International Mining Forum 2004

INTERNATIONAL MINING FORUM PROCEEDINGS OF THE FIFTH INTERNATIONAL MINING FORUM 2004 FEBRUARY 24–29, 2004, CRACOW—SZCZ...

Author: Jerzy Kicki | Jacek Sobczyk

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INTERNATIONAL MINING FORUM

PROCEEDINGS OF THE FIFTH INTERNATIONAL MINING FORUM 2004 FEBRUARY 24–29, 2004, CRACOW—SZCZYRK— WIELICZKA, POLAND

International Mining Forum New Technologies in Underground Mining Safety in Mines Edited by

Jerzy KICKI AGH—University of Science and Technology, Department of Underground Mining, Cracow, Poland Polish Academy of Sciences, Mineral and Energy Economy Research Institute, Cracow, Poland Eugeniusz J.SOBCZYK Polish Academy of Sciences, Mineral and Energy Economy Research Institute, Cracow, Poland

A.A.BALKEMA PUBLISHERS LEIDEN / LONDON / NEW YORK / PHILADELPHIA / SINGAPORE

Copyright © 2004 Taylor & Francis Group plc, London, UK. 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 written prior permission from the publisher. Although all care is taken to ensure the integrity and quality of this publication and the information herein, no responsibility is assumed by the publishers nor the author for any damage to property or persons as a result of operation or use of this publication and/or the information contained herein. Published by: A.A.Balkema, a member of Taylor & Francis Group plc. http://www.balkema.nl/ and http://www.tandf.co.uk/ This edition published in the Taylor & Francis e-Library, 2005. “To purchase your own copy of this or any of Taylor & Francis or Routledge’s collection of thousands of eBooks please go to http://www.ebookstore.tandf.co.uk/.” ISBN 0-203-02413-3 Master e-book ISBN

ISBN 90 5809 607 6 (Print Edition)

Table of Contents

International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 Preface Organization

viii xi

Lectures Application of IntelliMine® for Underground Mine Management J.Paraszczak Development Trends in Mine Hoisting and Drainage G.G.Litvinsky Technology of Underground Coal Mining in the U.S.A.—State of the Art C.Bruniany Prospects of Development of Highly Productive Coal Extraction Technologies V.I.Bondarenko, O.M.Kuzmenko & R.O.Dychkovsky The Effectiveness of Support of Weak Rock with TA2 Tubular Anchors V.I.Bondarenko, G.A.Simanovich, I.A.Kovalevska & V.V.Porotnikov Equipment Selection for Mechanized Mining as a Function of Physical, Mechanical and Deformation Properties of Rock M.Ljubojev, M.Stjepanovic, M.Ivkovic & S.Perendic The Influence of Physical Parameters on Mechanizability of Longwall Mining of Coal K.Oraee & R.Pourkhandani A Fuel-Energy System Based on Mining Preparation and Underground Burning of Coal Layers G.Gayko Exploitation of Technologically Generated Methane Deposits by Means of Surface Wells A.E.Vorobyov, T.V.Chekushina, G.A.Balykhin & A.D.Gladush

1 13 23 36 46 51

63

71

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Methane as a Source of Energy in an Air-Conditioning System in “Pniowek” Coal Mine N.Szlązak, A.Tor, A.Jakubów & K.Gatnar New Technologies of Coal Bed Methane Preliminary Recovery S.V.Slastounov Geomechanical Problems in Simultaneous Exploitation Both Open Pit and Underground at Minera Michilla, II Region Antofagsta—Chile A.R.Carvajal, C.G.Fernández & J.M.Carmona Underground Geotechnology of Exploitation of Polymetallic Ores A.E.Vorobyov, K.G.Karginov & T.V.Chekushina Technological Advances in Underground Miing of a Stratified Copper Deposit in Poland J.Butra Possibility of Employing Mechanical Extraction in Thin Copper-Ore Deposit Conditions L.Horoszczak & L.Ziętkowski Experience and Practical Aspects of Utilizing a Shrinkage Method of Extraction at “Kazimierz-Juliusz” Coal Mine in Sosnowiec S.Gajos, M.Urbaś & T.Lamot Application of the SF6 Tracer Gas in Identifying Mine Air Flows through Abandoned Workings Sealed from the Ventilation System P.Buckwald & Z.Jaskólski Risk of Coal Dust Explosion and its Elimination K.Lebecki, K.Cybulski & A.Szulik Methane Control in Coal Mining Safety System N.O.Kaledina Rock Bursts—Preventive Measures Undertaken in the Polish Mines J.Dubiński & W.Konopko Technical and Economic Possibilities for Permanent Limitation of Water Inflow to Mine Workings R.Kuś The Attempt to Apply Radar Interferometry InSAR in the Monitoring of the Impact of the Ore Deposit Exploitation in LGOM (Lubin Copper Mining Area) E.Popiołek, C.Bachowski, A.Krawczyk & P.Sopata The Control of Mining Damage in China Yu Xueyi & Zhao Binchao Limiting Inflow of Water to Operating Shafts by Application of Permanent HydroInsulating Screens R.Kuś Problems and Prospects of Development of the Coal Mining Enterprises in the Donetsk Region A.R.Vovchenko, V.G.Grinyov & D.M.Zhitlyonok Organizational—Economic Models of Investment Activity at the State-Owned Coal Mining Enterprises of Ukraine

93

112 120

137 151

168

178

192

203 215 223 242

257

269 276

286

291

V.I.Logvinenko & O.Yu.Kuzmich Author Index

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Preface

International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 We live in a most fascinating world that is constantly changing. The saying: “We are changing the earth even before we are able to understand it” is most appropriate. Unfortunately among the reasons of this change is also mining activities. Profound interference by mining in the environment doesn’t encourage many friends. However it is hard to imagine any well-functioning economy without mineral resources. So you may ask what kind of intentions were behind the organizing of the International Mining Forum. I believe that it is possible to combine the development of the mining industry with the implementation of sustained development (defined by the Brundtland Commission as “development that meets the needs of the present generation without undermining the capacity of future generations to meet their needs”). The reflection of a Polish scientist, former rector of the AGH—University of Science and Technology, professor Walery Goetel “What industry breaks down, technology has to repair, that what industry threatens, technology has to defend” remains valid. It is possible to accomplish this goal through changes and further development of mining technologies. This huge challenge has to be realized by the mining sector, continuously seeking new solutions in the area of its influence on the environment. Solutions that will also be compatible with work safety, of course. This work comprises technical papers that were presented at the International Mining Forum in Kraków—Szczyrk—Wieliczka, Poland, held on 25–28 February 2004. The two major themes of this seminar were: – New technologies in underground mining – Safety in mines This book is addressed to researchers and professionals who work in the fields of underground mining technology, rock engineering or management of mines. The topics discussed in this book are: 1. Trends in the mining industry 2. New solutions and tendencies in underground mining technology 3. Rock engineering problems in underground mines 4. Utilization and exploitation of methane 5. Prevention measures for the control of rock bursts in Polish mines 6. Current problems in Ukrainian coal mines

The International Mining Forum was held thanks to the support of the Chair of Underground Mining, the Faculty of Mining and Geoengineering of the University of Science and Technology (AGH), Mineral and Energy Economy Research Institute of Polish Academy of Science in Cracow, KGHM Polska Miedz SA, FTT Stomil Wolbrom, Jastrzebska Coal Company, Katowice Coal Holding, CUPRUM Ltd., BLASTEXPOL Ltd., GLINIK Ltd., MIDO Ltd., MMDE ZOK Ltd. The organizers would also like to express their gratitude to all the other persons, companies and institutions, who helped in bringing the Forum into being. We hope that the Forum will throw a new light on mining and will contribute to the exchange of interesting experiences. Jerzy Kicki Chairman of the Organizing Committee 2004

Organization

International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 Organizing Committee: Jerzy Kicki (Chairman) Eugeniusz J.Sobczyk (Secretary General) Artur Dyczko Jacek Jarosz Piotr Saługa Krzysztof Stachurski Katarzyna Stala-Szlugaj Advisory Group: Prof. Władymir Bondarenko (National Mining University, Ukraine) Mr. Wojciech Bradecki (State Mining Authority, Poland)—Chairman of IMF 2004 Prof. Jan Butra (CUPRUM Ltd., Poland) Dr. Alfonso Carvajal (Universidad de La Serena, Chile) Prof. Piotr Czaja (AGH—University of Science and Technology, Poland) Prof. Bernard Drzęźla (Silesian University of Technology, Poland) Prof. Józef Dubiński (Central Mining Institute, Poland) Prof. Jaroslav Dvoracek (Technical University VSB, Czech Republic) Prof. Paweł Krzystolik (Experimental Mine Barbara, Poland) Prof. Garry Litwinski (Donbass Mining and Metallurgy Institute, Ukraine) Prof. Eugeniusz Mokrzycki (Polish Academy of Sciences, MEERI, Poland) Prof. Roman Ney (Polish Academy of Sciences, MEERI, Poland) Prof. Jacek Paraszczak (University of Laval, Canada) Prof. Janusz Roszkowski (AGH—University of Science and Technology, Poland) Prof. Stanisław Speczik (KGHM Polska Miedź S.A., Poland) Prof. Anton Sroka (Technical University—Bergakademie, Germany) Prof. Mladen Stjepanowic (University of Belgrade, Yugoslavia) Prof. Antoni Tajduś (AGH—University of Science and Technology, Poland) Prof. Kot F.v.Unrug (University of Kentucky, USA) Dr. Leszek Wojno (Australia)

Application of IntelliMine® for Underground Mine Management

Jacek Paraszczak Department of Mining, Metallurgical and Materials Engineering, Université Laval Quebec City, Canada International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: Mining cost depends to a significant degree on equipment availability and utilization. Both may be substantially improved through better management of the equipment fleet. An introduction of mine management systems in open-pit mining has been in a focus of attention for some time, but until recently, no such solutions were available for underground mines. This has changed however with an introduction of the underground version of IntelliMine®, a real-time mine management system developed by the American company Modular Mining Systems International (MMSI). This paper overviews its objectives and principles. System components and utilities are described in detail. The paper presents also some information on several applications in different mines all over the world, including an input from end-users. The paper concludes with a critical evaluation of system advantages and its current status.

1. INTRODUCTION The trends toward mechanization, increasing level of automation, and larger equipment have resulted in an increased reliance on the equipment used in the mines. As the equipment is getting more productive, but at the same time more complex and capitalintensive, deficient reliability and maintenance as well as inefficient operation often prevent utilization of its full capacity. As equipment-related costs constitute a high percentage of a total production cost (with maintenance alone accounting for 30% and more; Campbell 1995, Knigths 1999), it is extremely important that it is highly efficient and provides as much useful work as possible (Wiebmer and Widdifield 1997). This can be achieved, among the others, through the optimized asset management. In other

International mining forum

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industries such as for example manufacturing, the performance and utilization of equipment are usually closely monitored down to the lowest level of detail. Modern plants are controlled and optimized using programmable intelligent devices that enable displaying real time performance to operators and prompt them to make informed decisions to maximize production. Through continuous monitoring of equipment performance in real time, it becomes possible to quantify decision making, react quickly to unpredictable changes in circumstances, effectively manage and optimize the steps in the production cycle. Even though the mining industry is lagging behind manufacturing sector in this respect, a lot of R&D work has been done on this aspect in the last several years. With regard to open-pit operations a significant progress has been achieved in loading and haulage equipment matching as well as truck dispatching. Following the implementation of a Global Positioning System (GPS), mine management systems have become far more effective in maximizing the effectiveness of loading and haulage equipment fleets. Underground mining environment however, is by far more restrictive in what concerns an application of several technologies and more demanding in the sense of working conditions and constraints. Therefore, a direct technology transfer of open-pit mine management systems to underground workings represents a serious challenge. It seems however, that this changes gradually. The American company Modular Mining Systems International (further in the text referred to as “MMSI”) that has developed a well-known IntelliMine® mine management system for open-pit mines has made considerable progress in adapting it to underground mine environment. 2. OVERVIEW OF INTELLIMINE® AND ITS UTILITIES IntelliMine® is a real-time mine management system. Its underground package provides a platform to manage the underground production cycle and associated information collection. Its main role consists in allocation of load-haul-dump machines (LHDs), trucks or trains, controlling traffic, as well as monitoring equipment and manpower. It enables also comprehensive reports. It functions similarly to its surface counterpart, except in data transmission and equipment location. A typical system uses the mine’s existing radio communications network for data telemetry, reducing infrastructure investment. Given the impossibility to use GPS underground, mobile equipment tracking is achieved using radio frequency identification (Voss 2000). Basic components of the system as well as its main utilities will be briefly described in the following sections. 2.1. System components The main components of the system include (Hackwood et al. 1999, Voss 2000, Lewis 2003): – central computer work station, – field computer systems (FCS), – data radio network based on the leaky feeder and/or microcell radios with advanced token-ring technology, – RFID—radio frequency identification tags (beacons).

Application of intelliMine® for underground mine management

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A central computer in the system information center sends assignments and other data to LHDs, trucks or trains, and allows production controller (dispatcher) interaction. Field computer systems (FCS) are located at every piece of mobile equipment (LHDs, trucks or trains) that is supposed to make part of the whole system. FCSs receive and display data from the central computer and accept driver input. They consist of a touch-screen Graphics Console (GC) and a microprocessor-based communications hub. A data radio network passes information between the central workstation and the FCSs at the speed of 9600 bps. In most cases, this network is based on leaky feeder technology. Finally, Radio Frequency Identification (RFID) transponder tags serve to locate mobile equipment. They are mounted at strategic points such as: draw points or loading areas, weigh bridges, dumping points (tips), drifts and haulage ways rims as well as in other places if necessary (figure 1). At the earlier stages of development (Finsch mine in South Africa, see also later), equipment tracking was done by means of infrared beacons, linked together via twisted pair cable in a loop configuration (Mining Technology 2003).

Figure 1. Equipment tracking with Radio Frequency (RF) identification tags (after Lewis 2003) 2.2. Fleet management Fleet management is achieved by using a Dispatch®—a key component of the IntelliMine® underground package. It provides automatic work assignments to equipment operators based on a “tramming schedule” or a set of rules. With the former method, the shift foreman simply enters a list of appropriate loading and dumping locations for each LHD, truck or train. During the shift, the system works through the list, generating assignments accordingly. The second method (by “rules”) is more


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complicated. It involves developing an algorithm that can calculate equipment assignments based on the mining plan, which provides a list of active loading areas, material types, tonnage requirements, and dumping locations. The system ranks the loading areas by priority and assigns the closest available piece of equipment to the highest priority area that is not presently served by another piece of equipment. Loaded equipment is directed to the most efficient dumping location that is set to accept the equipment’s loaded material type. Equipment can also be “locked to” or “barred from” loading and dumping points as well as regions within the mine, allowing the production controller to fine-tune assignments. Upon receipt of disturbance information and other status information, the system automatically presents the next priority assignment. The production controller may reallocate resources based on other changes that occur during the course of the shift. The shift database records the change in status and updates the mine database with reason codes. Messages are displayed on transaction and exception screens to the production controller on the control computer. Incomplete activities roll over into the next shift for prioritizing. The production controller may reprioritize scheduled tasks at any time. It is important to stress that he can override either method and provide assignments to operators “manually”. At any point, the Production Controller can use graphical displays on the central computer to determine (in real time) the status, the location, and productivity of each piece of equipment compatible with IntelliMine®. The system collects real-time information from the payload systems on the loaders and trucks and it may prepare equipment productivity reports in terms of tonnage loaded and dumped (see figure 2), production tonnage reported by source and destination as well as average loading, hauling, dumping and traveling times.

