
A. A. Balkema
1992 Book
22 Dictionary of Drilling and Borehole
Moureau M.
436
French and European Publications, Inc.
1990 Book
23 Drillability – Drilling Rate Index Catlog
Bruland A., Eriksen S., ohannessen O., Sandberg B.
177
Dept. of Geology, Norwegian Inst. of Technology
1990 Book
24 Drilling and Excavation Technologies for the Future
Several Authors
176
National Academy Press
1994 Book
25 Drilling Practices Manual
Moore P. L.
448
Petroleum Publishing Co.
1974 Book
26 Drilling Technology
Vozniak J. P.
357
ASME
1995 Book
27 Engineering Geology Field Manual
Anonymous
491
U.S. Department of Interior, Bureau of Reclamation
2001 Book
28 Excavation Handbook
Church H. K.
1024
McGraw Hill Companies Inc.
1980 Book
29 Explosives and Rock Blasting
Morhard R. C.
662
Atlas Powder Company
1987 Book
30 Face Drilling
Ed – Mike Smith
160
Atlas Copco Rock Drills AS
2004 Book
31 Geology
Smith A. J.
128
Hamlyn
1974 Book
32 Handbook on Surface Drilling and Blasting
Editorial Committee
320
Tamrock
1984 Book
Publication Year Type
(Continued)
Book.indb 718
11/22/2010 2:44:15 PM
Table A15.1 (Continued) Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year Type
33 Hard Rock Miner’s Handbook
de la Vergne J.
330
McIntosh Engineering Limited.
2003 Book
34 History of Oilwell Drilling
Brantly J. E.
1525
Gulf Publishing
1971 Book
35 Improving Compressed Air System Performance
Several Authors
128
U.S. Department of Energy
2003 Book
36 Introductory Mining Engineering
Hartman H. L., Mutamansky J. M.
584
John Wiley & Sons
2002 Book
37 Mine Planning and Equipment Selection 1998
ED – Singhal P. K.
830
A. A. Balkema
1998 Book
38 Mining Engineers Handbook
Ed – Hartman H.
2394
Society for Mining Metallurgy & Exploration
2002 Book
39 Modern Techniques of Rock Blasting
Langfors U., Kihlstrom B.
406
J. Wiley and Sons
1973 Book
40 Novel Drilling Techniques
Maurer W. C.
114
Pergamon Press
1969 Book
41 Open Cut Blasthole Drilling
Anonymous
218
The Australian Drilling Industry Training Committee Limited
1987 Book
42 Open Pit Mine Planning and Design Vol. 1&2
Hustrulid W., Kuchta M.
864
A. A. Balkema
1995 Book
43 Petroleum Engineering Drilling and Well Completion
Gatlin C.
341
Prentice Hall
1960 Book
44 Physical Properties of Rocks and Minerals
Touloukian Y. S., Ho C.Y.
548
McGraw Hill Companies Inc.
1981 Book
45 Pneumatic Handbook
Barber A.
659
Elsevier Press
1997 Book
46 Pneumatic Handbook
Warring R. H
433
Guld Publishing Co.
1982 Book
47 Practical Handbook of Physical Properties of Rocks and Minerals
Carmichael R. S.
760
CRC Press
1990 Book
48 Preventive Maintenance for Rock Drills
Anonymous
107
Compressed Air and Gas Inst.
1963 Book
49 Principles of Rock Drilling
Clark G. B.
91
Colorado School of Mines
1979 Book
50 Principles of Rock Drilling
Rao Karanam U. M., 272 Misra B.
A. A. Balkema
1998 Book
(Continued)
Book.indb 719
11/22/2010 2:44:15 PM
Table A15.1 (Continued) Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year Type
51 Principles of Rock Fragmentation
Clark G. B.
610
John Wiley & Sons
1987 Book
52 Reciprocating and Rotary Compressors
Chlumsky V.
534
E & FN Spon Ltd.
1965 Book
53 Rock Bit Design, Selection and Evaluation
Bentson H. G
36
Smithtool
1966 Book
54 Rock Blasting Terms and Symbols
Ed – Rustan Agne
204
A. A. Balkema
1998 Book
55 Rock Drilling Data
Anonymous
224
Padley & Venables Ltd.
1974 Book
56 Rock Drilling Manual – Drill Steel Applications
Anonymous
84
Sandvik – Atlas Copco
1980 Book
57 Rock Drilling Manual – Theory and Technique
Anonymous
44
Sandvik – Atlas Copco
1977 Book
58 Rock Excavation
Dessureault, Sean
217
Dept. of Mining and Geological Engineering, Univ. of Arizona
2003 Book
59 Rock Excavation Handbook for Civil Engineering
Ed – Matti Heinio
364
Sandvik Tamrock
1999 Book
60 Rock Fracture Mechanics; Principles, Design, and Applications
Whittaker B. N., Singh R. N., Sun G.
570
Elsevier Press
1992 Book
61 Rotary Drilling Handbook
Brantly J. E.
825
Palmer Publications
1961 Book
62 Sintered Metal Carbides
Romonova N., Chekulaev P., et al.
332
Mir Publishers
1972 Book
63 SME Mining Engineering Handbook Vol. 1&2
Ed – Cummins A. B., Given I. A.
App 1600
SME of AIME
1973 Book
64 Surface Blast Design
Konya C. J., Walters E. J.
303
Prentice Hall
1990 Book
65 Surface Blaster’s Certification Study Guide
Anonymous
94
Commonwealth of Virginia, Dept. of Mines, Minerals and Energy, Division of Mineral Mining
2009 Book
66 Surface Drilling
Anonymous
114
Atlas Copco
1998 Book
67 Surface Mining
Pfleider E. P.
1061
SME of AIME
1968 Book
(Continued)
Book.indb 720
11/22/2010 2:44:15 PM
Table A15.1 (Continued) Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year Type
68 Surface Mining
Kennedy B. A
1175
SME of AIME
1990 Book
69 Surface Mining Equipment
Martin J. W., Bennett T. P., Martin T. J.
455
Martin Consultants Inc.
1982 Book
70 Swedish Blasting Techniques
Anonymous
328
Sprang Tekniska Institutet AB
1974 Book
71 Systematic Drilling and Blasting for Surface Excavations
Anonymous
145
Department of the Army, U.S. Army Corps of Engineers
1972 Book
72 The Drilling of Rocks
McGregor K.
306
CR Books Ltd.
1967 Book
73 The Physics of Rocks
Rzhevsky V., Novik G.
320
Mir Publishers
1971 Book
74 Excavation of Hard Ground Using Drill and Blast Techniques
Graeme Jones
13
AUCTA Tunnel Course
2003 Notes
75 Blaster’s Training Modules
Anonymous
402 Slides
Office of Surface Mining Reclamation and Enforcement, U.S. Department of Interior
2008 Present
76 Blasthole Drilling
Anonymous
27 Slides
Training Module – Office of Surface Mining U.S. Department of Interior, Denver
Present
77 Blasting Rock Practical Exercise
Anonymous
57 Slides
Web Presentation – BlastProd1.pt
Present
78 Drilling of Rock and Earth
Assakkaf I.
72
Department of Civil and Environmental Engineering, University of Maryland
2003 Present
79 Drilling of Rock and Earth
Anonymous
67
McGraw Hill Companies Inc.
2007 Present
80 Excavation Methods
Eberhardt E.
52 Slides
UBC Geological Engineering
2007 Present
81 Explosive and Blasting Agents
Annette Slyman
35
Institute of Mining Engineering, University of Aachen
Present
82 Explosives
Anonymous
42 Slides
Training Module – Office of Surface Mining US Department of Interior, Denver
Present
(Continued)
Book.indb 721
11/22/2010 2:44:16 PM
Table A15.1 (Continued)
Book.indb 722
Name of the book/ Sr. article/product no. literature
Author/editor
Pages Publisher
Year
83 LaserAce Burden Finder
Jagdeepak Sharma
12 Ultra EnviroSlides Systems (P) Ltd.
2005 Present
84 Machine Guidance System – Moving More With Less
Annette Slyman
45 CAT Global Mining Slides
2007 Present
85 Presentation on MREL High Speed Video and Borehole Survey Instruments
Anonymous
20 Web Presentation – Slides Presentation High Speed Video.ppt
Present
86 SimQuarry
Anonymous
27 Sandvik Mining and Slides Construction
Present
87 VOD (Velocity of Detonation) Instruments
Jagdeepak Sharma
22 Ultra EnviroSlides Systems (P) Ltd.
2005 Present
88 Design of Surface Blasts – A Computational Approach
Ashutosh Mishra
77
Dept of Mining Engineering – National Institute of Technology Rourkela
2009 Thesis G
89 Blasting Design Using Fracture Toughness and Image Analysis of the Bench Face and Muckpile
Kwangmin K.
124
Virginia Polytechnic Institute
2006 Thesis M
90 Comparison of Rotary and In-Hole Motor Techniques for Drilling Horizontal Boreholes in Coal
Kravits S. J., Sainato A., Finfinger G. L.