Figure 2. Example of an LHD production chart generated by the system (after Lewis 2003)


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Moreover, since the system records all equipment activity, it provides also other valuable information that can be used to improve operations and enhance productivity. Other available reports include equipment availability and utilization, as well as total equipment “ready”, “delay” and “down” time sorted by reason codes (figure 3).

Figure 3. Truck performance pie chart generated by IntelliMine® (after Lewis 2003) The information is available to standard reporting utilities for analysis. Analysis of the reports leads to the establishment of best practice and Key Performance Indicators (KPI) for use in continuous improvement initiatives. 2.2.1. Maintenance The Maintenance Utility (Hackwood et al. 1999) consists in data entry-based software programs designed to help mines respond quickly to equipment failures (equipment “down”), and to closely monitor and record equipment repairs that occur. It also offers a wide range of real-time and historical maintenance reports. This utility includes: – Pre-shift safety checks by operator. – Downtime management. – Resource allocation. – Service schedules. – Defect management. – Work order interfaces. – Work order “light”. – Downtime analysis: MTBF, MTTF, MTTR, reliability, frequency & duration calls. Tracked breakdowns (failures) trigger appropriate measures to bring a disabled unit back to an “up-state” (“available”) quickly. The system tracks equipment status and alerts the


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Production Controller to problems. The shift database records “delay” and “down” transactions for display to the Production Controller on the central computer. Specific delays drive adequate actions based on priority, schedule, rules or other criteria accordingly to user requirements. Depending on the breakdown, the system posts messages directly to maintenance (electrical, mechanical) personnel, who in turn dispatch the right personnel to fix the equipment. The Maintenance Utility monitors and records equipment status and enables assignment of resources and priority by the Production Controller. The system optimizes also the Preventative Maintenance (PM) schedule for the week and measures compliance. The system accepts date or hour-driven calendar and can reschedule PM when the machine is “down”. The utility tracks also maintenance backlog in the Maintenance Utility database. 2.2.2. Fuel service utility The system computes fuel consumption for each piece of haulage equipment based on tank capacity, and equipment type (manufacturer and model). Equipment “idle” or “operational” status data forms the basis for fuel consumption values. When a piece of equipment is operational (hauling), travel distance, road grade, and “empty” or “full” status determine the fuel consumption value. The system assigns fuel stops when required versus time-of-day assignment, improving equipment and personnel utilization.s 2.2.3. Vehicle Health Monitoring System (VHMS) It is a real-time preventive maintenance tool that monitors and reports engine and other machine parameters, if equipped with third party (for example Original Equipment Manufacturer—OEM) monitoring devices and sensors. Performance data from the latter is collected and then relayed to a central computer via radio link and, optionally to a Graphics Console (GC) for an operator to view. Mine personnel can view real-time graphics and text monitoring screens from a central computer work station and use these for maintenance reports. The system displays collected data on PCs or workstations in a variety of text reports, strip charts, and histograms. (Hackwood et al. 1999). Maintenance personnel can also use VHMS to monitor equipment and maintenance indicators from a PC at the workshop (Voss 2000). Up to the best knowledge of this author, the underground application of VHMS has not yet been reported. 2.3. Mine Manager 2.3.1. Ore constrol utility It contains all the logic required to manage the material flow (movement) in the mine. The Ore Control Utility allows the mine to define its ore body in the main mine database and to store associated composite parameters for blending and reporting, as well as stope tonnage for inventory control (Hackwood et al. 1999).


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2.3.2. Muck Flow Utility The system monitors and records the ore and waste flow, respective passes levels, as well as source and destination status. Material tracking is automatic from source to destination. The system calculates pass levels by monitoring and recording “tonnage in” and “tonnage out” values. The actual level may be reconciled at any time. The records for the stope indicate its status at any given point. The system accepts a daily fill schedule into the activity schedule database to plan a particular stope backfilling. Reporting capability provides reconciliation of actual status to plan (Hackwood et al. 1999). 2.3.3. Blend Utility The Blend Utility is a valuable production tool that allows mines to control material blending at crushers, passes and temporary re-muck bays. The utility provides two main blending methods: continuous and batch. Continuous blending uses a control mass variable and component or category parameters to provide a flow of material from the loaders or trucks to the crushers, ore passes and re-muck bays, so that the material remains continuously within set blending limits. This method is for use in mines whose current production scheme requires crusher material to be within blending limits at any point in time. In contrast to it, batch blending only ensures that material production will meet the blending limits upon production target achievement. This method allows maximum efficiency of loader resources (Hackwood et al. 1999). 2.4. Crew management 2.4.1. Crew Scheduling Utility It allows a mine to enter and schedule personnel in the respective database. The utility tracks, correlates and prints data regarding a crew’s schedule (including also approved vacation dates). This utility tracks personnel profiles, their equipment operating qualifications and the shift rotation. The system verifies also personnel qualifications and safety training with regard to a given piece of equipment (Hackwood et al. 1999). 2.4.2. Lineup Utility It allows a mine to schedule personnel and equipment for the upcoming shift. Operator qualification, seniority, and any other mine requirements govern the lineup procedure. This utility allocates personnel to their assigned equipment at shift startup. It verifies personnel qualifications for the equipment and whether personnel log-in procedures ensure compliance with the plan. Lining up the next shift usually occurs during the last hour of the current one. The utility features a separate form for lining up haulage equipment and auxiliary equipment. The latter category also includes re-muck bays, crushers, ore and waste passes, chutes and stopes (Hackwood et al. 1999).


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2.4.3. Safety Training Utility It allows tracking the training of operators on all types of equipment throughout the mine. It records also the safety training and certification of all its employees, their retraining and recertification. Personnel’s reporting is accomplished by this utility. Updates of status are limited and depend on the source of data input; either by the GC mounted on the assigned equipment or by the shift supervisors or leaders. The production controller may enter voice status reports as required (Hackwood et al. 1999). 3. INTELLIMINE® APPLICATIONS IN UNDERGROUND MINES 3.1. Finsch diamond mine—South Africa The first reported application of IntelliMine® in an underground mine environment was a Finsch diamond mine operated by DeBeers in South Africa. The system intended to dispatch LHDs was installed and commissioned in 1990s and was based on the open-pit dispatching code. At that time the underground voice communication systems were still relatively new. At present, the underground production comes from the open-stope blocks 2 and 3 on the 430 m and 510 m levels, and a block-caving system in block 4 on the 630 m that began production in 2003. The mine uses 12 t-capacity LHDs that dump the broken ore to eight ore passes (Mining Technology 2003). At the Finsch mine, safety is improved by means of a unique application of proximity detection readers. These are installed on LHDs and interfaced to the onboard FCSs. Tags that are fitted to personnel’s cap lamps and underground equipment transmit a signal that is detected by a reader. A warning message is displayed on the operator’s Graphical Console (GC) indicating the proximity of the person or vehicle. At the same time the warning message is transmitted via data radio to the mine management system, where it is stored in the central databases (Hung et al. 2001). 3.2. Stillwater mine—USA Stillwater palladium/platinum mine is located close to the city of Nye in the state of Montana. Its mining methods include ramp-and-fill and cut-and-fill. The main benefits of the system have been the tracking and scheduling of waste-muck handling using LHDs and haul trucks having a capacity of 15–20 tonnes. In order to determine the last position of any piece of equipment fitted with the FCS, over 400 strategically placed RFID tags were used. The equipment location is graphically displayed on the surface control room computers where production controllers are aware of equipment movement in real time. In a large, extensive and growing mine such as Stillwater, this feature saves time and effort, helping increase equipment utilization. The system schedules also waste remucking activities by assigning trucks as a group, along with an LHD as a working team. Once a pile of muck is successfully removed the group moves on to the next most important assignment (Hung et al. 2001).


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3.3. Kidd Creek mine—Canada Falconbridge Kidd Creek is copper, zinc and silver mine situated in Timmins, Ontario. It employs long-hole stoping with cemented rock-fill or waste fill methods. Mucking is performed using 5 LHD machines with a 5,8 m3 buckets. Backfilling is performed by 4,2 and 5,8 m3 LHDs, using waste hauled in with several 26- and 30-tonne trucks from the development ends of the new Deep Mine Project. Through the implementation of the MMSI system the mine saw opportunity to improve substantially LHD and truck utilization. The equipment operators logged on to the system at the beginning of the shift by entering their payroll numbers on a touch-sensitive graphic console (GC) in the cabin. They received work assignments from the production controller sent via a data radio transmission to the console. All “transactions” to and from the production controller and mobile equipment were logged on the surface computers, allowing the mine to follow the actual utilization of the fleet. The system has been found helpful to limit the negative impact of “unavailability” of draw points (once empty or blocked with oversize blocks) and ore passes dumps (when full or hung-up). In such situations, it enabled the production controllers to reassign LHDs to other work areas, saving time and money. It has had also a positive impact on safety, since via real-time communications with machine operators, the production controllers knew right away if there were any problems, and they were able to deal with them immediately (Hung et al. 2001). Although the system as such has had its advantages, underground communication was a critical issue. The Kidd Creek’s system relied on leaky feeder technology. The mine’s leaky feeder network is considered the largest and the most complicated in North America and, unfortunately, it did not lend itself to reliable data communication. After several attempts to improve its use, the mine decided to put IntelliMine® on hold (Mracek 2003). 3.4. Olympic Dam mine—Australia Olympic Dam mine (Southern Australia) operated by WMC Resources is the country’s largest underground mine, producing copper and uranium plus gold and silver as byproducts. Its production rate is approximately 10 million tonnes per year. Mining method is by large sublevel open-stoping. Implementation of IntelliMine® was completed in June 2000 and the initial focus was for data collection from the fleet of so-called “ore pass” LHDs, composed of Caterpillar-Elphinstone R2900s. They work on the extraction level of the stopes, and feed one of the 12 main ore passes, which lead to the major rail haulage horizon. The project has been substantial and included installation of 50 hubs and graphic consoles on LHD machines, trucks and drill jumbos, placement of RFID tags at 700 locations, setting up a separate data channel, computer servers in the mine control room, as well as personnel training. Vehicle location and status information is stored on the servers. This allows managers to locate equipment on computer screens, view shift production, direct the equipment fleet either manually or automatically, identify and correct problems and run reports on total shift production (Voss 2000).


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3.5. Palabora mine—South Africa Palabora (operated by Palabora Mining Company, a subsidiary of Rio Tinto Group plc) is a copper mine that switched from open-pit to underground production in 2002. It was to operate a full block cave as opposed to the more common panel caving technique. Reserves in excess of 250 Mt lie above the production level and the production tunnels will have a long active life of 20 years. Its expected full capacity was expected to reach 30000 tonnes/day. In order to meet this target, 344 drawpoints arranged in 21 production tunnels were planned. In the time of writing however, due to inability to clear efficiently drawpoints blocked by poorly fragmented large rocks, the actual production figures were far lower at the level of 10000–12000 t/day (Anon 2003a). A fleet of 6 R-1700G Supa 14S LHDs (Caterpillar Elphinstone), having a bucket capacity of 14 tonnes (Anon 2003b), is mucking from drawpoints and delivering ore, via a grizzly, directly into one of four large jaw crushers each equipped with two tipping points. Crushed ore is then conveyed to the shaft and hoisted to surface stockpiles feeding the metallurgical plant. MMSI has been working with Palabora Mining Company to implement an underground Auto-Dispatch® system. Its main objective is to maximize efficiency while meeting operational constraints such as for example blending, minimum and/or maximum production for drawpoints, demand for all dump points or pre-existing conditions (for example locking of LHDs to dump points), etc. It ought to meet also several operational constraints Auto-Dispatch® is intended to generate assignments of haulage equipment to loading and dumping locations, assign automatically: refueling and scheduled maintenance as well as secondary breaking equipment to production tunnels. This is supposed to be done in the way to maximize material throughput (MMSI 2001, MineAutomation web page 2003). 3.6. Other reported applications Other reported applications of IntelliMine® system included assigning LHD machines between loading points and dumps by following pre-determined work schedules (the mines: El Teniente in Chile and Koffiefontein in South Africa) and locomotive/trains dispatching (Great Noligwa, Kopanang, Val Reefs mines in South Africa). In the latter ones, locomotives are assigned to pull ore from every chute while minimizing queues, following priority, meeting tonnage level goal, and minimizing chute waiting times (Hung et al. 2001). In the Val Reefs mine a locomotive fleet has been reduced from 22 to 8 units, while the productivity rose by 30% (Lewis 2003). IntelliMine® was also supposed to be implemented in the Craig mine (operated by Falconbridge) situated in the Sudbury basin (province of Ontario, Canada). For this nickel mine using cut-and-fill method, optimal face and equipment utilization are critical to reduce mining cost and increase efficiency (Hung et al. 2001). The implementation of IntelliMine® however, has been slowed down however by a lengthy strike and no information about current status of the system in this mine has been obtained.


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4. FINAL REMARKS AND CONCLUSIONS The equipment and mine management system developed by MMSI for underground applications is aimed at the substantial improvement of the mining process, particularly in what concerns loading and haulage operations. It is expected to assist mine executives with the tasks concerning management of equipment fleet and crew: – Production task management. – Mobile supervisor. – Monitoring. – Maintenance. – Task management (auxiliary equipment, fuel service). – Crew location and security. – Operator feedback. – Incident reporting. – Performance analysis and organizational Key Performance Indicators (KPIs). Concerning the cases where it has been implemented, it seems to have met the expectations in some mines such as Finsch and Val Reefs (South Africa), as well as in Olympic Dam (Australia). In advantageous circumstances and environment, the system performs well and it is capable to bring substantial benefits in terms of: – Better utilization of equipment and manpower. – Minimization of contention time. – Drawpoint equalization. – Improved management and planning through comprehensive reporting. – Better vehicle monitoring and periodic maintenance. It looks however, that communication may become a key concern in the places where mine layout is complex as it has been the case in Kidd Creek. This proves that the technology transfer from open-pit to underground is not always as seamless as we would like it to be. In all, it may be stated that Dispatch® and IntelliMine® have not yet reached the stage of the “off-the-shelf” product that could be successfully implemented in every underground mine. The system however, has definitely some interesting potential to help improve equipment management and utilization and, consequently, reduce mining cost. For this reason all the future developments of the MMSI systems should be watched closely by a mining community. REFERENCES Anon 2003a: Crossing Bridges. Mining Magazine 189 (2):54–57. Anon 2003b: Helping Palabora. Mining Magazine 189 (1):17–18. Campbell J.D. 1995: Uptime—Strategies for Excellence in Maintenance Management. Portland: Productivity Press.