36
1985 Report U.S. Bureau of Mines Report of Investigation No. 8933, U.S. Department of Interior
91 Noise,Vibration and Airblast Management Plan (NVAM) Eagle Rock Quarry Project
Anonymous
7
For Polaris Minerals Corporation by Explosives and Rockwork Technologies Ltd.
92 Rotary Drilling Holes in Coalbeds for Degasification
Cervik J., Fields H. H., Aul G. N.
25
1975 Report U.S. Bureau of Mines Report of Investigation No. 8097, U.S. Department of Interior
92 Rotary Drilling Techniques Used in the Beckley Coalbed
Goodman T. W.
16
1989 Report U.S. Bureau of Mines Report of Investigation No. 9238, U.S. Department of Interior
94 Vibration Assessment – Eagle Rock Quarry Project
Anonymous
24
For Polaris Minerals Corporation by Explosives and Rockwork Technologies Ltd.
Publication Type
2002 Report
2002 Report
11/22/2010 2:44:16 PM
Table A15.2 List of articles covering drilling and blasting.
Sr. no.
Name of the book/ article/product literature
Publication Author/editor
Pages
Publisher
Year
Type
92
Petroleum Transactions AIME Vol. 216
1959 Article
1
A Laboratory Study of Rock Breakage by Rotary Drilling
Somerton W. H.
2
A New Approach to the Design of Drilling Tools
Mishniavsky L. Jr., 6 Schmauder S.
International Journal Rock Mechanics Mineral Science & Geomechanics
1996 Article
3
A Preliminary Theory of Static Penetration by a Rigid Wedge Into Brittle Material
Paul B., Sikerskie D. L.
372
Transactions SME/AIME Vol. 232
1965 Article
4
A Quiet Series of Innovations Helps “KEEP IT TURNING TO THE RIGHT”
Zaburunov S.
11
World Mining Equipment
5
A Simple Criterion of Machinability of Hard Rocks
Singh S. P.
257–266
International Journal, Mining Geology Engg.
1989 Article
6
A Theoretical Description of Rotary Drilling for Idealized Down-Hole Bit/ Rock Conditions
Paul F., Gnirk J. B., Cheatham J. R.
443
Journal Society Petroleum Engineers
1969 Article
7
A Two Component Model of Blast Fragmentation
Djordjevic N.
13-Sep
Proceedings Australian Institute Mining and Metallurgy
1999 Article
8
Advantages and Disadvantages of Down the Hole Hammer Blasthole Dirlling vs Rotary Drilling in Large Scale Open Pit Surface Mining
Raitt G., Lyon R.
365–376
International Society of Explosives Engineers 1998 G
1998 Article
9
An Empirical Approach for Relating Drilling Parameters
Cunningham R. A. 987
Journal Petroleum Technology
1978 Article
10
An Empirical Equation for Drill Performance Prediction
Rabia H., Brook N.
103–111
21st Symposium on Rock Mechanics
1980 Article
11
An Explosion in Mining
Eloranta J.
12
Eloranta & Associates Inc.
2000 Article
Article
(Continued)
Book.indb 723
11/22/2010 2:44:17 PM
Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
11 An Introduction to Drill Vibration
Schivley G. P. Jr.,
647–655
International Society of Explosives Engineers 1997G
1999 Article
12 An Introduction to Drilling Vibrations
Schivley G. P. Jr.,
647–655
International Society of Explosives Engineers 1997G
1997 Article
13 Assessment of Abrasive Wear Resistance Potential in Rock Excavating Machinery
Atkinson T., Cassapi V. B., Singh R. N.
151–163
International Journal of Rock Mechanics and Mineral Science
1986 Article
14 Asymmetric Blasting: A Rock Mass Dependent Blast Design Method
Segui J. B.
6
Proceedings of EXPLO 2001, Hunter Valley, NSW
2001 Article
15 Bench Blast Modeling Using Numerical Simulation and Mine Planning Software
Firth I. R., Taylor D. L.
4
Mackay School of Mines, University of Nevada, Reno
16 Bench Blast Optimization
Budin B., Newton G., Thompson J.
21
Colorado School of Mines
2003 Article
7
Mining Engineering
2000 Article
17 Bit Geometry Effects on Failure Characteristics of Rock
Type
Article
18 Bit Penetration Into Rock – A Finite Element Study
Wang J. K.
16-Nov
International Journal of Rock Mechanics, Mining Science., & Geomechanics
1976 Article
19 Blasthole Bits – Virtues of Simplicity
Anonymous
3
World Mining Equipment
1994 Article
20 Blasthole Drill: Better by Performance
Carter R. A.
4
Engineering & Mining Journal
1993 Article
21 Blasthole Drilling, Rotary Drilling and the Four Kingdoms
Praillet D.
4
World Mining Equipment
1998 Article
22 Blasthole Drills
Anonymous
Mining Journal
2000 Article
23 Blasthole Drills: Better by Design
Carter R. A.
5
Engineering and Mining Journal
1993 Article
24 Blasting Cap
Anonymous
2
NationMaster.com
Article
(Continued)
Book.indb 724
11/22/2010 2:44:17 PM
Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Year
Type
Author/editor
Pages
Publisher
25 Blasting in Surface Mines
Anonymous
20
Web Article – H_BLASTING IN SURFACE EXCAVATIONS.pdf
26 Blasting Review & Maintenance Program
Konya Calvin J.
3
Engineering and Mining Journal
1996 Article
27 Blasting Vibrations – Mining Engineer’s Toolbox
Anonymous
2
Web Article
2005 Article
28 Characterization of the Pre and Post Blast Environments
Eloranta J.
10
Proceedings Twenty Ninth Conference of Explosives and Blasting Techniques, ISEE
2003 Article
29 Choice of Optimal Sizes of Rock Destructing Elements and Their Arrangements on the Drilling Bits
Eigheles R. M., 136–150 Strekakive R. M., Mustafina N. N.
Rock Fragmentation (Ed. Eigheles R. M.), VNIIBT Moscow
1975 Article
Article
30 Choosing the Proper Dick R. A., Borehole Size for Olson J. J. Bench Blasting
41–45
Mining Enginnering
1972 Article
31 Concept of Blastability – An Update
Kaushik Dey, Phalguni Sen
8
Indian Mining and Engineering Journal
2003 Article
32 Control of Fragmentation by Blasting
Bozic B.
9
Rudarsko-geološko-naftni zbornik
1998 Article
33 Death of a Dirty Raitt G. Driller: The DM Story
3
World Mining Equipment
1994 Article
Canadian Institute of Mining Bulletin Vol. 75
1982 Article
34 Development Leighton J., Brawner C., of a Correlation Between Rotary Drill Stewart D. Performance and Controlled Blasting Powder Factors 35 Differential GPS Applications at the Morenci Copper Mine
Shchields S., Flinn J., Obregon A.
4
GPS World
2000 Article
36 Down to earth: Mike Woof reports on drilling innovations.
Woof M.
4
World Mining Equipment
2002 Article
37 Drill Accuracy
Harris J. E.
35
International Society of Explosives Engineers 1999G
1999 Article
(Continued)
Book.indb 725
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Table A15.2 Continued Sr. no.
Name of the book/ article/product literature
Publication Author/editor
Pages
Publisher
Year
Type
11
World Mining Equipment Conference
1999 Article
38
Drill Monitoring Developments and the Utilization of the Drill Data
Vynne J.
39
Drill Monitoring Investigations in a Western Canadian Surface Mine
Scoble M., Peck J., Kennedy D.
Proceedings 2nd Conference Large Open Pit Mining
1989 Article
40
Drillability and Physical Properties of Rocks
Singh D. P.
Proceedings Rock Mechanics Symposium Institute of Engineers, Australia
1969 Article
41
Drillability in Hard Rock Drill and Blast Tunneling
Thuro K., Spaum G.
103–109
Falsbau 14
1996 Article
42
Drillability Predictions – Geological Influences in Hard Rock Drill and Blast Tunneling
Thuro K.
426–437
Geological Rundsche 86
1997 Article
43
Drillability Predictions in Hard Rock Tunneling.
Blindheim O.T.
284–289
Tunneling – Institute of Mining and Metallurgy
1979 Article
44
Drillability Studies – Diamond Drilling
Paone J., Bruce W. E.
U.S. Bureau of Mines Report of Investigation No. 6324, U.S. Department of Interior
1963 Article
45
Drillability Studies – Laboratory Percussive Drilling.
Paone J., Madson D., Bruce W.E:
U.S. Bureau of Mines Report of Investigation No. 7300, U.S. Department of Interior
1969 Article
46
Drillability Studies – Percussive Drilling in the Field
Schmidt R. L.
U.S. Bureau of Mines Report of Investigation No. 7684, U.S. Department of Interior
1974 Article
47
Drilling – Mining Engineer’s Toolbox
Anonymous
Web Article
2005 Article
48
Drilling and Blasting in Open Pits and Quarries
Crosby W. A.