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Hackwood H.J., Henderson A., Fantin P. and Eisenhour B. 1999: Underground Mining Cycle Control System—a Platform for Management. Proc. 5th ISMMA and Telemin 1 Conference, Sudbury, Ont., Canada, June 14–17. Hung J., Gerhart D., Pix A. and Hackwood J. 2001: Mine Management System Development in page: www.mining-technology.com/projects/finsch/ Modular Mining Systems International 2001: Palabora Mining Company to Install Modular’s IntelliMine® System at its New Mine in South Africa. News release posted on the MMSI Web site: http://www.%20mmsi.%20com/ne%20ws_06.%20shtml Mracek B: Underground Mines. Knights P.F. 1999: Analysing Breakdowns. Mining Magazine 181(3):165–171. Lewis M.: (Modular Mining Systems Intl., Tucson, AZ, USA) 2003. Private communication. Mine Automation webpage:www.mine-automation/mmsi/mmsi.html Mining Technology Web (Falconbridge—Kidd Creek Division) 2003. Private communication. Voss B. 2000: Underground Intelligence. World Mining Equipment, Vol. 24, No. 10, pp. 15–16. Wiebmer J. and Widdenfield L. 1997: Cost-Per-Ton Improvement Ideas for Underground Equipment. Paper No. 97–169 presented at SME Annual Meeting, Denver, Colorado—February 24–27. Paper No. 97–169.

Development Trends in Mine Hoisting and Drainage

Garry G.Litvinsky The Donbass Mining and Metallurgical Institute. Alchevsk, Ukraine International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: A brief analysis of the base problems and shortcomings both of mine hoisting and drainage is given. Inappropriateness of their utilization in future mining industry is shown. As alternate version a principally new engineering solution for overcoming engineering difficulties in mine hoisting and drainage is offered. The new scientific and engineering idea is based on usage of a force volumetric hydraulic drive for transport of mineral and mine water. The new engineering system executes and unites functions of mine hoisting and drainage. Hydro-jack hoisting and drainage (HJHD) removes shortcomings of current constructions, has considerably smaller powerconsumption and cost. Base design features are shown and a technological comparison of alternate versions demonstrates advantages of the new engineering system HJHD. KEYWORDS: Mining industry, analysis, rope mine hoisting, drainage by pumps, hydro-jack hoisting and drainage (HJHD), performance characteristics

1. INTRODUCTION Winding plants are used to hoist people, material and machinery, broken mineral and waste rock. They are erected to make exploitation possible and are utilized both in vertical and incline shafts, both in coal and ore mines. However recently with increase of depth of mining essential shortcomings of present hoisting systems and drainage have begun to be exhibited. Of them first of all it is necessary to mention major energy consumptions and price, sharp decrease of performance parameters at increase of depth of mining. As the main occurrences of mineral resources, in particular of coal, are accumulated primarily on depths of more than 1000–2000 m, there is a problem of their


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effective mining. A solution to this problem to no small degree depends on an overall performance of hoisting and drainage. The purpose of this work is to present to the mining industry; both the construction and operation of a new, alternative engineering solution to mine hoisting and drainage. The main principle of the new system of hoisting of weights and water from deep levels is to dispose of rope hoisting and substitute it by hoisting with a force volumetric fluid drive by means of systems of hydraulic jacks hoisting a continuous string of kibbles. Especially it is necessary to point out a possibility of solving all the complex problems of mine drainage, if it is combined with new hydro-jack hoisting. Hydro-jack hoisting and drainage, in abbreviated form referred to as HJHD, is such a combined engineering system. Major advantages of HJHD are: independence of performance indexes from shaft depth, compactness, low cost, low mass, decrease in energy consumptions etc. These qualities give a new urge to re-estimate possibilities of undertaking mining operations at great depths and essentially to increase performance of mining of minerals. 2. ANALYSIS OF PERFORMANCE CHARACTERISTICS OF TRADITIONAL MINE ROPE HOISTING All present types of hoisting are divided to major (for transportation of broken mineral) and supplementary (for lifting of waste rock, hoisting of material, machinery, people). Additionally, taking into account mine shaft inclination it may be divided to vertical and inclined; number of ropes—to one and multi-rope; conveyances—to skip, kibble and cage winder; use of counterweight—to balanced and unbalanced (Fedorov 1979). Constructions of winding plants to the present time have achieved high level of sophistication. A hoisting system comprises among others (Mining Encyclopaedia 1989): – hoisting equipment: winding engine with actuator, conveyances, rope/ropes, shear wheels, safety devices, tilting platforms and landing keps, handling machines etc., – mining buildings and works: a winder building, headgear, silos, rail and conveyor transport near the shaft, a shaft equipped with guides for skips and cages, shaft sump, loading bins etc. Key feature of all winders is that they use hoisting ropes. The rope or ropes are the weakest link of this complex engineering system; they impose stringent restrictions on its main performance parameter—mass of lifted weight and winding depth. At the depths of mining (1000–1500 m and more), reached to the present time, the socalled critical length of rope L0, which equals to the depth, at which rope is torn up under its own dead weight, starts to be reached. So, for man-hoisting the critical rope length equals: L0=σz/mγ0=160·103/(9·78)=2280 m where σz—tensile strength of metal of strands of a rope, σz=160−180 MPa; m—factor of safety of a rope, m=9; γ0—specific weight of a rope, γ0=78 kN/m3. Thus, already at depths of more than 1000–1200 m half of a rope’s strength is spent on carrying its dead weight. And if take into account the weight of hoisting block and skip, the performance of rope hoisting appears on these depths even less (about 0,4–0,5).

Development trends in mine hoisting and drainage

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Relevant parameter of performance of hoisting is the energy consumption to lift 1 ton of weight 1000 m up. Theoretically, minimum required energy to achieve this equals: E0=10 kN·1000 m/3600=2,78 kWh Because of periodic nature of the operation due to alternating cycles of acceleration and braking, currently used winding engines use much more energy. For example, for a multirope winder MK 3,25×4 with an installed power of motors of 3 MWt and hourly hoisting capacity of 450 tons, the specific consumption of energy for hoisting 1 ton of weight for 1000 m equals E=6,67 kWh, which yields an engineering coefficient of hoisting performance equal to: η=E0/E=2,78/6,67=0,42 From here follows, that at depth of 1000 m only 42% of power used by a multi-rope winding plant will be utilized to lift the payload. Within the same limits between 0,4– 0,45 lies the engineering efficiency of other winding plants, being smaller for greater winding depth. Thus, the present engineering solutions of mine hoisting have severe key shortcomings that put in doubt a possibility of their usage as primary machinery for the future mines: 1) an ineffectiveness of usage of a cable rope to lift loads from great depths, 2) poor bearing capacity of cable rope imposing limitations on a volume of lifted weight, 3) technical difficulties of creation of high-performance hoists (800–1000 tons/hour and more), 4) high specific energy consumption per unit of hoisted weight, 2,2–2,4 times greater than theoretically calculated as required, 5) cyclic operational mode creating complications for automatic and manual control, dynamic loads on supporting members of construction and essential parts of machinery, 6) major area required for the installation and big weight of the construction as a whole, complicated design, high complexity and duration of construction work and servicing, small reliability, 7) prohibitively high cost of lease of the area of the mine surface which was taken up by hoist buildings, unhandiness and complication of mine technical buildings and hoist buildings, large number and high cost of underground mining work and underground communications (service lines) near the shaft. It is apparent, that the engineering limitations, intrinsic to modern rope hoisting, are severe inhibiting factor to modernizing underground mines. It is necessary to search for a solution to this engineering limitation by discarding the major component of construction of mine hoisting systems, the one which till now seemed unreplaceable—cable rope. 3. CONSTRUCTION AND OPERATION OF NEW MINE HOISTING An analysis of the most promising, from the point of view of the requirements of the future mining industry, development trends in mine hoisting was carried out in the Donbass Mining and Metallurgical Institute, with an overall objective to find a solution to


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engineering limitations intrinsic to existing rope hoisting. As the result a novel solution of this problem was offered—hydro-jack hoisting and drainage (HJHD), which revised the conventional approach of engineers to constructing mine hoisting systems. Let’s describe the basic structural components of the new hoisting system. Figure 1 shows HJHD in a cross-section of a shaft, in figure 2 the interaction between the side kibble and hydraulic jack is shown, in figure 3—magnification of unit A from figure 2, and in figure 4—part of the mine lifting apparatus in area of abutment station in shaft is shown. HJHD includes kibbles 1, built-in shaft guides 2 (for example, L-bars), hydraulic jack 3, installed on abutment stations 4, which are uniformly placed along the shaft (20–50 m apart from each other depending on HJHD’s work parameters). The telescoping rods of hydraulic jacks 3 are furnished with rotary stoppers 5, which can clamp to edges 6 of kibbles 1 on contact. Kibble 1 is built as a rigid tank, the back end of which has edges 6 and restricted on height lug 7, capable to enter extraction 8 below than disposed kibble and to snap beside were kibbles. The capacity of a kibble can be 1–2 cubic metres and more. The kibbles 1 are placed in shaft to form two columns, continuously driven through abutment stations 4. These are: load-carrying column 14, going up, and empty column 15, following downward. The abutment stations 4 are furnished with force braces 8, which are fastened to the shaft lining and stretched rope support 9 with turnbuckles 10, which retain station from above and are secured to rock-mass by anchors 11. Hydraulic jacks 3 are connected up with the help of a flaps system to pressure head and overflow hydromainlines 12, which are in turn linked to a stationary hydraulic pump installed on a surface (not shown). A self-propelled conveyance 13 is provided in the shaft for transport of material, machinery and people, and also for HJHD inspections. Together with a column of kibbles for hoisting weights 14, corresponding column 15 for hauling down empty kibbles 1 is provided in the shaft. The new hoisting system works as follows. When underground, kibbles 1 are loaded (e.g. with coal) with the help of a loading bin with a weighing device and move up between guides 2. A column of full kibbles 14 is thereby made up in the shaft. Hydraulic jacks 3, installed on stations 4, are periodicly hooked up to pressure head or over-flow hydromainlines 12 by flaps system and push the rods, on which rotary stoppers 5 are installed. At upward movement of rods of two hydraulic jacks 3, rotary stoppers 5 hook on sill 6 of kibbles 1 and force them up by multiple maximum length of travel of rods 5 of hydraulic jacks 3. The opposite movement of rods 5 results in a withdrawal of rotary stoppers 5 from sill 6 of kibbles 1 and their free slip on a side of a kibble 1 up to the moment, when they again hit under sill 6 of the next hoisting kibble 1.


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Figure 1. HJHD—view in a crosssection of shaft

Figure 2. Kibbles HJHD at interplay with rods

Figure 3. Close-up of unit A on figure 2—the principle of interaction between the stopper and the back end of the kibble


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As a rule, to maintain balance, two couples of hydraulic jacks 3 work simultaneously, placed across the diagonal, symmetrically in relation to the centre of the kibble column. Namely, when the rods of the first couple of hydraulic jacks 3 move forward, hoisting the kibble column, the rods of the second couple execute backward motion, retracting the rods. Thus, each couple of hydraulic jacks 3 hoists the kibble column 1 between stations 4 for the length of travel of hydraulic jacks (about 1 m) per thrust. Due to sequential operation by each couple of hydraulic jacks 3 kibble columns can be moved almost continuously, ensuring continuous operation. Pressure coming from the mass of the kibble column between stations 4 (from 40 up to 100 tons) is transmitted through hydraulic jacks partially to braces 8, which rest on the shaft lining, and partially to the stretched rope support 9 with turnbuckles 10, which support the stations from above and are attached by means of anchors 11 to the solid. All stations 4 (there can be 10–20 of them depending on depth of shaft and spacing intervals between stations) along shaft are identical, except for the lowest, where the exchange operations with kibbles are executed. Due to two identical columns of kibbles (full 14 and empty 15) working in parallel of hydromainlines, an ideally balanced circuit of mine hoisting is achieved, where the effective power is not spent for lifting of dead weight of conveyances. 4. PERFORMANCE CHARACTERISTICS OF THE NEW HOISTING SYSTEM Let’s determine productivity of HJHD for a shaft. If the mass of load in a kibble equals 1 ton, and the kibble moves in the shaft with a velocity of 0,3 m/s (i.e. every 3 sec to surface is hoisted 1 kibble with a load of 1 ton), the hourly capacity of hoisting equals: P=1·0,3·3600=1080 tons/hour. Extremely relevant for the future underground energy system is, that productivity of HJHD and its design features do not depend on winding depth, but only demand increasing the power of hydraulic pumps proportionally to depth. Taking into account that efficiency of a hoisting hydraulic-circuit system equals at least 0,8, we receive a required power of HJHD for a 1000 metres deep shaft: N=2,78−1080/0,8=3750 kW For depth of 2000 m it is necessary to double the power of motor engines. The engineering coefficient of performance of HJHD equals η=E0/E=2,78/3,48=0,81, and exceeds twice the value of this index for rope hoisting from depth of 1000 m. HJHD works continuously on hoisting weights without interruptions. Its output, depending on capacity of kibbles and power of a surface installation of hydraulic pumps, can vary from 300 to 1000 m3 and more of weight per hour that is 1800–6000 m3 per a 6hour shift. It considerably exceeds output of known present-day rope hoisting systems. The demanded continuous power of HJHD is considerably decreased (1,5–2 times) due to the continuous operation of driving motors, that slashes product costs, improves energy saving and ecology. In order to present performance data of a proposed HJHD system we shall compare it with a state-of-the-art multi-rope winder (CM 5×8) with a double-motor direct-drive


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actuator (see table 1). A generalized index of performance λ is introduced here. It allows estimating levels of performance of different mining techniques by means of the method explained in (Litvinsky 2003).

Table 1. Comparison of engineering indexes of HJHD and CM 5×8 winder Hoisting parameters

CM 5×8

HJHD

HJHD vs CM 5×8 index

1. Maximum depth H of hoisting, m

1600

unlimited

+

2. Hourly efficiency P, tons/hour

1000

1080

1,08

3. Installed power W, MW

10

6

0,6

4. Mass of machinery, tons

250

140

0,55

5. Mass of metallic structures Mo, tons

2500

1200

0,48

Generalized index of performance λ

0,56

1,44

2,50

5. PROBLEMS OF MINE DRAINAGE AND THEIR SOLUTION It is necessary to note, that due to high efficiency, HJHD will not be needed to hoist load for the whole shift. After the broken coal and waste rock are hoisted, HJHD can be utilised for the removal of underground water. Most importantly no restrictions are placed as to its acidity and solid particle content. Let’s review some problems of an underground mine using a traditional drainage system, which represents a complex and power-intensive engineering system. The main mine drainage is designed to pump out the total water inflowing into the workings. In deep shafts this is carried out in stages, with complex and costly pump stations situated on midlevels. They include intake pits, water-drainage installations, settlers and delivery lines with gate valves, apparatus of automation, check and security. The water-drainage installation, as a rule, consists of 3 identical pumps (working, standby and under repair), each of which is designed to pump-out the normal daily inflow in 20 hours. The efficiency of pumps vary between 68–78%, to safeguard the pumps against hydraulic impacts special shock arrestors are utilized. The basic shortcomings of a conventional drainage system are: – complicated construction of both pumps and auxiliary equipment and instruments, their low efficiency, high required power of motors, – unhandiness, high complexity, high cost of both servicing and reconditioning, – high requirements as to presence of solids in water and water acidity, – major quantity of costly pipelines in the shaft, – complicated inspections and servicing of pumping stations, – necessity to construct pump chambers and waterways, – high capital and running cost.