Australian Mineral Foundation, Presentation No152/81
1981 Article
49
Drilling and Blasting Pattern Selection for Fragmentation Optimization in Raycon Quarry Ore, Ondo State
Saliu M. A., Akande J. M.
Journal of Engineering and Applied Sciences
2007 Article
50
2
6
(Continued)
Book.indb 726
11/22/2010 2:44:18 PM
Table A15.2 Continued Sr. no.
Name of the book/ article/product literature
Publication Author/editor
Pages
Publisher
Year
Type
51
Drilling Machines, Surface
Morrell R., Unger H. F.
Sec. 11.4
SME Mining Engineers Handbook, SME/AIME
1973 Article
52
Dust Control in Surface Mines
Brown J. W.
57
International Society of Explosives Engineers 1999G
1999 Article
53
Dynamic Stability of an Impact System Connected With Rock Drilling
Fu C. C.
743–749
Journal Applied Mechanics
1969 Article
55
Earth
Arnett W.
5
Web Article
2003 Article
56
Effect of Charge Weight of Vibration Levels from Quarry Blasting
Devine J. F. et al.
U.S. Bureau of Mines Report of Investigation No. 6774, U.S. Department of Interior
1966 Article
57
Energy Balance in Rock Drilling
Simon R.
Journal Society of Petroleum Engr.
1963 Article
58
Engineering Classification and Index Properties for Intact Rocks
Deere D. U., Miller R. P.
Report No. AFWL-TR65–116 Air Force Weapons Lab.
1966 Article
59
Estimating Drilling and Blasting Cost and Analysis and Prediction Model
Borquez G.V.
7
Engineering and Mining Journal
1981 Article
60
Experimental and Theoretical Analysis of Roller Cone Bit Vibrations
Naganawa S., Tanaka S., Sugaya N.
6
Drilling Symposium
1995 Article
61
Explosive Advances
Anonymous
6
World Coal
2002 Article
62
Explosive Material
Anonymous
7
NationMaster.com
63
Explosive Materials
Anonymous
Wikipedia Article
2009 Article
64
Explosives
Anonymous
Web Article – Explosives’04.doc
2004 Article
65
Explosives and Blasting
Sinkala T.
Web Article – Module 6 Explosives and Blasting. pdf
Article
66
Factors in Rotary Drill Evaluation
Frank M. C.
20–33
Mining Congress Journal
1962 Article
67
Flame Jet Gains Ground as a Blasthole Drill
Browning J. A.
80–81
Engineering and Mining Journal,Vol. 170
1969 Article
68
Fluid Drag on a Tool
Rao Bharath
8
www.Ctes.com
1999 Article
69
Focus on Rotary Drilling Rigs
Schivley G. P. Jr.,
Compressed Air Magazine
1987 Article
54
298
Article
(Continued)
Book.indb 727
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Author/editor
Pages
Publisher
Year
70
Geologic Data for Blasting
Eloranta J.
11
Proceedings of the twenty-second conference of Explosives and Blasting Technique, ISEE
1996 Article
71
Going Straight – Straight hole systems boost drilling effectiveness
Woof M.
p. 8
World Mining Equipment
2002 Article
72
GPS Based Machine Guidance in the Mining Industry
Seymour
p. 15
12th International Symposium On Mine Planning And Equipment Selection
2003 Article
73
GPS Guidance System and Reduction of Open Pit Mining Costs and Revenue Loss
Jarosz A. P., Finlayson R.
p. 13
[email protected] 2004 Article
74
GPS Positioning and Equipment Monitoring technology for Blasting Operations
Hendricks C.
14
Proceedings of Minnblast 99 Surface Blasting Conference, Minnesota
1999 Article
75
Ground and Air Vibrations from Blasting
Siskind D. E.
Sec. 11.8
SME Mining Engineering Handbook, SME/AIME
1973 Article
76
Hard Iron Ore Yields Knoerr A. W. to Fusion Piercing
66–69
Engineering and Mining Journal Vol. 147
1946 Article
Proceedings Atlas Copco Bench Drilling Days, Symposium Stockholm Mine and Quarry Engineer
1975 Article
77
Hard Rock Blasting Developments and Possibilities 78 Hard Rock Drilling Techniques 79 Important Considerations When Selecting Rotary Blasthole Drills for High Altitude Use 80 Influence of Joint Direction and Position of Explosive Charge on Fragtmentation
Publication
Fairhurst C.
Fairhurst C., 157 Lacabanne W. D. Lyall Workman J. p. 16
Hafsaoui A., Talhi K.i
8
Calder & Workman, Inc.
The Arabian Journal of Science and Engineering
Type
1957 Article Article
2009 Article
(Continued)
Book.indb 728
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
Type
81
Introducing ‘Destruction Work’ as a New Rock Property of Toughness Referring to Drillability in Conventional Drill and Blast Tunneling
Thuro K., Spaum G.
707–713
Eurock 96, Prediction and Performance in R. Mech. and R. Engg. Vol. 2
1996 Article
82
Jet Piercing
Calaman J. J., Rolseth H. C.
325–337
Surface Mining (Ed. Pfleider E. P.), SME/AIME
1968 Article
83
Keep on Running: Rotary Maintenance
Woof M.
p. 6
World Mining Equipment
2003 Article
84
Large Diameter Holes in Overburden Blasting Large Hole Rotary Drill Performance
Piper W. C.
33–35
Mining Congress Journal
1963 Article
Lyall Workman J., 334–347 Calder P.
1996 Article
86
Lubrication of Rock Bits
MacPhail J., Gardner H.
87
Made in Milwaukee
Hammel K.
88
Made in Milwaukee
Woof M.
p. 4
89
Maximing Rotary Blast Hole Drills
Brown J. W.
p. 461
Paul B., Fu C. C.
301–307
International Society of Explosives Engineers 1996G International Society of Explosives Engineers 1996G World Mining Equipment World Mining Equipment International Society of Explosives Engineers 1998G Transactions. SME/AIME Vol. 232
Dobrilovic M., Ester Z., Jankovic B.
7
Rudarsko-geološkonaftni zbornik
2005 Article
Gstalder S., Raynal J.
991
Journal Petroleum Technology
1966 Article
85
90
Maximization of Footage Drilled Before Drill Steel Fails in Fatigue 91 Measurement in Blast Hole Stem and Influence of Stemming Material on Blasting Quality 92 Measurement of Some Mechanical Properties of Rock and Their Relationship to Rock Drillability
140–149
1996 Article
1999 Article 1999 Article 2000 Article
1965 Article
(Continued)
Book.indb 729
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
Type
Nauka, Novosibirsk
1985 Article
93 Mechanical and Physical Principles of Design of Rock Breaking Mining Tools
Artsimovich G.V.
94 Mechanical Properties and Drillability of Rocks.
Protodyakonov M. M.
103–118
Proceedings. 5th Symposium on Rock Mechanics
1962 Article
95 Modern Drilling and Blasting Technology at the Minntac Mine
Thompson D.
18
Proceedings of Minnblast 99 Surface Blasting Conference, Minnesota
1999 Article
96 Modern Methods of Rock Fragmentation in Hole Drilling
Suluashin S. S.
Nedra, Moscow
1964 Article
97 Monitoring While Drilling Production Blastholes: Applications in Surface Mining
Peck J., Scoble M. J., Hendricks C.
International Society of Explosives Engineers 1988G
1988 Article
98 New Blasting Technique to Eliminate Subgrade Drilling, Improve Fragmentation, Reduce Explosive Consumption and Lower Ground Vibrations
Chiapetta F. dd
6
Journal of Explosive Engineering
2004 Article
99 Normalization of Specific Energy
Mellar M.
661–663
International Journal of Rock Mechanics and Mineral Science
1965 Article
100 On Drillability of Rocks by Percussive Drilling
Selmar-Olsen R., Blindheim O. T.
Proceedings 2nd Conference. on Rock Mechanics ISRM
1970 Article
101 On the Penetration of Rock by Three Dimensional Indentors
Miller M. H., Sikerskie D. L.
375–398
International Journal of Rock Mechanics and Mineral Science
1968 Article
102 Open Pit Drilling
Bauer A.
175–180
28th Annual Mining Symp.
1967 Article
103 Open Pit Drilling and Blasting
Bauer A.
115–121
Journal South African Institute of Mining and Metallurgy
1971 Article
(Continued)
Book.indb 730
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
Type
104 Optimization of Shape of Drilling Tools
Mishnaevsky L.
1223– 1228
Proceedings. International. Congress Rock Mech.Vol. 2
1999 Article
105 Optimizing Blasthole Spacing
Konya K.
p. 3
World Mining Equipment
106 Optimizing Rotary Drill Performance
Schivley G. P. Jr.,
23-May
International Society of Explosives Engineers 1995 v2 G
107 Optimum Bit Selection and Operation for The Rotary Blasthole Drilling Through Horizontal Drilling Rig (HDR) – A Case Study at KBI Murgul Copper Mine
Ergin H., Kuzu C., Balci C., Tunçdemir H. & Bilgin N.
11
Istanbul Technical University, Istanbul
108 Perfect Cleaning Theory of Rotary Drilling
Maurer W. C.
1270
Journal of Petroleum Technology
1960 Article
109 Performance Characteristics of a Small Scale Rotary Boring Machine Instrumented with Large Drag Picks
Howarth D. F., Adamson W. R.
191–202
International Journal of Rock Mechanics and Mineral Science
1988 Article
110 Physical Properties and Drillability of Mine Rocks
Shepherd R.
468–470, Colliery Engineering Dec. 50, Jan. 51, Feb. 51, 28–34, Mar. 51 51–66, 121–126
1950 Article
111 Predicting Rotary Drill Performance
Schivley G. P. Jr.,
181–189
International Society of Explosives Engineers 1994G
1994 Article
112 Priming of Explosives
Anonymous
12
Web Article – G_PRIMING OF EXPLOSIVES.pdf
113 Principles of Drilling
Hartman H. L.
Sec 6.1
Surface Mining (Ed. Pfleider E. P.), SME/AIME
1968 Article
114 Principles of Drilling
Tandanand S.
Sec.11.3
SME Mining Engg. Handbook, SME/AIME
1973 Article
115 ProVision High Precision GPS Drill System
Anonymous
20
Komatsu Technology Company
2002 Article
Article 1995 Article
Article
Article
(Continued)
Book.indb 731
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature 116 Quantitative Assessment of Rock Texture and Correlation with Drillability and Strength Properties
Publication Author/editor
Pages
Publisher
Year
Howarth D. F., Rowlands J. C.
29
Rock Mechanics and Rock Engineering
1987 Article
4
117 Realising Capacity
Meyhew J.
118 Recording Drilling Parameters in Ground Engineering
Pfister P.
119 Relating Horsepower to Drilling Productivity
Givens R., Williams G., Wingfield B.
120 Relationship Between Thermal Stability and Molecular Structure of Polynitro Arenes
Type
World Coal
2004 Article
Geodrilling
1985 Article
International Society of Explosives Engineers 1996G
1996 Article
Keshavraz N. H., 4 Pouretedal H. R., Semnani A.
Indian Journal of Engineering and Material Science
2009 Article
121 Review of Rock Drillability and Borability Assessment Methods
Howarth D. F.
12
Transactions International Journal Mining and Metallurgy
1986 Article
122 Rock Abrasiveness Testing for Tunneling
West G.
10
International Journal of Rock Mechanics, Mining Science., & Geomechanics
1989 Article
123 Rock Cutting and It’s Potentialities As a New Method of Mining
Cook N. G. W., Toughlins N. C., Weibols G. A.
20
Journal African Institute of Mining and Metallurgy
1968 Article
124 Rock Drillability Related to a Roller Cone Bit
Morris R. I.
Society Petroleum Engrinners Paper 2389
Article
125 Rock Fragmentation and Optimization of Drilling Tools (Chapter 6 in Fracture of Rock Ed. Aliabadi M. H.)
Mishnaesky L. Jr., 36
Computational Mechanics Publications
1998 Article
126 Rock Fragmentation by Concentrated Loading
Simon R.
Failure and Breakage of Rock (Ed. Fairhurst, C.) AIME
1967 Article
20
15
(Continued)
Book.indb 732
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
Type
127 Rotary Club: A Turn Around the Latest Large Diesel and Electric Rotary Blasthole Rigs.
Watt J.
5
World Mining Equipment
2002 Article
128 Rotary Drilling
Willaimson T. N. 25
Surface Mining (Ed. Pfleider E. P.), SME/AIME
1968 Article
129 Ruggedized Computer Controls Mining Equipment
Mintchell G. A.
2
Control Engineering
2001 Article
130 Selection of Powder Factor in Large Daimeter Blast Holes
Eloranta J.
3
Eloranta & Associates Inc.
1995 Article
131 SSI Kaltex Drill Adviser
Kerr Cameron
7
World Mining Equipment Conference
1999 Article
132 Stemming Selection for Large Diameter Blastholes
Eloranta J.
4
Eloranta & Associates Inc.
1994 Article
133 Studies on Drillability of Rocks by Rotary Drills
Kinoshito, S
6
Journal Institute of Mining
1956 Article
134 Swedish Approach to Contour Blasting
Holmberg R, Persson R. A.
Proceedings Fourth Conference of Explosives and Blasting Techniques, ISEE
1978 Article
135 Taming the Dust Devil: An Evaluation of Improved Dust Control for Surface Drills Using Rotoclone Collectors
Page S. J., Organiscak J. A.
5
Engineering & Mining Journal
1995 Article
136 Technical Review of Aquila’s Drill Monitoring Control and GPS Guidance System
Hendrinks C.
14
World Mining Equipment Conference
1999 Article
137 Tests for Rock Drillability
Furby J.
6
Mining and Quarry Engineer
1964 Article
(Continued)
Book.indb 733
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
138 The Cercher Abrasivity Index and It’s Relation to Rock Mineralogy and Petrology
Sauna M., Peters T.
7
Rock Mechanics Vol. 15
1982 Article
139 The Concept of Specific Energy in Rock Drilling
Teale R.
17
International Journal of Rock Mechanics and Mining Sci. Vol. 2
1965 Article
140 The Drillability of Rocks
Singh D. P.
6
Mining Science and Engineering
1973 Article
141 The Effect of Apparatus Size and Surface Area of Charge on the Impact Strength of Rock
Rabia H., Brook N.
9
International Journal of Rock Mechanics and Mineral Science
1981 Article
142 The Effect of Some Drilling Variables on the Instantaneous Rate of Penetration
Outman H. D.
13
Transactions SME/AIME Vol. 217
1960 Article
143 The Effects of Blasting on Crushing and Grinding Efficiency and Energy Consumption
Workman L., Eloranta J.
10
Web Article
2005 Article
144 The Hole Picture
Carter Russ
6
Coal Age
2002 Article
145 The Mechanics of Rock Cutting
Nishimatsu Y.
10
International Journal of Rock Mechanics and Mining Science
1972 Article
146 The Perfect Cleaning Theory of Rotary Drilling
Maurer W. C.
10
Journal of Petroleum Technology
1962 Article
147 The Potential for Unifying Drilling, Blasting and Downstream Operations by the Application of Technology
Lyall Workman J. 14
Calder & Workman, Inc.
148 The Prediction and Reduction of Abrasive Wear in Mine Excavating Machinery
Atkinson, T., Cassapi V. B.
Journal Mechanical Engineering
7
Type
Article
1984 Article
(Continued)
Book.indb 734
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Publication Author/editor
Pages
Publisher
Year
Type
Mining Engineering
1978 Article
149 The Pro’s and Con’s of Rotary Blasthole Drill Design
Laswell G. W., Laswell B. J.
150 The Process of Thermal Spalling Behavior on Rocks – An Exploratory Study
Thirumlai K.
23
Rock Mechanics – Theory and Practice SME/AIME (Ed. Somerton)
1970 Article
151 The Stamp Test for Rock Drillability Classification
Wijk G.
8
International Journal of Rock Mechanics, Mining Science., & Geomechanics
1989 Article
152 Theory and Practice on Inclined Drilling for Surface Mining
Kochanowsky B. J. 18
Quarterly, Colorado School of Mines
1961 Article
153 Transfer of the Stress Wave Energy in the Drill Steel of Percussive Drill to the Rock
Simon R.
15
International Journal of Rock Mechanics and Mineral Science
1964 Article
154 Twist in the Tale (Rotary Rig Developments)
Woof M.
2
World Mining Equipment
1998 Article
155 Updating Air Practice for Better Open Pit Blasthole Drilling
Clarence Zink
19
International Society of Explosives Engineers 1997G
1999 Article
156 Wall Control Blasting in Open Pits
Crosby W. A.
Mining Engineering
1982 Article
157 Water Separator Shows Potential for Reducing Respirable Dust Generated on Small Diameter Rotary Blasthole Drills
Listak J. M., Reed W. R.
158 What’s Ahead for Blasting in the Next Millennium
13
Web Article
Article
Many Authors
Drill & Blast
2000 Article
159 What’s New in Drilling and Blasting
Freymuller J.
Rock Products
1999 Article
160 Which Blasthole Rig
Lyall Workman J., Szumansky M.
World Mining Equipment
1997 Article
(Continued)
Book.indb 735
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Table A15.2 Continued Name of the book/ Sr. article/product no. literature
Book.indb 736
Publication Author/editor
Pages
Publisher
Year
Type
161 Comparison of Rotary and In-Hole Motor Techniques for Drilling Horizontal Boreholes in Coal
Kravits S. J., Sainato A., Finfinger G. L.
36
U.S. Bureau of Mines Report of Investigation No. 8933, U.S. Department of Interior
1985 Report
162 Noise,Vibration and Airblast Management Plan (NVAM) Eagle Rock Quarry Project
Anonymous
7
For Polaris Minerals Corporation by Explosives and Rockwork Technologies Ltd.