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Figure 4. Longitudinal section of shaft with hydro-jack hoisting As an example the basic operational parameters of main mine drainage systems, conventional and new alternatives, are compared. Inflow of water is taken to be 500 m3/h and depth of 1500 m. For a conventional main mine drainage system it is necessary to arrange for 3 stages of three pumps (for example, such as 12 MC-7) on each with applicable mining and engineering structures. For pumps 12 MC-7 it is necessary that water pH is 6–8; content of suspended solids is not more than 40 g/litre and maximum size of solid particles—not more than 10 mm. As an alternative we shall consider HJHD. The comparative performances are shown in table 2. From table 2 the undisputable advantages of combining hoisting and mine drainage functions in one design engineering system are visible. It makes a rather cost-intensive underground traditional system of drainage with its underground installations, pumps, pipelines etc obsolete. Thus, one more complex mining stationary installation, just as rope hoisting interfering with development of operations at great depths—water drainage


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is solved. Therefore it is possible to consider HJHD expedient to be applied for drainage and to call it HJHD (Hydro-Jack Hoisting and Drainage).

Table 2. Comparison of conventional pumping system and HJHD 12 MC-7×5 pumps

Water drainage parameters

HJHD

HJHD vs 12 MC7×5 index

1. Maximum pumping height H, m

625

not limited

+

2. Hourly efficiency P, m3/hour

800

1080

1,35

3. Installed power W with allowance for 3 pumping stations, MWatt

16,2

9,0

0,56

4. Mass of machinery, tons

210

140

0,67

0,54

1,46

2,70

Generalized index of performance λ

6. CONCLUSIONS As a result of the completed investigations and engineering developments it is possible to make the following conclusions: 1. The tendencies of increase of mining depth and growing production capacity of underground production units present new heightened requirements to mine hoisting and drainage. 2. The conventional complex engineering systems of mine hoisting and drainage will not be able to match the increased requirements and will hinder the advance of mining industry. In particular, rope-hoisting systems become completely unsuitable for depths of 1500–2000 m, and drainage systems have to cope with problems of multi-stage pumping. 3. A perspective promising way to solve the mine hoisting and drainage problems is utilizing HJHD—Hydro-Jack Hoisting and Drainage—powered by a force volumetric hydraulic drive and utilizing a system of hydraulic jacks instead of rope for hoisting. 4. Special advantage of this new hoisting system is its capability to function as an underground drainage system as well, thus combining two complex engineering systems in one design. 5. Hydro-jack hoisting and the drainage (HJHD) engineering indexes much exceed traditional systems. In particular, its coefficient of performance is 1,5–2 times higher, generalized index of performance—2–2,5 times, output—2 times and does not depend on depth, the cost is 2–4 times smaller, the indexes of compactness, reliability, safety, ecological compatibility are much higher. Thus, the proposed solution, a deviation from the conventional approach and urged by stagnancy in engineering thinking at a solution of problems in the field of mining stationary installations—hoisting and drainage—allows to get across to a new level of


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designing, building and exploitation of future underground mines. The introduction of HJHD marks an important step in winding theory and practice by offering a solution of one of the problems of deep winding by increasing safety and reducing installation and running cost. ACKNOWLEDGEMENTS The investigations were conducted in the Donbass Mining and Metallurgical Institute within the framework of the grant and at financial support of Ministry of Education and Science of Ukraine. The author expresses great thanks to the Organizing Committee of IMF and the Chairman of the Organizing Committee Jerzy Kicki for the opportunity to report on the obtained results at the Conference School of Underground Mining organized by the Polish Academy of Sciences. REFERENCES Fedorov N.M.: Mine Winding Plants. M: Nedra 1979–385 p. Litvinsky G.G.: About a Method and Criterion of an Estimation of a Technological Level of a Mining Technique. In Proceedings: The Technology of Underground Excavations. The Bulletin of Ukraine Academy of Building. Vol. 3. Donetsk:Nord-Press 2003—pp. 62–77. Mining Encyclopaedia. /chapter the editor E.A.Kozlovsky/. M.: The Soviet Encyclopaedia. Vol. 4. Ortinum—Sociosfera 1989–623 p.

Technology of Underground Coal Mining in the U.S.A.—State of the Art

Cas Bruniany Joy Mining Machinery, Franklin, PA. U.S.A. International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 1. INTRODUCTION With 275 billion tons of recoverable coal reserves that can be mined using existing technologies, United States is ranked number one in the World. At the current rate of production, recoverable coal reserves are sufficient for 250 years. American coal mines are the most productive in the World, and their performance is still improving every year. This discussion will focus on some major technology developments that greatly contribute to the success of the American coal mining industry. 2. US MARKET OVERVIEW Three major coal-producing regions are: 1. Appalachian (states of Kentucky, Pennsylvania, Ohio, West Virginia, Virginia, Alabama); 2. Mid-West (Illinois, Indiana, Kentucky and Texas); 3. West (Colorado, Wyoming, Utah, New Mexico, Montana, North Dakota, South Dakota). Coal production in the United States has been more or less steady for the past few years, at the level of about 1 billion tons per year. Small variations in a level of production are related to the demand for electricity that in turn depends on the state of economy and fluctuations in the weather conditions. Over ninety percent of coal mined in the United States is used for power generation. Coal is used to generate fifty percent of all electricity, and is by far the most significant fuel source. Coal is and will remain the primary fuel for electricity generation, but producers of electricity are increasingly turning to other fuels, especially natural gas. Recently


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however, due to substantial increases in gas prices, that trend seems to be reversing. Excess coal production capacity drives coal prices down, forcing less efficient coal mines to close. This means that to stay competitive, coal producers will continue to look for ways to increase their productivity and reduce costs. An outcome of this competitive pressure is a significant shift in US coal production from Eastern, underground coal to surface mined Western coal (fig. 3 and 4). This shift is primarily due to the competitive advantages (i.e. lower sulfur and lower production costs) of the thick, seam subbituminous coals being mined in Wyoming, Montana and the Dakotas.

Figure 1. U.S. Coal Fields

Figure 2. U.S. Electricity Generation

Technology of underground coal mining in the U.S.A.—State of the art

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Figure 3. US Coal Production

Figure 4. Production shift East to West Another significant trend is a steady reduction in the number of active coal mines, number of people working in the industry, and related growth in productivity (fig. 5 and 6).

Figure 5. Number of U.S. Coal Mines


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Figure 6. Productivity in U.S. Coal Mines The price pressure has resulted in continuous productivity improvements in both, room and pillar and longwall mines. These improvements have been realized by the closure of old mines with large infrastructures, closure of less productive mines, improved management approaches, technological advancements, improved information systems, and re-engineering of mining plans. 3. UNDERGROUND MINING METHODS Approximately 650 underground coal mines are active in the United States. Fifty-two longwalls are in operation in forty-eight mines (four mines use two longwalls simultaneously). All other mines use room and pillar mining methods. Both of these methods have had considerable effort directed at improving productivity on a cost/ton basis in the competition for survival and profitability. Of the approximately 330 million tons of underground coal produced in the U.S., one half comes from longwall mines, and the other half from room and pillar operations. Almost without exception, the highest total productivity in terms of total output, output/man/shift, or cost/ton results from longwall mining systems. While the number of longwall faces in the United States is declining steadily, total longwalls keep growing in size, power and output. The average longwall panel is 280 m wide and 2900 m long. The best longwalls produce on average 500,000 to 600,000 tons per month, and 25,000 tons production days are quite common. Continuous miners working in Room and Pillar systems have been achieving high cutting rates, but the system itself is plagued by bottlenecks in the form of inadequate haulage and roof bolting techniques. A considerable effort is being devoted to elimination of those bottlenecks, and results are very encouraging. Some room and pillar sections, where continuous miners work in conjunction with continuous haulage systems, produce on average of around 2,500 tons per shift, and achieve peak production of over 4,000 tons per shift. Such results make room and pillar mining very competitive, because the system is much more flexible than longwall system, and requires much lower capital investment.


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Development of longwall gate roads is also a function of room-and-pillar continuous mining system. No roadheaders and steel arches are used in American mines. Typically, U.S. longwall development is for three entries on each side of a panel, with entries 4.5 to 6.0 m wide, however situations exist where two entries or four entry developments are also used. The law in the United States allows no single entry systems. Roof is supported by roof bolts, 1.2 to 1.8 m long. In more difficult conditions, straps, wire mesh, and/or truss bolts are used. The fastest development rates are achieved using a 3-entry system. Typical development rate in a 3-entry system is 40 to 50 m per shift, while in 2 or 4 entry system that rate drops typically by 20%. All United States longwalls require entry development, as there are no advancing longwalls in the country. Continuous Miner productivity in longwall development sections, as high as it is, is lower than in a typical room and pillar system, and very often cannot keep up with retreat rate of longwall faces. Equipment manufacturers are facing a challenge of developing a system capable of developing longwall gate roads at a rate of 100 m per shift. In certain conditions, continuous miners combined with roof bolting machines (miner/bolters) give promising results, but a total system capable of achieving required performance is still being sought. 4. TECHNOLOGY PROGRESS The US coal mining industry has developed and adopted a number of technological changes in each stage of production, that have contributed to overall productivity improvements. Examples include identification of reserves, mine planning, development of mining methods, design of mining and materials handling equipment, automation and monitoring. Joy Mining Machinery is approaching the future challenges in a systematic fashion. Development efforts are focused in two areas: room and pillar, and longwall production systems. Joy is refining its individual products, so that their output and performance is synchronized and optimized. In room and pillar mining, the combination of a 12CM27 continuous miner with 4FCT0 haulage and a roof-bolting machine is an example of Joy’s system approach to technology. Joy 7LS shearer with Ultratrac 2000 haulage system, in conjunction with high capacity Armored Face Conveyor, and state of the art Roof Supports with RS20s electronics, is another example of the success of a “system” oriented approach. 4.1. Longwall Mining Mining companies are benefiting from technological advances in longwall mining. More powerful cutting motors and VFD shearer haulage drives, wider, longer and higher capacity conveyors, wider and heavier shields, improved automation and monitoring, all contributed greatly to a steady progress in production and productivity. US longwall production is growing steadily, while the number of longwall systems is declined from its peak of 130 longwall faces, to the current number of 52 longwalls. At the same time, longwall production increased dramatically. Longwall automation is increasing the reliability of equipment, and has a potential to remove people from a potentially hazardous environment. It is especially important in


28

gassy and dusty faces. Some mining companies have successfully employed partial automation, like shearer memory cut, shield advance, or conveyor chain tensioning, but full automation on a continuous basis has not been achieved yet. 4.1.1. Longwall Shearers All but 4 longwall shearers working in U.S. mines are JOY machines. The trend is towards increasing the power of the shearers due to the retirement of older machines and their replacement with newer and more powerful shearers; average installed power reached the level of 950 kW. Four shearers working in American mines have total power exceeding 1500 kW, and only 5 have installed power lesser than 750 kW. All shearers are high voltage, either 2300 V, or 4160 V machines. Joy’s shearer developments have resulted in increased cutting and haulage power, significantly increased cutting and flitting speeds, deeper webs, increased life between rebuilds, enhanced diagnostics and improved automation and horizon control features. More powerful motors necessitated design of stronger gearcases, with better gears, bearings and seals. More powerful shearer ranging arms allowed for wider webs. The most popular web size is 0.92 m, and some shearers are cutting 1.07 m web. Cutting speed of 20 m/min. is quite common, and some shearers are cutting even faster—up to 30 m/min.

Figure 7. JOY J525 Ranging Arm The latest, most significant development in longwall shearers includes AC variable frequency drive, a family of high power ranging arms and much more reliable electric motors, that allow higher cutting and flitting speeds due to increased power level in a given size. Joy developed the inverter and power supply for the VVVF system specifically for use in the harsh environment of Underground mining. The units are designed to withstand high vibration and shock loads as well as the ingress of moisture. The components are contained within the controller case and are mounted on a watercooled heat sink.


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Both, J525 (525 kW) and J750 (750 kW) ranging arms offer 30% more power than their predecessors, with only slight increase in their physical size. They are a good match for the new generation of AC haulage shearers, as the increased power capability of their haulage motors is now requiring more from the ranging arms. 4.1.2. Longwall Roof Supports Several trends in roof support design add to the ever-improving productivity and safety of U.S. coal mines. Some of these attributes include: – Increased automation. – Increased set and yield capacity. – Increased support width. – Rock burst protection. – Water sprays for dust suppression. – Increased structural life. – Decreased lower advance set (LAS) cycle time. All new roof support installations in the U.S. utilize full electrohydraulic controls. This prevents face operators from working in the elevated dust concentration downstream of the shearer. Electrohydraulic controls also improve productivity (shorter support LAS cycle time), allow for better face alignment, improve ground control thanks to uniformly high shield setting loads, and reduce the face manpower. The RS-20 electrohydraulic control system has proven to perform almost flawlessly, and has had several faces commissioned with “shearer initiation” operating from the face start-up date.s

Figure 8. JOY RS20s Roof Support Control System The JOY RS20 system will be gradually replaced by newly introduced JOY RS20s system. The new system is combining Joy’s roof support automation flexibility and features with a new generation of state of the art electronic control technology. RS20s system allows faster data communication and higher processing power in a slim package. It utilizes the latest advances in microprocessor technology that make it much more powerful than RS20. The system is designed to accommodate also future development in


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technology, not available yet, such as video and integrated voice communication. RS20s system comes with a library of over 100 different face cutting methods, and can be fitted to any shield using a JOY hydraulic system. The increased set and yield load capacity of modern shields is improving roof control, and is making deeper web shearing possible even in otherwise difficult conditions. Also, product quality is improving by reducing out-of-seam dilution from roof falls. Joy has produced the largest capacity roof supports in service in the U.S., with an impressive 1020T rating. Leg diameters up to 400 mm are currently being produced. 1.75 m wide shield became the industry standard. Wider shields have improved productivity, economics, and safety. Increased support spacing has also contributed to the need for increased support set and yield-load capability. The 1.75 m wide shields require fewer units to complete a given face length, but are only slightly slower to move during an individual LAS cycle, or from panel-to-panel on a face transfer, than the 1.5 m wide shields. The reduced number of units to complete a face with 1.75 m wide shields vs. 1.5 m wide shields leads to lower initial acquisition costs, as well as lower total operating and ownership costs. Even wider, 2.0 m shields are now available on the market. At this time, only one face in the world (in Australia) is equipped with 2.0 m wide shields, but undoubtedly, other companies will also acquire such shields in a near future. 4.1.3. Longwall Armored Face Conveyors (AFC’s) To increase overall productivity, Armored Face Conveyors (AFC) capacity had to increase. Numerous conveyors have 1.0 m, or even 1.15 m raceways and capacities up to 4500 tons per hour. Large installations of this type commonly have twin-inboard 42 mm chain and installed power up to 2550 kW (3×850 kW). Four American longwalls use twin 48 mm chain. With face lengths up to 335 m, soft starting and active chain tensioning management is required. In response to these needs, Joy introduced the TTT coupler. This dynamically controlled, water-fill coupler allows the motor to start under no load conditions and gradually build up torque transfer to the AFC chain. Not only are individual AFC drives automatically controlled for soft starting, but the 2 or 3 drives installed on the face are coordinated to control chain slack, and ensure smooth AFC startup, even under fully loaded conditions. In conjunction with the TTT, an automatically controlled tensionable tail frame actively adjusts the effective length of the AFC nearly in real time. Such controls are almost mandatory on very long faces with large conveying capacity, because the elongation of the AFC chains under load can create both, excessive tension and excessive slack during one normal cutting cycle on the face. In addition, longer faces are subject to more length change resulting from seam undulations, than shorter faces of past days. As a result of active chain tension management, some customers report improved AFC chain life by up to 40%, and reduced sprocket wear. Other JOY innovations, such as machined pan ends and trapezoidal line pan construction also helps to optimize the longwall mining system. The chainless shearer haulage system known as Ultratrac 2000, the enhancement of the rack based Ultratrac system previously offered by Joy, has improved both shearer and AFC performance. Fatigue life of the shearer rack sprocket has improved 400% while the tunnel area under


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the shearer was increased, and the ability of the AFC to follow an undulating seam expanded tremendously. 42 and 48 mm conveyor chains forced increase of conveyor pan height that is not desirable in lower seams. To alleviate that problem, Joy designed the BROADBAND low profile conveyor chain, which features flat, forged vertical links, and conventional roundwire horizontal links. Because its design, 42 mm BROADBAND low profile chain can fit pans using 34 mm traditional chains, and 50 mm flat link chain can be used in 42 mm pans. The prototype of the chain is being tested, and the chain will be available on the market in 2004.