2002 Report
163 Rotary Drilling Holes in Coalbeds for Degasification
Cervik J., Fields H. H., Aul G. N.
25
U.S. Bureau of Mines Report of Investigation No. 8097, U.S. Department of Interior
1975 Report
164 Rotary Drilling Techniques Used in the Beckley Coalbed
Goodman T. W.
16
U.S. Bureau of Mines Report of Investigation No. 9238, U.S. Department of Interior
1989 Report
165 Vibration Assessment – Eagle Rock Quarry Project
Anonymous
24
For Polaris Minerals Corporation by Explosives and Rockwork Technologies Ltd.
2002 Report
166 Optimum Bit Selection and Operation for The Rotary Blasthole Drilling Through Horizontal Drilling Rig (HDR) – A Case Study at KBI Murgul Copper Mine
Ergin H., Kuzu C, Balci C., Tunçdemir H & Bilgin N.
9
Web Article
167 Blast Design – Chapter 19 of Field Manual
Anonymous
90
U.S. Department of Interior, Bureau of Reclamation
168 Explosives – Chapter 2
Anonymous
18
Web Article
Book Ch
169 General Rock Anonymous Removal – Chapter 9
32
Web Article
Book Ch
Article
2001 Book Ch
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Index
Advantages of BECO Threads 120–122 Advantages of Electronic Initiating System 518, 519 Advantages of GPS System 353, 369–371, 591 Advantages of Inclined Blastholes 601, 602, 612–616 Advantages of Mechanical Charging System 571–575 Advantages of Screw Compressors 248, 252–254 Advantages of Sliding Vane Compressors 248–250 Air Blast 535, 545–549, 584 Air Blast Tricone Bit 100 Air Circulation 72, 74, 75, 85, 86, 91–95, 113, 301–303, 648 Air Decking 576, 583–585 Air Pollution 11, 221, 512, 535, 550–552 Air Shock Waves see Air Blast Altitude 23, 110, 223, 326–338, 353, 359, 640, 651–655 Annular Space 16, 127, 182, 296, 300, 301, 305, 307, 318, 319 Bailing Velocity 47, 122, 183, 288, 301–308, 310, 313, 314, 317, 319, 320, 323, 326, 647–649, 689 Bailing Velocity - Effect of Annular Space 302, 305, 307 Bailing Velocity - Effect of Blasthole Inclination 302, 305 Bailing Velocity - Effect of Density of Fragments 302, 304, 307 Bailing Velocity - Effect of Particle Roughness 302, 305 Bailing Velocity - Effect of Quantity of Water Injection 302, 306 Bailing Velocity - Effect of Rate of Fragmentation 302, 305 Bailing Velocity - Effect of Roundness of Fragments 302, 304
Index.indd 737
Bailing Velocity - Formulation 301–307 Bedding Plane Inclination 556, 557 Bedding Planes 41, 556, 558, 559 Bench Height 5, 6, 147, 433, 598, 606, 610, 612, 613, 623–625, 627–634 Bench Nomenclature 6 Bit Breaker 137, 679 Bit Design Feature Summery 103–105 Bit Metallurgy 100–103 Bit Wear Test 60 Bits 14, 20, 21, 46, 68, 69, 71, 72, 75, 83–115, 123, 124 + many more entries Bits - Air Circulation 72, 74, 75, 85, 86, 91–95, 113, 296–301 Bits - Blade Type 22, 83–85, 114 Bits - Claw Type 21, 22, 85 Bits - Drag Type 21, 28, 58, 83–87, 114, 115, 157, 263, 270, 272, 274, 286, 290, 679 Bits - Sealed Bearing 87, 94, 95, 107, 177, 178 Bits - Tooth Spacing 98, 99, 104 Blast Design 347, 370, 473, 597–638 Blast Initiator for 520–525 Blast Initiator for Detonation Wave System 523 Blast Initiator for Electric Detonation 521–523 Blast Initiator for Electronic Detonation 525 Blast Initiator for Hercudet System 523 Blast Initiator for Nonal System 523 Blast Initiator for Safety Fuse 521 Blast Initiator for Shock System 523 Blastability Index 355, 358, 359, 564 Blastability Index - Ashby (1977) 566 Blastability Index - Ghose (1988) 568 Blastability Index - Gupta (1990) 568 Blastability Index - Hainen and Dimock (1976) 566 Blastability Index - Han, Weiya and Shouvi (2000) 569 Blastability Index - Hansen (1968) 565
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738
Index
Blastability Index - JKMRC (1996) 569, 570 Blastability Index - Langefors (1978) 566 Blastability Index - Lilly (1986) 566–568 Blastability of Specimen 350, 555, 563, 564, 633 Blaster’s Multimeter 519 Blaster’s Ohmmeter 519, 520 Blaster’s Tagger 520, 521 Blasthole Camera 139–141, 536, 587, 679 Blasthole Charging Patterns 571, 575–588, 603 Blasthole Deviation 83, 124, 173, 315 Blasthole Dewatering 143, 144, 340, 679 Blasthole Dewatering Pump 143 Blasthole Drill - Advantages of Computerization 365–373 Blasthole Drill - Air Conditioners 181, 186, 188, 189, 219, 340, 387 Blasthole Drill - All Electric 74, 75, 163, 165, 173, 189, 213, 218, 220, 222, 255, 396, 401, 403, 407, 408 Blasthole Drill - Angle Hole Attachments 190, 196–198, 199 Blasthole Drill - Automatic Fire Suppression System 212, 213 Blasthole Drill - Automatic Leveling System 213–215, 350, 355, 359 Blasthole Drill - Automatic Lubrication System 78, 178–181, 338, 399, 402, 409, 414, 429 Blasthole Drill - Auxiliary Winch 154, 155, 175, 176, 181, 190, 204, 205, 387, 429, 430 Blasthole Drill - Batteries 175, 176, 218, 340, 432, 536 Blasthole Drill - Bit Changer 196 Blasthole Drill - Bit Lubricating System 177, 178, 402 Blasthole Drill - Cable Reel 165, 173–176, 189, 220, 387, 426 Blasthole Drill - Carrier Mounted 156, 157, 160, 161, 163, 166, 176, 177, 188 Blasthole Drill - Centralizer 190, 194, 195, 205, 207, 311, 372, 373 Blasthole Drill - Churn 66–68, 74, 153, 597 Blasthole Drill - Computerization 343, 345, 346, 348, 355, 359, 362, 363, 365, 423 Blasthole Drill - Computerization Hardware 362–364, 505 Blasthole Drill - Computerization Software 80, 140, 236, 346, 348, 362, 364, 365, 569, 570 Blasthole Drill - Crawler Mounted 19, 21, 67, 71, 74, 75, 82, 155–157, 159–172, 174, 175, 178, 345, 376, 377, 382, 392, 396, 411, 428, 687–689
Index.indd 738
Blasthole Drill - Depth Indicator 209 Blasthole Drill - Details of 155, 163, 173, 176, 181, 190, 200, 206, 311, 362, 687–689 Blasthole Drill - Diesel Engine 163, 164, 173, 175–177, 215, 217– 222, 228, 255, 330, 333, 341, 397, 401, 403, 408, 432, 687–689 Blasthole Drill - Diesel Hydraulic 156, 157, 175, 219, 222, 397, 400, 402, 420 Blasthole Drill - Double Wall Machinery House 181, 218, 219 Blasthole Drill - Driver’s Cab 154, 164, 173, 177, 188, 189 Blasthole Drill - Dry Dust Collector 156, 162, 164, 184–186, 233, 307, 340, 387, 430 Blasthole Drill - Engine Power 154, 173, 175, 176, 220, 221, 228–231, 330, 335, 376, 386, 387, 687–690 Blasthole Drill - Engine Starting Aid 217, 218 Blasthole Drill - Fast Fuel Fill System 207–209 Blasthole Drill - Fast Retract System 209, 210 Blasthole Drill - Feed Mechanism 21, 117, 124, 141, 190, 200–204, 223, 314, 377, 384, 430 Blasthole Drill - Fire Extinguishers 210, 211, 213, 425 Blasthole Drill - Fuel Tanks 164, 173, 189, 207, 209, 220 Blasthole Drill - Glow Plug 218 Blasthole Drill - Ground Load Bearing 385, 386 Blasthole Drill - Heaters 181, 186, 188, 189, 213, 215, 218, 219, 387 Blasthole Drill - Hydraulic System 75, 154, 155, 173, 177, 222, 335, 391, 428 Blasthole Drill - Hydraulic Test Station 215, 216 Blasthole Drill - Inclinometer 137–139, 679 Blasthole Drill - Language Name Plate 206 Blasthole Drill - Layout 153, 155, 163–165 Blasthole Drill - Leveling Jacks 74, 154, 155, 173–176, 214, 222, 339, 344, 345, 373, 377, 385–387, 430 Blasthole Drill - Lighting 166, 175, 176, 189, 206–208, 214, 387 Blasthole Drill - Machinery House 21, 154, 155, 173, 175, 176, 181, 182, 188, 215, 218, 219, 339, 340, 426 Blasthole Drill - Machinery House Heater 188, 215, 219