Figure 9. BROADBAND low profile chain 4.2. Room and Pillar Mining Room and pillar mining has been practiced in U.S. mines for generations. Technology progress achieved in room and pillar mining almost matches progress in longwall systems. New generations of continuous miners, haulage equipment and roof bolters hardly resemble their predecessors. Still, lots have to be done to make room and pillar mining truly continuous, and match performance level of longwall faces. 4.2.1. Continuous Miners Continuous miners mine fifty percent of underground coal in the United States. The basic concept of a continuous miner has not changed since its introduction in 1948, and includes a multimotor design with electric tram, introduced long before it became available on longwall shearers. Between mid-1980s and mid-1990s longwall shearers made a successful switch from 1000 V to high voltage (3300 V or equivalent) power supply. Realized performance gained from high voltage longwall shearers encouraged designers of continuous miners to follow that lead. The decision to develop high voltage continuous miner was based on a number of key factors, like: performance potential, safety, maintenance and repair cost. Increased performance potential is based on possibility of installing more powerful motors on the machine, without detrimental voltage and torque drops. Since motor torque varies with the square of the voltage, any voltage drop has a drastic effect on motor performance. In addition to the increased torque and power, more stable voltage leads to reduced heating of cables, motors and electrical components, positively impacting maintenance and repair costs.


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Figure 10. JOY 12CM27 high voltage Continuous Miner Increased power of motors and related increased size of gearcases allowed for increased total weight of the machine that in turn improved its stability and reduced number of vibration related failures. Overall, high voltage machines offer increased tonnage between rebuilds, reduced operating cost, improved availability and increased output per shift. Despite their higher purchase price, cost per ton over the lifecycle of a high voltage machine is generally lower than for 1000 V continuous miners. The Joy Network Architecture (JNA) machine control system is based on Joy’s latest evolution of electronic hardware and software architecture. Key features of the system include enhanced machine diagnostics, electronic motor overload protection, surface data logging/communication and automation. Flooded bed dust scrubbers, stationary water sprays and ventilation methods were used for suppressing dust generated by continuous miners. Increasing output from continuous miners and related increased generation of dust forced development of alternative methods. Only recently, thanks to new generation of seals, installation of water sprays on continuous miner drums become a reality. 4.2.2. Haulage Equipment A haulage system is used to transport coal from continuous miner to a panel belt. Four haulage systems, split into two groups, are available: – Continuous Haulage: — Flexible Conveyor Train, — Chain Haulage System. – Batch Haulage: — Shuttle Car,


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— Articulated Hauler (battery or diesel). Selection of a haulage system depends mostly on mining system and conditions. A Shuttle Car, first introduced in 1938, is still the most popular machine for transporting coal between continuous miner and a panel belt. The latest development, installation of VFD drives on shuttle cars, brought a new life to that form of transportation. Articulated Haulers have a big advantage over Shuttle Cars—no trailing cable, that is the biggest reason for downtime on Shuttle Cars. Thanks to a cable-free design, machines are more flexible than Shuttle Cars. On the other hand, Diesel Haulers generate noise and pollution, and Battery Haulers use costly batteries that have to be charged every shift. Flexible Conveyor Train (FCT). In order to get the most from today’s high production continuous miners, a continuous haulage system must be just that, truly continuous. It was this type of thinking that led Joy to develop the 4FCT. This product is the latest offering in a long line of JOY haulage products that enhances the capability of the continuous miner by becoming an integral part of a matched system. The 4FCT can simultaneously convey material and tram to follow the continuous miner’s every move, while providing a continuous flow of material to the panel belt. The 4FCT offers a oneman operation, radio remote control, and a variety of features and options to maximize the productivity of the mine, and at the same time decreasing the downtime associated with maintenance and repairs. Whether for Longwall entry development or room and pillar extraction, the 4FCT is the haulage system of choice for high production applications. Both the traction system and belt drive utilizes VFD for a soft start. JNAII System adds valuable information on machine operation, troubleshooting and maintenance on both, the Traction and the Belt system. Chain Haulage Systems (CHS) are popular in low seam applications, where capacity of other haulage machines is limited. The biggest disadvantage of CHS is the fact that it requires 3 operators. 4.3. Variable Frequency Drives (VFD) Variable Frequency Drives are presently available on almost all of Joy’s mobile machines. VFD drives became standard on JOY longwall shearers, and are now the drive of choice on Joy’s new generation of continuous miners, continuous haulage products, and shuttle cars. The latest generation 1000 AC drive is a major step forward in development of VFD technology. Joy’s 1000 Volt AC drive uses the latest flux vector technology for state-ofthe-art control. It also features compact water-cooled construction, rugged and robust construction with extreme overload capability (400 V for 15 seconds), essential for mining applications. The drive is also fully regenerative, providing superior electrical braking compared to its AC and DC drive predecessors. Among numerous benefits realized with Joy 1000 V VFD are: – High available torque at higher speeds providing improved cutting and flitting speeds for JOY shearers, – Faster tramming and sump rates on continuous miners, – Superior performance on steep grades,


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– Improved heading control and accuracy on continuous miners due to the precise speed regulation. 4.4. JNA Control System A major electronic innovation for underground machine control and diagnostics is the Joy Network Architecture (JNA) system. JNA serves as a common hardware and software platform across all Joy products, offering improvements in both capability and reliability for the most effective coal production. The JNA Headgate System, for example, has evolved to add significant value to longwall mining operations. Benefits include reduced downtime while maximizing production, significantly reduced complexity by minimizing components and cabling, and the ability to interface with virtually any mine-wide monitoring or communications system for the highest levels of system integration. Flexible application options include either full longwall system operation or individual equipment, depending on operation requirements. In Joy’s longwall systems some key features of the JNA System include enhanced equipment diagnostics, face equipment control and integration, underground data logging, and a surface data communication link. Diagnostic information is displayed for all underground Joy equipment from a single location. Products including the shearer, roof supports, AFC, pumps, and programmable logic controllers can be uniformly controlled and monitored. All face equipment is also centrally controlled through features such as SIRSA, anti-collision logic, and external feedback. Operational information from each subsystem is stored for a defined period allowing playback of many mining cycle parameters underground. This capability permits a detailed analysis of long-term production and maintenance information. Finally, real-time operational information can also be assessed on the surface. With greater data storage above ground, longer time periods can be captured and analysed without affecting production. Memory Cut is becoming increasingly popular feature of longwall shearers. Memory Cut automates the cutting cycles through the creation of a cutting profile which the machine follows until cutting conditions change at which time a new profile is then created. The entire mining operation is programmed through a teaching procedure, with JNA basically memorizing the initial manual operation on subsequent cuts.


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Figure 11. Display screens for JNA system 5. CONCLUSION The American coal mining industry is facing extreme challenges, because of the competitive pressures from other energy resources, and impact of economy and environmental regulations. It is expected that the leading mining companies will continue to lead, and the leaders will absorb less productive firms, or force them to exit the business. The level of technology progress experienced in the last two decades, that made coal mining as efficient as it is now, is levelling off. Steadily declining price of coal limits resources available for research and development. No major revolutionary changes in technology are expected in a near future. Research and development will devote limited available resources to project with potential for significant changes in mining efficiency, and reduction of mining cost. Older, less efficient equipment will be systematically retired and replaced by new equipment, closing the gap in productivity between mines and companies. Despite those challenges, it is expected that coal production in the United States will grow on average by 1% per year for a foreseeable future. Underground coal production however will be steadily decreasing. Closer alliances will be formed between equipment manufacturers, coal producers and electricity generating companies. Only the fittest of the fittest companies will survive in the future, and for those, the future looks very bright. REFERENCES Department of Energy—Energy Information Administration Website. National Mining Association Website. Coal Age—February 2003. Joy Mining Machinery.

Prospects of Development of Highly Productive Coal Extraction Technologies

Volodymyr I.Bondarenko, Olexander M.Kuzmenko, Roman O.Dychkovsky National Mining University. Dnipropetrovsk, Ukraine International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 The increase of an overall performance of mines requires consecutive and economically justified complex change of coal extraction technologies according to possibilities given by the modern mechanized longwall sets. The important direction of research in order to increase the economic parameters to uncover the internal unused opportunities to influence the quantity and quality of extracted coal. Intensification of coal production is a prime task for the Ukrainian coal industry. The increase of mines’ capacity is possible to achieve in two ways: intensive and extensive. The extensive way of development means that the increase of coal production is based on the increase in the number of production faces. Such way demands an essential capital investment. So, extent of mining enterprise operational expenses for coal production is increased. Such development is economically not favourable, if there is only numerical increase of production faces without qualitative change of engineering and technologies of coal extraction. The intensive way of development provides introduction of technologically advanced mechanized systems and modification of mining technologies. The most essential technological parameters, which influence coal production, are: face advance, thickness of the coal seam, and length of the longwall and face mining limit. Changes of these parameters have influence on the mine’s technical situation. Therefore, decisions should be based on the analysis of actual geological conditions of mines. Taking into account complex structure of mining deposits, limitations of the mining field and mining width, it is necessary to approach the various production parameters on their own merit. Such approach guarantees adequate technological change of coal extraction being adopted for specific geological situation. As a result, the intensive way of development will decrease coal production cost, time of return on the expenditure on equipment, and increase of profitability of the coal mining enterprises. Its introduction is impossible without attracting new scientific development, new mining engineering and updating of some technological parts of mine.

Prospects of development of highly productive coal extraction technologies

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The technological flow chart of coal flow on a mine, offered by the authors (figure 1), enables to reveal “bottlenecks” at different stages of coal production and provides ways of their elimination {Bondarenko 2003). Technological flow chart of coal flow is divided into three levels: – production site, – main transport of a mine, – shaft and auxiliary technological complex. An analysis of the first level of the flow chart established, that increase of coal production is restrained by the following reasons: – long time required to equip and strip a longwall face and the discrepancy of this length comparing to the time of the whole technological process, – conditions of high stress, seismic activity of rock-mass and geological disturbances, – absence of modernisation of technological operations that would allow increasing coal production in difficult geological conditions and effective usage of modern mechanized longwall sets.

Figure 1. Technological flow chart of coal flow in a mine From these, the most important is modernisation of technology of coal mining and intensification of extraction work. It is achieved by increasing face length and rate of face advance. There is a necessity to revise the design of faces’ mining limits. The mining limit in favourable conditions should correspond to the time when technical deterioration


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of the mechanized longwall takes place before major overhauls are required. With the use of modern equipment it is possible to increase the distance of faces’ mining limits up to 2000–3000 m due to the technical characteristics of such equipment. Research of formation of lateral breeds in wall faces established, that increase in length of the longwall face from 150–200 m up to 250–300 m, practically does not change stress condition of the rock-mass and has no significant influence on the support of development headings. It is essential to prevent stoppages of production faces as they have negative impact on changes in rock-mass stress conditions. Technological parameters of mining operations should correspond to the parameters of rock-mass management. Now, the problem of face transport is solved by the domestic and foreign industry of mining mechanical engineering. A number of highly productive, reliable armoured conveyors with big motor potential that are compatible with new mechanized support are developed. It is necessary to stop a longwall face with the relaxation of rock-mass and lag of breeds on long distances. Current repairs of the elements of the mechanized system are carried out as required. Ventilation problem could be solved by gas drainage and making methane extraction a compulsory part of the technological process. The distance between degasification holes must be 20–40 m depending on actual geological conditions. The basic scientific and technical task of the first level involves utilization of modern mechanized systems aimed at increase of coal output and development of technological process to maximize use of the mining equipment. The second level of the technological process differs from the first due to its relative stability. Transportation of coal in mine workings is carried out with the use of belt conveyors over the whole distance. The main conveyor lines are loaded to 30–40% of their capacity. Increase of volumes of extraction does not cause necessity of re-equipment of mine transport systems. At this level the basic problem is reduction in quality of coal due to its mixing with waste rock if it is transported by means of the same transport lines. Therefore obligatory separation of coal and country rock transport systems in mine workings should be made. Technological process of mine transport should be aimed at full conveyerization of coal transport and trains could be used to move waste rock from the mine’s development ends. It is important that the staff adhere to such organization of work. The attitude of personnel employed in production workings to performance at all three levels of the technological process is one of the important problems of mining. Social transformations of the environment of mine workers directly depend on the volumes of coal being mined. When there is increase in extracted coal miners are seen in different perspective. Preservation of quality of the extracted coal and elimination of its losses during the transport in mine workings is the basic scientific and technical problem of the second level. Capacity of haulages was not found to be a deterrent to increasing coal production. Daily inspections of haulage equipment are compulsory under the safety regulations. They are carried out during a service shift and cause production and development faces to stop. Creation of storage capacities in shaft area for streamlining the process is proposed.