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Index Blasthole Drill - Main Frame 138, 154, 155, 157–159, 169, 173, 174, 176, 179, 181, 182, 186, 197, 200, 204, 214, 241, 296, 339 Blasthole Drill - Manifold Flame Heater 218 Blasthole Drill - Mast 21, 64, 66, 78, 80, 117, 137, 138, 153–155, 157–160, 162–164, 172–177, 179, 181, 182, 184, 186–190, 193–198, 200–205, 209, 231, 241, 242, 315, 339, 344–346, 357, 358, 368, 372, 376, 381, 385, 387, 392, 395, 407, 410, 411, 426–430, 602, 687–691 Blasthole Drill - Mast Ladder 190, 191 Blasthole Drill - Mast Raising Cylinders and Mast Braces 164, 189, 191, 196–199 Blasthole Drill - Oil Cooler 154, 164, 165, 173, 176, 181, 253, 254, 335 Blasthole Drill - Operator’s Cab 21, 79, 138, 154–156, 164, 165, 173, 174, 180, 182, 186–188, 191, 215, 216, 219, 355, 358, 363, 364, 368, 424, 430 Blasthole Drill - Pipe Changer 13, 78, 117, 123, 155, 158, 177, 186, 190, 191–195, 311, 313, 348, 384, 385, 392, 410, 411 Blasthole Drill - Primary Requirements 375, 376 Blasthole Drill - Prime Mover 154, 155, 173, 175, 176 also see Engine Power Blasthole Drill - Procurement 368, 419–424 Blasthole Drill - Protective Coating on Steel Components 220 Blasthole Drill - Radiators 173, 181, 340, 416 Blasthole Drill - Remote Propel Control 214, 215, 344 Blasthole Drill - Rotary Head 19, 70, 74, 78, 117, 129, 130, 142, 143, 154–156, 158, 176–178, 181, 182, 186, 190, 193, 194, 198–200, 202, 203, 215, 220, 223, 232, 242, 344, 352, 377, 379, 383, 384, 387 Blasthole Drill - Specifications 408, 421, 422, 687–691 Blasthole Drill - Tool Handling Jib 190, 204, 206 Blasthole Drill - Tow Hook 210 Blasthole Drill - Transformer 165, 173, 175, 176, 189, 215, 220, 221, 333–335, 403, 409 Blasthole Drill - Truck Mounted See Carrier Mounted Blasthole Drill - Types 153–163, 220, 420 Blasthole Drill - Video Camera System 215–217 Blasthole Drill - Welding Outlet 213
Index.indd 739
739
Blasthole Drill - Wet Dust Control 182–184, 186 Blasthole Drilling - Choice of Method 27–29, 420 Blasthole Drilling - DTH 19, 20, 26–28, 77, 129, 248, 254, 311, 420, 631, 687 Blasthole Drilling - History 61–82 Blasthole Drilling - Peculiarities 22, 23 Blasthole Drilling - Rotary 13 15, 20, 21, 23, 26, 28, 29, 46, 64, 68, 71, 72, 74, 75, 77, 78, 80, 82, 83, 86, 117, 119, 121, 124, 125, 128, 129, 132, 134, 136, 144, 145, 147, 153, 155–161, 163, 164, 168, 169, 172, 174, 176–178, 180–182, 184, 187, 189, 190, 196, 198, 204, 207, 209, 210, 213, 215, 217, 221, 223, 229, 230– 232, 241, 247, 248, 251, 264, 270, 280, 286, 295, 308, 317, 322, 326, 330, 335, 336, 339, 341, 343, 345, 346, 348–350, 355, 357–359, 362, 375–379, 381, 383, 385–387, 389, 392–394, 396, 401–404, 419–423, 425 Blasthole Drilling - Rotary Percussion See DTH and Top Hammer Blasthole Drilling - Top Hammer 18–20, 24, 26–28, 75, 77, 260, 631 Blasthole Drilling Patterns 13, 22, 566, 589–591, 603, 608, 633 Blasthole Firing Patterns 520, 588, 589, 592–595 Blasthole Inclination 15, 138–140, 147, 148, 160, 197, 198, 213, 302, 305–307, 318, 319, 352, 359, 377, 380, 381, 384, 385, 420, 422, 433, 456, 553, 556, 559, 565, 597, 598, 603, 609, 610, 612–616, 619, 632 Blasthole Inclinometer 137–139, 679 Blasthole Plugs 134, 583, 584, 588, 620, 679 Blasthole Sampler 145–147 Blasthole Spacing 6, 12, 13, 64, 157, 433, 513, 530, 600, 601, 603, 607, 608, 617, 618, 623–625, 627–630, 632, 633, 635, 637, 638 Blasthole Stemming Height 421, 474, 508, 536, 548, 549, 579, 580, 584, 586–588, 603, 620, 624–629, 632, 633, 637 Blasting - Cast 569, 570, 599, 600, 604, 605, 632–634 Blasting - Conventional Bench 580, 598, 618, 626, 627 Blasting - Hazards 493, 507, 533–553, 595, 597, 615 Blasting - Presplitting 599, 600, 601, 632 Blasting - Rip Rap 583, 603, 629–631
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740
Index
Blasting - Secondary 124, 509, 556, 588, 599 Blasting - Snake Hole 549, 597, 598, 601, 602, 630–632 Breakout Torque 384–385 Burden 6, 12, 147–150, 442–444, 475, 513, 525, 530, 548–550, 563, 565, 568, 598, 603, 607, 608, 612, 614–619, 621–637 Burden Calculation of 621–624 Burden Measuring Instrument 147–150, 525 Cerchar Abrasiveness Test 57, 58 Collaring 215, 340, 344, 345, 369, 370, 431, 534, 588, 614, 623 Comparison of Blasthole and Oilwell Drilling 13 Comparison of Blasthole Drilling Methods 24, 25 Comparison of Crawler and Carrier 161, 162 Comparison of Drill Types 123 Comparison of Slewing and Pivoted Undercarriage 168, 169 Comparison of Water Injection with Dry Dust Collection 181–184 Compressed Air - Discharge 110, 229–231, 236, 238, 247–252, 254–257, 301, 308, 314, 317–320, 322, 323, 325, 326, 336, 337, 386, 655, 657–666 Compressed Air - Flow in Hose Pipes 240, 241, 245, 309, 657–659 Compressed Air - Flow in Pipe Fittings 231, 241, 243, 244, 651, 661–666 Compressed Air - Flow in Steel Pipes 233–240, 651–655 Compressed Air - Flow through Nozzles 91, 93, 182, 232, 233, 245–247, 296, 300, 309, 320–326, 336, 338 Compressed Air - Leakage 247, 248 Compressed Air - Measurements 256, 257 Compressed Air - Pressure Loss 117, 118, 177, 235, 236, 238–247, 251, 256, 257, 309, 310, 322, 323, 336, 338, 352, 373, 652, 655, 657–666 Compressor 47, 154–156, 159, 164, 165, 173, 175–178, 181, 182, 186, 188, 223, 225, 227–231, 233, 236, 238, 246–257, 295, 296, 301, 308–310, 314, 317–320, 322, 323, 326, 330, 335–337, 386, 387, 401, 403, 407, 409, 410, 414, 415, 423, 428, 429, 432, 433, 480, 583, 651, 688, 690 Compressors Rotary Screw 176, 229, 230, 248, 250–254, 337 Compressors Sliding Vane 248–250, 254 Concepts of Drilling 16–22
Index.indd 740