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At the National Mining University (NMU) a number of technologies directed at reduction of raw coal ash is developed. They are simple enough and easy to apply in the constrained conditions of mines. One of priority directions is dry processing of coal. It is done on small distances and can be introduced in any mining working with the area of cross-section larger than 13,7 m2. Actually, using this method of processing in underground workings enables to increase mining width by up to 50% above seam thickness by taking out waste rock, without loss of quality of production. Thus the problem of ventilation of faces is solved due to increase in cross-section of faces. Improving coal quality and streamlining the system of coal mining is the basic scientific and technical task of the third level. The common results of the analysis of all levels of the technological process and ways of realization of the proposed actions are shown in table 1. As an estimation of practical development an analysis of parameters of work of some mines of the State Holding Company “Pavlogradugol” has been carried out. The history of coal mining in this region is short and the mine officials offered to change the utilized there technologies of coal mining. For the purpose of the research the faces were chosen according to the program of modernization of the State Holding Company “Pavlogradugol” in 2003. These are the following faces: #560 (“Juvilejna” mine), #519 (“Pavlogradska” mine), ##169, 534 (“Samarskaya” mine), #1076 (“Dneprovska” mine), #531 (“Stashkova” mine), #506 (“Blagodatna” mine), and #345 (“Ternovska” mine). Three variants were chosen for comparison: the equipment required by mines, offered by manufacturers of mining equipment, new hardware necessary when increasing face length to 250 m. The analysis of the required equipment allowed arriving at a conclusion, that there is a certain stereotype of thinking present among mine workers that makes the mechanized systems seems very attractive. The preference is given to the technical equipment tested and used in mines. Innovations and changes in technical and technological maintenance of mining processes cause big fears. The following basic parameters of work of the mechanized systems were addressed: volumes of annual coal output, money value of the produced coal, difference between the income from coal sales and expenses for purchase of the equipment, and time of return on the capital spent on equipment (figures 2, 3, 4, 5). The obtained results have shown, that implementation of modern equipment in the State Holding Company “Pavlogradugol” enables to extract additional 1,3 million tons per year. The volume of coal sales increases by 127,3 million UAH/year, and the time of return on capital expenditure on modern mechanized systems is reduced by 15%. Change of technological parameters of extraction (face length increased to 250 m) combined with the purchase of modern equipment will increase the output by 2,2 mln t/year in the State Holding Company “Pavlogradugol”. The value of produced coal will increase by 220 mln UAH./year, with reduction in the time of return on capital by 2,1 months.


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Table 1. Results of the analysis of all levels of the technological circuit of coal movement in mine Level

1

The name

Longwall face

Direction of development

Increase of coal production

“Bottleneck”

Ways of realization

– length of the face, Scientific and technical development of technological circuits directed at – face advance, increase of length of production faces, rates face advance and – short mining reduction in a share of manual limits, labour. Introduction of modern – high stress, mining technical equipment. seismic activity of Application of preliminary gas rock-mass and extraction, extraction of roof, geological carrying out the creation of anomalies, workings of different cross-sections. – increased gas presence, – the organization of operation. – common coal and waste rock transport flow,

2

3

Main transport of a mine

Reduction – inappropriate of losses and improvement circuits of coal quality of transport, – the organization of operation.

Increase Shaft the capacity and auxiliary of coal storage and technological improvement complex of quality of final product

Development of technological circuits of transport with separate rock and coal flows.

– absence of storage Creation of additional storage capacity capacity underground for long-term underground, accumulation of extracted coal. Development and introduction of – absence of technologies of coal processing technologies of underground. processing and improvement of coal quality in underground conditions.

The obtained results allow decreasing the volume of mining works by 30%, which approximately corresponds to constructing 3 extraction sections with 180 m long faces. Following this it is possible to make a conclusion, that change of technological parameters of mining operations together with attracting modern equipment is the correct way of re-structuring of mines the State Holding Company “Pavlogradugol”.


Figure 2. Coal annual output from faces working in different technical and technological conditions

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Figure 3. Value of the produced coal for faces working in different technical and technological conditions


Figure 4. Difference between the income from coal sale and expenses on purchase of the equipment for faces working in different technical and technological conditions

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44

Figure 5. Time of return on the equipment for faces working in different technical and technological conditions


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CONCLUSIONS The main directions of technological and organizational improvement of technologies of coal mining providing concentration of coal mining are the following: – increasing face length from 180 m up to 250 m, – increasing face mining limit from 1200 m up to 2000–3000 m (before the first major overhaul of the mechanized sets), – increasing waste rock overcutting up to 40–50 cm, – leaving of waste rock underground, due to separation of transport systems and dry preliminary processing with total backfilling, – maintaining continuous coal extraction, – eliminating methane hazard without additional work on degassing by using rock overcutting and so increasing the cross-sectional sizes of faces, – streamlining the technology of coal mining with face shut offs for performance of service works combined with roof control, – creating storage capacities (5000–7000 t) for streamlining of coal mining. All that will allow achieving: – daily output of commodity coal of 2500–3200 t/day from one production face and 7500–10000 t/day from a mine, – good ventilation that meets the requirements of Safety Regulations, without additional work on degassing, – 1,5–2 times increase in labour productivity in a mine, – improved working conditions of miners in production sections.

REFERENCES Bondarenko V.I., Sally V.I., Kuzmenko A.M., Porotnikov V.V., Dychkovskij R.E., Kotov Y.V., Sytnyk V.V.: The Ways of Re-Structuring of Mines of the State Holding Company “Pavlogradugol” at Investment of Production. Proceedings of NMU of Ukraine, Dnepropetrovsk: EIC of NMU of Ukraine, 2003. № 17. Vol. 1. PP. 140–149.

The Effectiveness of Support of Weak Rock with TA2 Tubular Anchors

Volodymyr I.Bondarenko, Gennadiy A.Simanovich, Irina A.Kovalevska National Mining University. Dnipropetrovsk, Ukraine Vjacheslav V.Porotnikov State Holding Company “Pavlogradugol”. Ukraine International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 One of the perspective types of anchor support are tubular anchors, installed with the help of energy obtained by detonating an explosive charge. They are relatively cheap to manufacture and install and have good support characteristics. However, the basic difficulty restraining wide implementation of tubular anchors lies in the fact that the process of interaction of tubular anchors with rock surface in the borehole is not studied enough and is not fully described by mathematical equations and hence proper methods of calculating basic parameters of support are not available. So, studying the process of interaction of forces in the system “tubular anchor-rock”, creating on this basis an engineering method of calculating tubular anchor parameters and improving construction and technology of their installation are the actual tasks for theoreticians and practitioners constructing underground workings with the aim of increasing their stability. One of the basic criteria of anchor support performance is the value of its holding force, which is determined by the smaller of two values: force required to pull the anchor out from a borehole and tensile strength of the unit (bar, pipe and so on). From the point of view of providing maximum support resistance with minimum material consumption per support unit, anchor’s rational construction would be such that provides condition of equal strength—anchor’s holding force would be the same according to both abovementioned criteria. As it was shown by analytical and experimental research while expanding tubular anchor pipe by detonating an explosive charge that adhesion of anchor’s material to rockwall of the support hole is not observed and its supporting strength is determined by friction force along contact surface, resulting from radial stress on “anchor-rock” contact. Naturally, with other conditions the same, the greater the value of radial stress, the bigger the value of the friction force (which is a product of radial stress and metal-on-rock coefficient of friction). Hence, value of anchor holding force in hole changes according to cross-sectional area on condition that construction of uniform strength is selected.

The effectiveness of support of weak rock with TA2 tubular anchors

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On the other hand, the value of anchor’s holding force depends on the area of distribution of friction force in a borehole (determined by radial stress distribution) and the greater the area is, the bigger the value of the anchor’s holding force, while value of radial stress stays constant. “Anchor-rock” system interaction mechanism research allowed determining the relation of the abovementioned parameters to strength and deformation properties of system elements. With time the value of radial stress reduces under the action of rheological factors and different structural planes of weakness (Manual…1983), (Bokij 1972) from maximum value at the time of anchor’s collision with rockwalls of hole to minimum constant value when “anchor-rock” system becomes a balanced one. It was determined that the minimum (set constant with time) value of radial stress on the “anchor-rock” contact surface corresponds to the value of the rock’s uniaxial compression strength, taking into account main rock strength reducing factors—structural weakening and influence of humidity. That is, with time a condition is reached when radial stresses are equal to the calculated uniaxial compression strength of rockwalls of the borehole. That’s why, the stronger the rock in which tubular anchor is installed, the higher the value of its holding force. Surface of “anchor-rock” contact on which radial and tangential stresses are distributed depends on correlation between strength and deformation properties of anchor’s material and rockwalls of borehole. The basic underlying reason for this dependence is that anchor compresses and “gets loose” from rockwalls of borehole under the action of axial load. This leads to radial stress decreasing or disappearing on “anchorrock” contact and, as a result, anchor’s holding force reduces. The greater is the difference between the values of strength of anchor’s material and the rock of borehole’s walls (determining radial stress value), the more intense the process of anchor’s radial compression under the axial load is going to be with the resultant diminishing of the contact surface. So, the disadvantage of this method of installation of tubular anchors TA1 is reduction in their holding force if the are installed in weak rocks as a result of unsatisfactory regime of anchor’s interaction on contact with rockwalls of borehole due to the distinct difference (10–30 times) of strength and deformation characteristics of anchor’s material (steel) and rock. This leads either to reduction in anchor’s holding force (if it is determined by the value of pull-out force) as the result of its cross-sectional compression under axial load, or to partial use of anchor’s material supporting capacity (to material overdesign) when support design was done according to the criteria of anchor’s failure from loads due to mining-induced stress. Using TA1 tubular anchors characterized by low values of holding force to support underground workings, especially in weak rock, where significant support load is observed, leads to the necessity of using dense pattern of rock reinforcement, that is to a higher density of anchor’s installation, that leads to the increase in volume of work for drilling and anchor installation. There are two ways to eliminate disadvantages caused by low supporting strength of tubular anchors TA1 in weak rocks. The first one is to strengthen rockwalls of boreholes during installation; the second is to use more completely the strength properties of boreholes’ rockwalls along “anchor-rock” contact surface by improving the regime of anchor’s interaction with BOREhole’s rockwalls on contact. It covers the limitation of


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the “anchor-rock” contact process weakening while anchor’s cross narrowing by means of given value pressure created on its inside surface. The first method has important disadvantages, the essence of which is the following. Existing tubular anchors’ constructions (Sentsov 1981, 1982) foresee strengthening of the external rock layer with help of hardening solutions, and sand-cement mixture is usually used for that purpose. When an explosive charge is detonated, the mixture is injected into the cracks and holes in rock under the pressure of expanding blast products or under the pressure created by deforming pipe wall. It is supposed that a zone of increased strength is created on the outside edge of rock and increased anchor’s supporting strength is achieved. The following factors should be underlined as an opposition to this opinion. First, injecting hardening solution into borehole’s rockwalls, while blasting an explosive charge, is a fast process, taking a very short time under speed of 100 and more meters per second. The importance of viscosity of the cementation mixture, which is rather high when it is injected into holes and cracks under static conditions, increases even more under blast-forced injecting. That is, the resistance to injecting significantly increases and, therefore, the depth of outer rock layer strengthening decreases proportionally to it. Not depth strengthening but spreading of the mixture over the surface of borehole’s walls takes place. Second, the strength properties of external rock zone can be increased only by such substance, which has higher strength than that of the rock. However, due to the abovementioned reason the strengthened rock layer has insignificant thickness and when radial stress from the anchor’s side act, this layer plays a role of a load transmitting element because of its insignificant ability to withstand the stress. That is, the greatest part of radial stress is transmitted through the strengthened outer layer to unstrengthened rock, which has much lower compression strength. So, unstrengthened rock again becomes the main element determining strength properties of borehole’s rockwalls and the effect of its low strength on anchor’s supporting capability is not removed. Third, the mixture for any significant outer rock layer strengthening must be rather flexible (in order to reduce its viscosity) that is to have high water-cement ratio, which means sharp reduction of concrete’s strength properties (Skramtaev 1953). In practice such a concrete’s compression strength doesn’t exceed 3–5 MPa. It determines low efficiency of borehole’s rockwall strengthening and, therefore, low supporting strength of anchors installed in weak rock with this method. That’s why in the case when the rock strength is much lower than that of concrete, the introduction of intermediate, weaker layer on the contact between anchor and rock leads to significant loss of its supporting capability. But if the intermediate layer made of hardening material reduces anchor’s supporting capability as far as its supporting strength in borehole is concerned, the holes made in anchor (Sentsov 1982) in order to facilitate injecting cement-sand solution into the rockwall, on the contrary 1) reduce anchor’s cross section, 2) help to create stress concentrations around them and while anchor is loaded with tensile forces. But, on the other hand, they also reduce anchor’s supporting capability concerning its material strength.

The effectiveness of support of weak rock with TA2 tubular anchors

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Figure 1. The construction of tubular anchor before detonating the charge (a) and tubular anchor TA2 after complete installation (b): 1—borehole, 2—steel pipe, 3—clay plug, 4—detonating cord DCA or DCB, 5—electrical detonator, 6—electrical wires, 7—filling material (sand, water), 9—material expanding while solidifying That’s why another method is the most expedient. Its aim is to improve “anchor-rock” system force interaction regime by means of installing into the anchor’s inside cavity material with such deformation properties that would ensure providing necessary forces acting on the inside surface of anchor. An anchor is expanded in a borehole by detonating an explosive charge with the aim to increase its support capacity in weak rock by improving the regime of its interaction with rock. After that the anchor’s inside cavity is filled with material expanding while hardening (figure 1b). Volumetric expansion coefficient of the material is calculated according to the formula:


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where µ, µΠ, µM—accordingly Poisson ratio of steel, rock and solidified material, E, Eп, EM—accordingly elasticity modules of steel, rock and solidified material, d1 and— accordingly anchor’s inside diameter and borehole’s diameter, β—a coefficient which takes into account anchor’s material plasticity limit increasing during its expanding by means of energy of the detonation, Rcж—the uniaxial compression strength of rock. Increasing of TA2 tubular anchors’ holding force in weak rocks is achieved by preventing the possibility of anchor’s cross-sectional compression under axial load. This is done by filling its inside cavity with expanding solidifying material after installation. And, therefore, the possibility that the radial stresses acting along the “anchor-rock” contact surface will decrease is eliminated. As a result, the value of anchor’s supporting capability in borehole stays constant in time and equal to the calculated value. It provides uniform strength of the construction concerning both the anchor’s material failure and its holding force in borehole under increased stress conditions. REFERENCES The Directory on Designing of Underground Mining and Calculation of Support Parameters. M.: Stroyizdat, 1983. P. 273. Bokij B.V., Zimina E.A., Smirnjakov A.V.: Technology and Complex Mechanization of Mining. M.: Nedra, 1972. P. 81. Sentsov P.I., Sagalayev Ju.I., Chursin B.N. & Andreev A.V.: A Way of a Tubular Anchor Installation. A.S. 796448 (USSR). Institute KuzNII, 22.03.79, № 2740224/22–03. Publ. in B.I., 1981, № 2, MKIE 21D 20/00. Sentsov P.I. & Korschun A.P.: A Way of Hardening of a Rocky Massif Around Mining Workings. A.S. 969903 (USSR). Institute VostNII, 01.04.81, № 3267529/22–3. Publ. in B.I., 1982, № 40, MKIE 21D 20/00. Skramtaev B.G. et al.: Building Materials. M.: State Publishing House. 1953. P. 643.

Equipment Selection for Mechanized Mining as a Function of Physical, Mechanical and Deformation Properties of Rock

Milenko Ljubojev Copper Institute. Bor, Serbia & Montenegra Mladen Stjepanovic University of Belgrade. Serbia & Montenegra Mirko Ivkovic EPS JP for PEU. Resavica, Serbia & Montenegra Savo Perendic Director of “Lubnica” Coal Mine. Serbia & Montenegra International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: Physical, mechanical and deformation properties of a coal seam and accompanying rocks have an exceptional importance in selection of mining method and technology, in selection of mining equipment; the way of equipping the face and construction of underground mining facilities. This paper presents the results of research of the influence rock-mass characteristics has on selection of equipment for mechanized mining of the floor part of the first coal seam, in specific case of “Lubnica” mine coal deposit. KEYWORDS: Geomechanical parameters, coal, mechanized mining equipment selection

INTRODUCTION Basic directions in research of technical and technological solutions for all stages of PPS technological complex, in modern mining practice and science aim towards complex mechanization and automatization of all processes and operations at all stages in underground mining of coal deposits.


52

In PPS system as a whole, the basic and starting problem is the face, as the smallest and basic production unit, therefore the problem of research of rational choice of mining methods and technology is a prerequisite for optimization of PPS technical and technological parameters. Introduction of mechanized mining technology utilizing complex mechanization with self-advancing hydraulic support (figure 1) has been the basic direction in coal production in underground mine for several years. Self-advancing hydraulic support is about 20 times more expensive than the previously used at longwall faces steel support. It is very complex movable machinery with mechanical, hydraulic, electric and electronic elements, which in permanent operation has to withstand the load of the overlying rock, ensure safe working environment at the face and allow achieving rapid mining rates. Higher price of the selfadvancing hydraulic support and of the face equipment as a whole has to be compensated with increased coal production and much higher labour productivity.

Figure 1. Mechanized longwall equipment complex disposition. 1— hydraulic support, 2—direction conveyer, 3—excavating machine, 4— stage loader, 5—crusher, 6—electrical plant, 7—driving station, 8—driving plant, 9—hydraulic pump, 10—winch There are three variations of mining longwall systems today: – mining with increased face length (adit and surface coal winning according to the principle of vertical concentration) utilizing mechanized longwall sets, – caving of the roof in the system of horizontal concentration, with mechanized longwall sets, – panelling in the system of horizontal concentration, with mechanized longwall sets. Physical characteristics of a deposit are the basis for selection of technical solutions. Even so the organization of duties and work for all technological operations and

Equipment selection for mechanized mining

53

production stages within the whole longwall require constant adjustment to changing working environment conditions. Physical and geological conditions of a deposit have dominant influence and are most important in selection of technical and technological variants and solutions. This influence shows itself in technological process through technical and economical production indicators. Self-advancing hydraulic support is designed as a retaining, shield and retaining-shield support. Retaining support is intended to support the longwall working space over a relatively large span of roof (more than 4 m). Supports of this type usually consist of 2 or 3 rows of props in one support section and accept the load from immediate and main roof. Shield-type supports are intended to support working space over a small span of immediate roof, from 2 to 3, 5 m and they as a rule, consist of one row of props in one support section. Shield supports accept the load coming mostly from the immediate roof. Retaining-shield supports are intended to cover longwall face working space over a roof span of 3–4 m, and they usually consist of two rows of props in one support section. There is obvious production decline, lagging in technological development and high losses in business operations in “Lubnica” brown-coal and lignite mine, and therefore justification of further mining is questionable. Investigations to the possibilities of utilizing mechanized mining were carried out in order to overcome such complex problems. The investigation results have shown that mechanized mining can be introduced in specific conditions with the use of shorter mechanized faces. This is a significant advance in the sense of production, economics and safety, compared with currently used room-and-pillar methods. INVESTIGATION OF WORKING ENVIRONMENT CONDITIONS In selection of a rational system of mining in our mines the current conditions of mining have predominant influence. Owing to significant difference in geological age and intensive tectonics of deposits (The Tertiary coal deposits instead of carboniferous deposits), the mining conditions are different from conditions in other mines in the world from which we are trying to transfer mechanised mining technology. It should be emphasized that the mining conditions in our deposits are relatively difficult, with significant changes from one deposit to another and frequently between different mining areas within the same deposit. These changes are particularly pronounced as mining operations advance deeper. Since physical and geological conditions cannot be influenced directly, technical and technological conditions of mining process represent the subject of research and studies. Choice of an adequate mining method and technology, design of the best ore winning and transport system, selection of the type of roof support and work organization, present a number of potential possibilities for improvement in effectiveness of mining operations. Carbonaceous area of the “Lubnica” mine is a part of Lubnica—Zvezdan tertiary carbonaceous basin which encompasses the area of 14 km2 and which Lubnica river and upper cretaceous bar of meridian direction divide into three parts (Lubnica, Zvezdan and Northwestern part).


54

Coal mining is currently being performed in “Centralno Polje” of the “Stara Jama” pit. In this area, the roof cut of the first coal seam was mined out in the 50’s and 60’s and since 1992 mining of the floor cut of this coal seam is under way. The neogene basin is located in depression of vulcanogenic and upper cretaceous sedimentary strata which form the edge of the basin. Volcano-clastics of andesite magma are predominant in the structure of these rocks, whereas sandstone and marl are more are represented by basal conglomerate and sandstones rare. Sediments of helvetian and by clay marl rocks on higher levels. Carbonaceous horizon is basically made of marl rocks with partially present thin seams of argillites and sandstones. Upper parts of the helvetian carbonaceous series are of clay-sandstone composition dominated by unstratified clays, sandy clays and clay sandstones alternating in most cases. Sediment thickness in the area of “Stara Jama” is various and depends on location. The whole area represents a depression where coal seams (the first and the second) were formed in a syncline. Thickness of the sediments of the helvetian age is about 100 m. The first coal seam is at the distance of 30–40 m in direction of stratygraphy and has productive thickness of 10 m. A cross-section of the seam structure is shown in figure 2. Thickness of sediments formed during the Tortonian—Sarmatian stage is different and depends on sediment and tectonic-erosive processes and ranges from several tens of meters (at the edge of depression) to 100 m and more (in central parts of depression). Lubnica—Zvezdan basin is divided into a large number of larger and smaller units by intensive radial faults of different strikes. The faults of meridian or approximately meridian strike belong to the oldest generation of faults, which influenced the structure forming. Faults of the second generation, striking predominately in east-west direction, belong to the second tectonic stage, which extended throughout all the Tertiary.

Figure 2. Geological structure of the first coal seam. 1—roof marls tuff sandstone and tuff microconglomerate, 2—excavated roof part of the coal seam, 3—roof part of the coal seam functioning as a protective slate, 4—


55

main thin seam of the tuff sandstone, 5—unexcavated floor part of the coal seam, 6—laminated lime sandy aleurite floor sandstones The third and the youngest generation of faults of different throws had a significant impact on the formation of the structure of the deposit, and the impact was greatest on elevation of the seam. Each generation of faults caused changes in the dip and strike of the strata, frequently over short distances. Tectonic movements are the cause of the dip variations in different parts of the deposit. Dips in the area of “Stara Jama” range from 16 to 25°. No problems with ground water were encountered during prospecting and mining in the area up till now. According to volume of water inflow, the deposit was classified as not affected by ground water problems. The coal from Lubnica—Zvezdan basin belongs to the group of xylitic humic coals, which the degree of their carbonification places between regular brown lignites and soft lignites. The coal is of brown-dark colour with shades depending on impurities. The coal crumbles to pieces and under atmospheric influence decays and crumbles. Estimated coal reserves in the area of “Stara Jama” and “Osojno” pits are approximately 12 million tons in categories A+B+C1. Methane presence was observed during mining operations and because of that all work is performed in “methane duty”. Coal propensity to spontaneous ignition was determined by performed investigation and confirmed by frequent mine fires of endogenous characters. It was established by laboratory investigation that under certain conditions coal dust shows flammable and explosive properties. Samples for investigation of physical and mechanical properties were taken in the shape of blocks from the coal seam in advanced galleries IH-1 and IH-2, which were advanced to form the first mining field in the floor cut of the first coal seam in the Northwest part of the pit. Development headings are driven on strike and are located in the floor. The immediate roof is formed by a seam of fine-grained 0,5–0,8 m thick tuff sandstone and a 0,5–0,8 m thick coal layer situated above it. Above that is goaf of the previously excavated roof cut of the same coal seam. Light grey-greenish sands and marl sandstones form the immediate floor of the coal seam. All investigations of physical and mechanical properties were carried out in the laboratory for soil and rock mechanics in Copper Institute Bor on small samples, using methods that are standardized for each kind of mining investigation, according to recommendations of the International Society for Rock Mechanics (ISRM). All the necessary investigations were carried out on such number of samples that based on the obtained results, an average representative value of an investigated property was determined for each investigated area. Average values of the investigated properties were calculated using formulae with standard deviation and medium datum variation coefficient. Tables below show the investigation results of the properties of the coal and accompanying rocks.


56

Table 1. Physical properties of the floor cut of “Lubnica” coal seam Lithologic unit symbol Volumetric weight Water content Longitudinal Transverse (KN/m3) wave (%) V wave velocity velocity B1 (m/s) B1 (m/s) 15,04

20,22

Kvar=2,53%

Kvar=8,45%

12,45

36,13

Kvar=1,00%

Kvar=6,44%

16,62

21,62

Kvar=CM 5×8

Kvar=8,62%

16,87

26,60

Kvar=2,53%

Kvar=11,46%

Immediate roof

Coal seam

Floor board

Immediate floor

1698 II

1898 II

1685 II

1815 II

Table 2. Mechanical properties of the floor cut of “Lubnica” coal seam Lithological unit symbol Immediate roof Coal seam

Floor board

Immediate floor

Unaxial Tensile Shear strength σ1 compressive strength σx (MPa) strength (MPa) (MPa) 8,11

0.79

3,18

Kvar=15,12%

Kvar=11,73%

Kvar=16,65%

1948

2,01

4,65

Kvar=8.99%

Kvar=7,29%

Kvar=15,81%

7.21

0,69

4,86

Kvar=7,5%

Kvar=8,35

Kvar=6,23%

9,32

0,87

4,98

Kvar=8, 53

Kvar=13,10%

Kvar=18,83%

Cohesion (MPa)

Internal friction angle φ°

1,25

38

2,90

35

Kvar= 10,24% Kvar= 7,10% 1,23

33

1,78

36

Table 3. Deformation and technical properties of the floor cut of “Lubnica” coal seam Lithologic Tangent unit modulus symbol E1 (MPa) Immediate

480

Secant modulus Es (MPa)

Deformation modulus at failure E1d (MPa)

Poisson Ratio V

240

240

0,21

Resistance 2

(N/cm)

(N/cm )

383

58

Coefficient of crushability f


roof

57

Kvar=8,84% Kvar=13,78%

Coal seam

614

325

Kvar=8,39% Kvar=8,55%

325 Kvar=7,29%

0,33

1536

211

Kvar=10,22% Kvar=12,25% Kvar=15,33%

Floor board

519

330

318

0, 36

65 510 Kvar=13,62% Kvar=14,68%

Immediate floor

535

330

319

0,27

99 600 Kvar=21,37% Kvar=8,66/0

7,99 Kvar=15,33%

MODELLING OF STRESS-DEFORMATION CONDITIONS IN THE ROCK-MASS AROUND MECHANIZED WORKINGS The main problem in determining the choice of equipment for mechanized workings is selection of support elements that is, SHP section. There are several approaches to solving the problem of correct selection of necessary SHP load-carrying capacity for specific conditions. One of the approaches is to select SHP of such capacity that would prevent excessive convergence. Another approach is to select support that would be able to provide resistance high enough to completely prevent separation of roof strata. Limiting factor in the latter approach is stress on the floor and roof, which must not exceed strength of the immediate floor and roof rock. Geomechanical considerations during SHP design process are related to selection of magnitude of reaction, which the support exerts on the roof, and also with selection of its load during relaxation. When the values are set too low excessive convergence may take place. Values of support resistance that are too high may on the other hand cause failure of roof and floor rock. Complexity of roof control in mechanized longwall faces demonstrates itself, among other things, in the necessity to achieve stability of the roof beam in front of the support, and in its good breakability behind the support. In most cases, the strongest rock in the productive series is coal, therefore the problem amounts to establishing the thickness of the roof beam as a function of set requirements. In specific case of the first seam of “Stara Jama” pit the problem became even more complex due to the fact that the upper cut of the seam was mined out. The sill pillar that separates goaf from the floor cut by the seam, which is to be mined now, is formed by 0,8 m thick layer of coal that was left in foot during primary mining operations and 0,4 m thick tuff sandstone below it. The floor cut of the seam, which is to be taken out in secondary mining operation is 2,8–3,5 m thick. Since the thin roof beam was found not to be sufficiently strong, the aim of the research was to determine the thickness of coal layer that would need to be left in the roof in order to provide stability of excavation in front of the support. The problem was solved with the use of finite element numeric models. Phase 2 software was utilized (2D finite element program for calculating stresses and estimating support around underground excavation), Rock Engineering Group, University of Toronto (Ljubojev et al. 2000). The basic model of finite elements on which the research was carried out is shown in figure 3.


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Figure 3. Look of the finite element basic model During research this model was transformed in about ten separate models and mining phases were simulated in each. A 0,3 m thick coal layer was modelled. It was assumed that a 0,3 m thick layer of coal would be left in floor during mining, and thickness of coal layer left in roof was variable (0,3, 0,5, 0,8 m). Depth of cut was 1,0 m, and maximum unsupported span in front of the support was 2,0 m. Coal disintegration by the cutting machine was modelled. The finite element method was used with numeric method modelling, and elasto-plastic model for rock-mass with Mohr-Coulomb material. The first model was run with 0,3 m thick coal layer in roof, the input data based on actual engineering experience and preliminary model research. After a certain number of cuts, after fracture of the roof beam behind the support, progressive fracture and failure of rock-mass occurred. Soon after that the roof beam in front of the support was gradually breaking. The seam of sandstone was broken first, then the 0,5 m and 0,8 m thick coal beam is broken. At this stage the complete simulation of mining is carried out on several consecutive models (figures 4–7).

Figure 4. Graphic illustration of 0,3 m thick roof plate stability condition


59

The figures show isolines of safety factors, with two different types of markers used. Dots mean elements broken due to extension stress, and crosses mean elements breaking in shear. Plastification zone was above the support in goaf as shown, whereas both coal layers, even sandstone show satisfactory stability.

Figure 5. Graphic illustration of 0,5 m thick roof plate stability condition

Figure 6. 0,8 m thick roof plate stability condition isolines


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Figure 7. 0,8 m thick roof plate Plastification zones According to the calculated stress values in front of the face, maximum stress concentration was 10,65 MPa, and virgin stress condition with the value of 2,75 MPa existed 10 m ahead of face in the solid. Up to 1,5 m ahead of the face, the sandstone seam which is located in the roof between the two layers of coal, will be crushed, therefore the maximum bearing stress was at 1,5 m in front of the face with value of about 8 MPa. The calculated required support resistance of the frame type support is F=1,37 MPa, whereas for shield support, depending on the shield type, it amounted to 1,5–3,0 MPa on the floor and 1, 5–2, 5 MPa on the roof. Application of two-prop retaining-shield support is adequate according to physical, mechanical, deformation and technical properties of the coal seam. Coal in the investigated floor cut of the first coal seam belongs to hard brown-lignite coal with compressive strength of σpsr=19,48 MPa and modulus of deformability Et=614,0 MPa. Required cutting force was established to be KI=1536 N/cm2 and cutting resistance KF=211 N/cm2, therefore the coal can be excavated easily with a loader. Stress and deformation analysis using finite element method with 0, 3, 0, 5 and 0,8 m thick coal layers in roof, indicated that the 0, 3 m one is unstable. CONCLUSION On the basis of the performed environment condition investigations, it was established that mechanized mining with the use of SHP as support of the working space at the face and excavating machine with cutting technology, can be applied, in the specific case of the floor cut of the No.1 coal seam in “Stara Jama”—“Lubnica”. REFERENCES Ivkovic M: Rational Systems of Underground Excavation of Brown Coal Seams under Complex Mining Conditions. Doctoral Dissertation. School of Mining and Geology, Belgrade, 1997.