Delay Element 512, 513, 515–517, 588 Detonating Cord 456, 472, 494, 508–513, 516, 517, 530, 536, 548, 583 Detonator 15, 62, 449, 451, 454, 472–475, 496, 498, 502, 507–510, 512–523, 525, 530, 535–537, 575–577, 580, 581, 588, 626 Detonator - Electric 62, 498, 513–516, 520, 521 Detonator - Electronic 518–520 Detonator - Non Electric 473, 516, 517 Dozer 9, 11, 210, 348, 350, 370–372, 426 Dragline 11, 74, 153, 222, 317, 348, 350, 355, 371, 393, 420, 422, 570, 586, 599, 600, 604, 605, 609–612, 616 Drill Collars 71, 315 Drill Operation 343–345, 349, 355, 372, 400, 408, 429 Drill Operation - Pipe Handling 78, 344, 345, 386, 387 Drill Operation - Leveling 111, 154, 174, 213, 214, 343, 344, 346, 354, 369, 386, 387, 423, 428, 429 Drill Operation - Tramming 343, 344, 346, 350, 372, 423, 426, 427, 434, 689 Drill Pipe 13, 16–19, 21, 24, 72, 74, 78, 82, 108, 110, 111, 113, 117–124, 128, 130, 134, 136, 137, 141, 143–146, 153–155, 157, 162, 185, 186, 191–196, 201–205, 207, 209, 232, 236, 256, 296, 300, 301, 309–320, 322, 323, 345, 350, 352, 359, 362, 368, 372, 379, 383, 392, 405, 410, 411, 416, 426, 429, 430, 433, 648, 649, 679, 687, 689, 691, 693–697 Drill Pipe - Choice of 122, 123, 310–319 Drill Pipe - Discarding 317, 320 Drill Pipe - Fabricated 13, 117, 118, 315, 679, 694, 696 Drill Pipe - Integral 13, 117, 118, 125–127, 315, 326, 679, 694, 696, 702, 703 Drill Pipe - Sizes 693–697 Drill Pipe - Surface Treatment 122, 310–313 Drill Stem Wrench 136, 137, 186, 345, 679 Drilling Angle Hole 25, 128, 184, 194 Drilling Cost 307, 389, 392, 393, 404, 406–408 Drilling Cost - Operating 199, 398–404, 406, 407, 413, 418 Drilling Cost - Owning 369, 393–399, 406, 412, 413 Drilling Cost - Salvage Value 112, 393, 396, 397, 413 Drilling Diamond 71, 73, 260, 263, 284–286, 290, 348, 358 Drilling Oilwell 12–14, 23, 64, 71, 86, 119, 132, 306, 315, 348, 349
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Index Drilling Methods 12, 17, 21, 23, 24, 26, 29, 259, 260, 393 Drilling Recorders 78, 108, 110, 111, 349–351 Drilling Records 432–434 Effect of Extremely Cold Weather 25, 215, 217–221, 250, 339–341, 407 Effect of Extremely Hot Weather 338–340 Effect of Heavy Rainfall 327, 338, 340, 377 Effect of Heavy Snowfall 327, 338–340 Effect of Scarce Air on Air Compressors 335, 336 Effect of Scarce Air on Blasthole Flushing 336–338 Effect of Scarce Air on Diesel Engines 330–333 Effect of Scarce Air on Drill Lubrication 338 Effect of Scarce Air on Electric Motors 333–335 Effect of Scarce Air on Hydraulic System 335 Effect of Stormy Wind 338, 339 Explosive - Ammonium Nitrate 448, 460, 466, 472, 473, 475, 477, 478, 480, 482, 485, 487, 488, 490, 498, 572 Explosive - ANFO 447–449, 451, 452, 460, 461, 473, 475, 477–486, 488, 490, 491, 525, 534, 565, 566, 568, 571–573, 575–579, 584, 585, 587, 603, 604, 608, 609, 617, 626, 628, 629, 632 Explosive - Blasthole Pressure 450, 452, 555 Explosive - Boosters 474, 475, 485, 507–509, 515, 559, 575, 576, 582 Explosive - Brisance Value 450, 455 Explosive - Charging Density 450, 455, 456, 555, 626 Explosive - Detonation Pressure 440, 450–452, 476, 477, 484, 487, 489, 491, 576, 577, 580, 622 Explosive - Dynamite 63, 447, 449, 450, 472, 473, 475, 477, 479, 485, 486, 488–491, 498, 499, 501, 507, 617 Explosive - Emulsion 452, 473, 475, 477, 479, 485–487, 502, 507, 534, 572–575, 577–580, 584, 587, 628, 629 Explosive - Fuel Oil 448, 473, 475, 478–481, 487, 572 Explosive - Heat of Explosion 449, 450, 453, 459, 464–466, 468, 469, 484, 487, 489 Explosive - Heavy ANFO 452, 482, 484, 485, 572 Explosive - Nitrocellulose 448, 450, 466, 475, 477, 488, 489, 501, 515 Explosive - Nitroglycerine 63, 488, 489
Index.indd 741
741
Explosive - Oxygen Balance 465–467, 484, 487, 489 Explosive - PETN 446–448, 450, 452, 455, 468, 475, 476, 507, 510, 512, 576, 578 Explosive - Primary 446–448, 473, 475, 476, 512, 514–516, 518 Explosive - Properties of 445, 449–461, 471, 493, 503, 628 Explosive - RDX 447, 450, 455, 468–470, 475, 476 Explosive - Slurry 448, 451, 461, 473, 475, 477, 479, 485, 486, 502, 534, 573–575, 577, 579, 617 Explosive - Sodium Nitrate 448, 475, 478, 485, 488–490 Explosive - Specific Gas Volume 450, 453 Explosive - Strength of explosive 452, 453, 465, 469, 470, 496–498 Explosive - Sulfur 448, 457, 472, 475, 479, 488–490, 573 Explosive - Thermochemistry 453, 463–470 Explosive - TNT 446–448, 450, 452, 455, 468–470, 472, 475, 476, 507, 576, 578 Explosive - Volume of Products of Explosion 465, 467, 468 Explosive - Water Resistance 450, 458, 459, 476, 484, 485, 487, 489, 491, 507, 535 Explosive Strength 452, 453, 465, 469, 496–498, 500 Explosive Strength - Absolute and Relative 452, 453 Explosive Strength - Weight and Bulk 452 Feed Force 17, 18, 21, 24, 25, 28, 71, 77, 80, 85, 89, 95, 105, 112, 117, 118, 123, 124, 128, 130, 141, 142, 154–156, 157, 159, 162, 174, 198, 200, 209, 261, 273–280, 282–284, 287, 289, 311, 314, 315, 317, 318, 344, 345, 348, 349, 352, 357, 358, 360, 361, 376–378, 382–385, 405, 615, 672, 679 Feed Force Measuring Kit 679 Feed Force on Drag Bits 21, 85, 86, 274 Feed Force on Tricone Bits 105 Feed Mechanism - Chainless Rack and Pinion 200, 203, 204 Feed Mechanism - Double Drum and Wire Rope 200 Feed Mechanism - Fixed Sprockets and Chain 200–202 Feed Mechanism - Hydraulic Cylinder and Chain 200, 201 Feed Mechanism - Hydraulic Cylinder and Wire Rope 200, 201 Feed Mechanism - Rack and Pinion with Chain 200, 202, 203
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742
Index
Firing Patterns - Channel Delay Pattern 592–595 Firing Patterns - Echelon Delay Pattern 592, 593 Firing Patterns - Flat Face Pattern 592, 593 Firing Patterns - Sinking Hole Pattern 592, 595 Firing Patterns - V Pattern 592, 593 Flushing 13, 47, 59, 71, 72, 84, 107, 113, 123, 124, 145–147, 154, 181, 183, 186, 223, 231, 251, 254, 275, 279, 288, 295–314, 322, 323, 330, 336–338, 348, 380, 402, 419, 429, 615 Formation Fracture Below a Drill Bit 274–280 Formation Logging Units 348, 349 Gas Laws 225 Gas Laws - Amagat’s Law 225, 227 Gas Laws - Avogadro’s Law 225, 227 Gas Laws - Boyle’s Law 225, 227, 229 Gas Laws - Charles’ Law 225, 227, 468 Gas Laws - General Gas Law 227 Gas Laws - Guy Lussac’s Law 225, 226 Gas Laws - Joule’s Law 225, 226 Gas Laws - Poisson’s Law 225, 226, 229 Geological Cycle 34, 35 GPS 80, 110, 346, 347, 350–356, 362, 365, 369–372, 591, 679, 681, 683 Ground Vibrations 452, 527, 534, 537–545, 550, 595, 603, 614 Ground Vibrations - Damage 537, 541–545 Ground Vibrations - Prediction 539–541 Hardness 9, 17, 27, 46, 50–58, 60, 75, 90, 103, 105, 106, 113, 145, 160, 203, 204, 273, 282, 285, 357, 379, 380, 567, 605, 616, 667, 668, 670–673, 675, 676, 710, 711 Hardness Scales 675–678 Hercudet System 517, 521, 523–525 High Speed Camera 525, 529, 549 Hot Hole Meter 525, 526, 534 Humidity 327, 329–333, 337, 340, 480, 482, 503 Influence of Rock Density on Blasting 554, 555 Influence of Rock Porosity on Blasting 555 Influence of Rock Strength on Blasting 554 Kelly 64, 71, 72, 157–159, 429, 489 Laser Measuring Instruments 679 Lifting Bails and Hoisting Plugs 134, 135
Index.indd 742
Loading ANFO and Heavy ANFO by Truck 572, 573 Loading Slurry and Emulsion ANFO by Truck 573, 574 Machine Digging 7, 8, 11 Maintenance 78, 130, 162, 163, 169, 172, 186, 199, 201–204, 219, 221, 241, 254, 369, 390, 399, 400, 408, 414, 421, 423, 424, 428, 429, 431, 432, 434, 536, 571 Measurement While Drilling 349, 355, 359 Metamorphism 34, 37–39, 45 Mineral Occurrence 44–46 Minerals 1, 2, 3, 10, 12, 32, 34, 35, 37–39, 43–46, 52, 54, 55, 61, 62, 263, 264, 311, 389, 440, 460, 551, 561, 586, 587, 614, 675, 713 Miniature Drill Test 57–60 Misfires 534, 535–537, 603, 614 Movement of Drill by Using GPS 353, 354 Operating a Rotary Blasthole Drill 428–431 Penetration of a Wedge 272 Penetration of an Indenter 292 Penetration Rate Estimation of 282–294 Periodical Check Ups 432 Power for Air Compression 228–231 Priming 458, 580–582, 586–588 Priming - Bottom 580–582, 586 Priming - Multi Point 580, 582 Priming - Special Rock Mass Conditions 586–588 Priming - Top 580–582, 586 Pulldown 71, 111, 129, 133, 165, 176, 181, 188, 201, 319, 433, also see Feed Force Pullout Force and Speed 384 Recovery Tools 144, 145, 679 Ripping 5, 7–9, 11, 12, 74, 137, 555 Rock - Abrasiveness 46, 57–60 Rock - Brittleness 46, 54, 57, 341 Rock - Coefficient of Internal Friction 46, 55, 56, 265, 710 Rock - Compressive Strength 2, 4, 8, 46–49, 51, 54, 57, 105, 111, 273, 273, 285–288, 294, 421, 554, 563, 569, 570, 582, 605, 628, 672, 673, 710, 712, 715 Rock - Density 31, 32, 36, 42, 45–47, 282, 293, 297–302, 304, 307, 308, 318, 539, 553–555, 563, 566, 569, 571, 577, 604, 616, 617, 710, 712–714 Rock - Drillability 46, 281, 282, 292, 379 Rock - Failure Theories 265–270
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Index Rock - Igneous 10, 34–36, 39, 45, 304, 313, 314, 318, 710, 712 Rock - Metamorphic 10, 34, 35, 37–39, 304, 313, 314, 318, 710, 712 Rock - Sedimentary 7, 10, 11, 27, 34–38, 45, 48, 304, 313, 314, 318, 444, 556, 557, 581, 586, 710, 712 Rock - Swell Factor 57 Rock - Tensile Strength 46, 53, 50–52, 262, 282, 442, 510, 511, 554, 622, 675, 710–712 Rock - Toughness 17, 46, 54, 101, 273, 282, 421, 672, 673, 711 Rock Breaking 7, 11, 12 Rock Fracture by Crushing 11, 39, 95, 104, 112, 181, 258, 261–264, 279, 281, 440, 469, 583, 599, 600, 606, 614 Rock Fracture by Impacts 262 Rock Fracture by Scratching 262–264 Rock Fracture by Spalling 263, 440, 442, 561 Rock Fracture by Vaporization 263 Rock Mass - Classification 560–569 Rock Mass - Dip 557, 558, 563, 603, 604 Rock Mass - Folds 42, 43, 556, 557, 608 Rock Mass - Formation of 18, 37, 45, 273, 437, 440, 560, 608 Rock Mass - Joint 282, 554, 561 Rock Mass - Joint Spacing 282, 554, 561 Rock Mass - RMR 562–564 Rock Mass - RQD 562, 563 Rock Mass - RTQI 562, 564 Rock Mass - Strike 557, 559, 563, 603, 604 Rock Mass - Unconformities in 42, 43, 556 Rock Mass - Voids in 123, 146, 449, 533, 556 Rotary Deck Bushing 136, 186, 679 Rotary Speed of Drag Bits 85, 86 Rotary Speed of Tricone Bits 105, 110–112, 277, 278, 281, 282, 287, 289, 290, 292, 294, 320, 344–346, 348–352, 357, 359, 360, 366–368, 378–380, 382, 406, 428, 429, 433, 570, 688, 690 Safety Fuse 62, 472, 474, 509, 520, 521, 535 Safety Items - Close Fitting Shirt and Long Pants 150 Safety Items - Ear Plugs or Ear Muffs 150 Safety Items - Face Shield 151 Safety Items - Rain Coat 151 Safety Items - Respirator 151 Safety Items - Safety Glasses 150, 151 Safety Items - Safety Toed Shoes 150 Safety Items - Safety Vest 150
Index.indd 743
743
Safety Precautions 434 Seismograph 472, 473, 525, 527, 538 Sensitivity of Explosive 450, 453–456, 459, 467, 476, 477, 480, 481, 484–487, 489, 500–503, 511 Sensitivity of Explosive to Detonator Strength 454 Sensitivity of Explosive to Friction 454 Sensitivity of Explosive to Gap 455 Sensitivity of Explosive to Heat 454 Sensitivity of Explosive to Shock 454 Setting Up the Rotary Blasthole Drill 428 Shock Absorbers 128–134, 199, 405, 707, 708 Shock Absorbers - External 130–133, 707, 708 Shock Absorbers - In the Hole 132, 134 Shovel 7, 11, 27, 74, 78, 153, 170, 222, 317, 348, 350, 355, 370–372, 393, 406, 420, 422, 570, 586, 592, 593, 598, 599, 604, 605, 609, 610, 612, 616 Size and Weight of the Drill 376–378 Soil Profile 43, 44 Specific Energy 280, 281, 289, 290 Stabilizer 21, 85, 91, 110, 113, 117, 123–128, 137, 183, 309, 316, 317, 380, 384, 405, 417, 426, 679, 699–706 Stabilizer - Integral Blade 125–127, 679, 702, 703 Stabilizer - Replaceable Blade 125, 126, 704, 705 Stabilizer - Roller 125, 127, 128, 679, 699, 700 Stabilizer - Welded Blade 125, 126, 679, 700–702 Structures - Extrusive 36, 40–42, 45 Structures - Hypabyssal 36, 40, 41 Structures - Plutonic 36, 40–42, 555–557, 560 Sub - Crossover 117, 128, 137, 241, 405, 407, 679 Surface Activities 43 Tests on Explosives 493–505 Tests on Explosives - Ballistic Mortar Test 497 Tests on Explosives - Bullet Impact Test 505 Tests on Explosives - Cap Sensitivity Test 502 Tests on Explosives - Chronograph Test for VOD 494, 495 Tests on Explosives - Cratering Test 497 Tests on Explosives - Critical Diameter Test 504
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Index
Tests on Explosives - Critical Height Test 505 Tests on Explosives - Cylinder Compression Test 499 Tests on Explosives - D’Autriche Test for VOD 494 Tests on Explosives - Effect of Wetness Test 503 Tests on Explosives - Electrostatic Discharge Sensitivity Test 503 Tests on Explosives - Fiber Optic Sensor Test for VOD 495 Tests on Explosives - Friction Sensitivity Test 502 Tests on Explosives - Gap Sensitivity Test 502 Tests on Explosives - Heat Resistance Test 500, 501 Tests on Explosives - Internal Ignition Test 503 Tests on Explosives - Koenen Test 505 Tests on Explosives - Material Compatibility Test 504 Tests on Explosives - Optical Measurement of VOD 495 Tests on Explosives - Plate Dent Test 497 Tests on Explosives - Shock Sensitivity Test 500 Tests on Explosives - SLIFER Test for VOD 495 Tests on Explosives - Traulz Lead Block Test 498 Tests on Explosives - Underwater Test 497 Tests on Explosives - Vacuum Stability Test 504 Thread Protectors 134, 135 Tool Joints 118–122 Tool Joints - API Regular 91, 100, 119, 120, 122, 128, 706 Tool Joints - BECO 100, 119–123, 128, 385 Toxic Fumes 450, 456–458, 551
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Transporting the Drill 431 Tricone Bits 14, 21, 68, 72, 75, 83, 86–115, 128, 177, 196, 262, 270, 276, 277, 279, 280, 286, 290, 294, 311, 320, 322, 379, 383, 402, 405, 417, 420, 669 Tricone Bits - Carbide Insert 20, 84, 87, 88, 95, 101, 102, 105–109, 115, 276, 292, 293, 669–673, 679 Tricone Bits - Cones 18, 21, 69, 71, 86–91, 95, 98–102, 108, 111–114, 275, 294, 383, 429 Tricone Bits - IADC Code 105–107, 379 Tricone Bits - Milled Tooth 87, 88, 95, 101, 112 Tricone Bits - Nozzles 92, 93, 107, 320–323, 325 Tricone Bits - Performance 88, 91, 95, 103, 109–112, 282, 286, 379, 404 Tricone Bits - Reasons for Failure 112–114 Tricone Bits - Rejection of 108, 109, 111, 115 Tricone Bits - Shirttail 92, 95, 97, 109, 113, 124 Tricone Bits - Steel Tooth 87, 105–107, 115, 276, 667, 679 Truck for Blasthole Charging 571–575 Tungsten Carbide 10, 20, 21, 46, 59, 60, 68, 75, 83, 84, 87, 88, 94, 99, 101, 102, 112, 115, 125–127, 286, 292–294, 667–673 Undercarriage 154, 156, 162–170, 174, 181, 376, 395 Undercarriage - Pivoted Crawler 169 Undercarriage - Slewing Crawler 166–169 VOD meter 525, 528 Volcanic Activities 35, 40 Water Jetting 7, 8 Why Compressed Air for Flushing 295
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