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Ivkovic M.: Methodology of SHP Selection for Excavation of Brown Coal Seams Using Longwall Method According to Vertical Concentration Principle. V Symposium on Mining Mechanization. Proceedings. Pages 157–164. School of Mining and Geology, Belgrade, 1999. Ljubojev M. et al.: Study on Preliminary Geomechanical Working Environment Research of SZ Part of “Stara Jama” of the Floor Part of the First Coal Seam of “Lubnica” Mine. Copper Institute, Bor, 2000. Stjepanovic M.: Scientific Base of Determination of the Laminal Deposites Underground Exploitation Main Parameters Determination. Monography. TF Bor, 1993.

The Influence of Physical Parameters on Mechanizability of Longwall Mining of Coal

K.Oraee & R.Pourkhandani University of Tarbiat Modarres. Tehran, Iran International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: Coal has always been considered to be an important source of energy and despite short-term fluctuations, its long-term total demand in the world, shows an upward trend. Longwalling remains the most attractive method of winning coal, mainly due to the availability of efficient machinery. In order to produce coal economically, optimal adoption and efficient utilization of mechanization is of utmost importance. In this paper, the importance of coal is first briefly discussed and it is then argued that the most important parameters influencing the viability of utilization of mechanized mining (or in short “mechanizability”) of coal seams are: roof and floor rock quality, seam thickness and dip. A distinct mathematical model is introduced that quantifies the effect of each parameter on mechanizability of coal seams. Combination of these component models by appropriate methods produces a comprehensive model that measures the mechanizability of coal seams in longwall mining. These component models could serve as an important tool for the mine design engineer in deciding the level of mechanization to be adopted and hence the most appropriate machinery to be utilized, in order to attain economy of production.

INTRODUCTION Coal has always been one of the most important sources of energy and despite short-term fluctuations, its total production and consumption in the world shows an increasing trend. Amongst others, longwalling is the most widely used method in deep coal mines. It has


64

been the predominant method in Europe and it has become more popular in most other parts of the world in the post war era. The reason for increased popularity in the use of longwall method can be sought in the fact that economic forces, coupled with fierce competition amongst coal-producing countries, have necessitated increases in productivity. This could only have been achieved by increasing the size of mining entities, in order to increase the volume of output obtained from each production unit. Such objective has proved to be attainable only when intense use of mechanization is made. Coal faces with low degree of mechanizability have therefore disappeared from the list of producing faces and small mines have combined to provide large mining complexes. The economy of scale achieved in this way, has resulted in higher productivity all over the world. LONGWALL MINING Longwalling is known to have the highest level of mechanizability amongst all other methods. It is highly flexible under varying geological conditions and with the modern machinery now available, it is capable of producing sufficiently high levels of ore to justify deployment of expensive equipment. It is for this reason that the exact mechanizability of a coal seam becomes an important characteristic, since any interruption in production lowers the coefficient of utilization of expensive machines and consequently results in significant amounts of financial lasses. Whilst there are numerous criteria affecting the mechanizability level of a typical coal seam, the most important ores are of the interest of this paper. Roof rock quality, on the other hand, hinders the advance of powered supports and hence decreases the volume of production expected from a face that is equipped with expensive machinery. The effect of seam thickness is somewhat different, in that, the possibility of using powered support and shearers in virtually any seam thickness exists, but, from economical point of view, it is only feasible to use power support in seams that are neither thinner nor thicker than a certain value. The angle of dip in the coal seam is also an important parameter, although again, the nature of its effect is different from others. It is argued that conventional machines can only be used on steep seams if ancillary equipment is resorted to. This assumption limits the scope of this study to moderate dip angles since steep seams are incomparably more expensive to equip than moderately dipping ones. ROOF ROCK QUALITY An important advantage of longwall method, as compared to room-and-pillar, is that it can be used in weak strata. Modern powered supports can virtually cover the whole area of the roof and hence provide a safe working environment, both for men and equipment. Although the method, with the use of advanced powered support, can be applied where the roof rock is weak, moderate or strong, the degree of success varies in different situations.

The influence of physical parameters on mechanizability

65

If for example, the roof is excessively weak, then unexpected minor falls of rock will hinder smooth and continuous advance of the powered support. On the other hand, if the roof rock is hard and stable, then after the advance of the powered support, the roof may not cave behind the face line. In such situations, where the caving process does not occur systematically and continuously, excessive load is exerted on the roof of the working face, which again limits the efficient operation of the mechanized face. The best conditions therefore, for mechanization to be applied efficiently, are where the roof is moderately weak to moderately strong. Coal seam mechanizability therefore depends on the quality of the roof. Amongst some possibilities available, the approach adopted by Unrug and Szwilski (Unrug, Szwilski 1982) seems to fit this situation best. They introduced a model that produces a classification for the roof rock quality according to the in-situ compressive strength of the rock and thickness of the immediate roof (Equation 1). Qr=0,016σ·d (1) where Qr—the roof rock quality index, Kg/cm, σ—the in-situ compressive strength of the immediate roof, Kg/cm2, d—the thickness of the immediate roof, cm. Since measurement of a can be an expensive process, it has been suggested that it is calculated using equation (2). σ=σC·K1·K2·K3 (2) where σc—uniaxial compressive strength of the immediate roof rock, Kg/cm2, K1— coefficient of in-situ strength (table 1), K2—coefficient of creep (table 1), K3—coefficient of moisture (table 1). Indicative values for all these coefficients are given in table 1.

Table 1. Coefficient values for different type of rock Claystone & Siltstone

Mudstone

Sandstone

0,5

0,42

0,33

K1

0,6

0,65

0,7

K2

0,4

0,5

0,6

K3

Thickness of the immediate roof (d) can also be calculated using equation (3). (3) where t—thickness of the coal seam, m, K—swelling factor of the roof rock. In this way, the quality of the roof rock can be quantified and hence its effect on the mechanizability of a particular coal seam determined. Table 2 shows approximate values obtained from such calculations.


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Table 2. Classification of immediate roof in terms of quality index (Unrug, Szwilski 1982) Roof rock description

Qr (Kg/cm)

Immediate fall of roof when exposed

0–18

weak

Difficult to control

18–35

fairly stable

Caves readily

35–60

semi-stable

Good caving characteristics

60–130

stable

Requires system to initiate caving

>130

very firm

Not suitable for caving

>250

unsuitable for caving

FLOOR ROCK QUALITY The quality of the floor rock is also another important factor in mechanizability of coal seams. In order to utilize machinery at the face, the floor rock must be strong enough to withstand the load exerted on it through the powered support units. If the floor is not sufficiently strong, the support legs punch into the floor. Research into the bearing capacity of the floor rock was first carried out by Prandtel (Prandtel 1921). According to his findings, due to plastic behavior of weak floor rocks that are typically composed of clay minerals, for purposes of analytical studies, they could be assumed totally plastic. The presence of such behavior is recognized when considering the stress exerted on the floor rock after failure. To anticipate the behavior of the floor rock, the knowledge of the stress-strain relationship and hence parameters such as Young’s modulus and Poisson ratio is not necessary. In the theory formulated by this method, the floor of the face is assumed to be a base whose length approaches infinity, so that plane strength assumptions hold true. In order to quantify the bearing capacity of the floor rock, in addition to the stressstrain relationship (Equation 4) (Prandtel 1921), required a third equation so that material failure characteristics could be analyzed. He consequently used Mohr-Coulomb (Equation 5) failure criteria to arrive at a model that determines the bearing capacity of the floor rock and analyzes its failure criteria. In this way, his first model (Equation 6) was produced. Equation 8 provides approximate values for the bearing capacity of floor rock. (4)

where x—the horizontal component, z—the vertical component, γ—the density of floor rock.


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τ=c+σf·tanφ (5) where τ—the shear stress before failure (ton/m2), σf—the normal stress before failure (ton/m2), c—the cohesion on the failure plane (ton/m2), —the angle of internal friction of the floor rock (Radian). σb=c·Nc+q·Nq (6) where σb—bearing capacity of the floor rock (ton/m2), q—the uniform load on both sides of the base (ton/m2) (Equation 11). A new parameter is added to equation (6), according to [3], being the effect of powered support unit base dimensions and hence produced a new model (Equation 7). (7) where B—the width of powered support unit base, (m), L—the length of powered support unit base in (m), Nr, Nc and Nq—dimensionless bearing capacity factors (Equations 8 to 10). (8) Nc=(Nq−1)·ctgφ (9) Nr=2(Nq+1)·tanφ (10) (11) In addition to the above parameters, the magnitude of the load applied on the powered support is also dependent on the span of the face roof, powered support unit width, density of the intermediate roof, maximum height of the stress ring, seam dip and the quality of the roof rock. One of the most important of these parameters is the seam dip. To include the effect of this parameter in equation (12), the model introduced in (Deepak 1986) is adopted here. (12) where αf—apparent dip of the face, δ—the coal seam dip, ψ—internal friction angle of the roof rock. Finally, the effect of face roof quality should ideally be included in the model. For this purpose, a safety factor (Sf) is introduced. The values of Sf are given in table 3. Multiplying the original model by the magnitude of the load applied on the face (Pf), equation (13) is obtained.


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Pf=1,15·wf·S′·γ·Cd·Hmax·Sf (13) where Pf—the vertical load on one powered support unit, ton/m2, wf—the span of roof loading one powered support unit, m, S′—width of the powered support unit, m, γ— density of the roof rock, ton/m3, Cd—the dip effect coefficient, Hmax—maximum of the stress ring height, m, Sf—safety factor of the powered support (table 3).

Table 3. Safety factor for powered supports (Sumi Kumar 1995) Safety factor

Immediate roof rock type

1,5

resistant sandstone

1,2

other rocks

The effect of the maximum height of the stress ring above the coalfaces is also shown by equation (14). (14) where M—the seam thickness, m, γaov—the density of roof rock, ton/m3, γf—density of the broken rock in goaf, ton/m3. When the load exerted on the powered support unit’s roof is equal to the bearing capacity of the rock floor, then according to (Deepak 1986) rock failure begins to take place. When this load is increased by the amount of 1,4 tons per m2, floor rock actually breaks and the powered support legs punch into the floor. Equation (15) obtained in this way, shows the effect of floor rock quality on the mechanizability of the coal seam. (15)

where Qf—the floor rock quality index, ton/m2. SEAM THICKNESS Thickness of the coal seam to be extracted is another important determinant of the level of mechanization that could be applied on a typical coal face. It is postulated that seams thinner than 0,6 m are not mechanizable. Limitations in the availability of equipment, also impose an upper limit on the coal seam thickness in order for mechanization to be applied efficiently and economically. This upper limit is taken here to be 3,0 m.


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Peng and Chiang (Peng, Chiang 1984) and Fettweis (Fettweis 1979) also arrived at the same figures, when attempting to find limits of the seam thickness by subjective quantification means. The result of model to be used to determine the effect of seam thickness on the mechanizability of coal seams will therefore have to assume values equal to zero for seam thicknesses either below 0,6 or above 3,0. Hence equation (16) is introduced. F(t)=t−0,6+|t−0,6| (16) where F(t)—seam thickness function, m, t—the seam thickness, m. The value of F(t) obtained from equation (16) does become zero when t is less than 0,6, but with values of t between 0,6 and 3,0 its effect on mechanizability of coal seams is exaggerated by a factor of 2 and for this reason,. F(t) in equation (16) is divided by two. The resulting equation, coupled with a value of zero for situations where seams are thicker than 3,0 m, result in equation (17). This double-component model determines the effect of seam thickness on mechanizability of coal seams in a quantitative manner. (17)

SEAM DIP Another deciding factor as to what level of mechanization can be adopted in a coalface, is the seam dip. Zero or near-zero dip angles are assumed to be ideal for the purpose of this study. As the dip angle increases, the application of mechanization becomes more difficult. The effect of dip on mechanizability of coal seams is, however, different from that in the case of seam thickness, since even in steep seams, some mechanization can be applied or, in other words, mechanizability in these situations, never actually becomes zero. The aim of this research is to provide help in selection of conventional, fully integrated and highly productive set of longwall machinery. In steep dipping seams, such equipment cannot be used on its own and without ancillary equipment, therefore the mechanizability of seams that dip at angles of more than 40° is assumed to be zero. It is plausible to assume that mechanizability of coal seams is inversely proportional to their dip. The most suitable functions that could show such relations in a uniform way are trigonometric functions. Equation (18) is therefore introduced that shows the effect of seam dip on its mechanizability. (18)

where F(α)—the seam dip function, α—the seam dip


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CONCLUSIONS It was concluded that the most important parameters influencing the mechanizability of a coal seam are: roof and floor rock quality, thickness and dip of the seam to be worked. In an attempt to quantify the influence of these factors, mathematical models have been produced that allow presenting them quantitatively. The values obtained from the model of roof rock quality show that when the roof is moderately weak to moderately strong, the mechanizability of the coal seam is high. In the case of floor rock quality, however, the model does not set the upper limit. Mechanizability of coal seam simply increases with the index introduced by the model. Mechanizability of coal seams is most sensitive to seam thickness and it was shown that seams below 0,6 m thick are technically difficult to mechanize and those above 3,0 m in thickness are economically non-mechanizable. The mechanizability index, however, increases in the range of 0,6–3,0 m. It can also be deduced from the model of the effect of seam dip. In the range of 0–40°, mechanizability of a typical coal seam decreases with the values of the seam dip. These models provide indicative values of mechanizability of a coal seam for the mine design engineer, so that the most appropriate and economical level of mechanization may be chosen. REFERENCES Bieniawski Z.T. 1987: Strata Control in Mineral Engineering. Balkema, pp. 110–117. Deepak D. 1986: Longwall Face Support Design-A Micro Computer Model. Journal of Mines. Metals and Fuels, pp. 56–61. Fettweis G.B. 1979: World Coal Resources, Methods of Assessment and Results. Elsevier Scientific Pub., pp. 69–86. Peng S., Chiang H. 1984: Longwall Mining. John Wiley, pp. 215. Prandtel L. 1921: Uber die Eindringungs festigkeit plastischer Baustoffe und die Festigkeit, Von schneuten.Z. Fur Angewandte Mathematik and Mechanik. Sumi Kumar D. 1995: Determination of Optimum Capacity of Powered Roof Supports in the Longwall Face. Journal of Mining Magazine, pp. 111–115. Unrug K., Szwilski B. 1982: Methods of Roof Quality Prediction. State-of-the-Art of Ground Control in Longwall Mining and Mining Subsidence, AIME, pp. 13–29.

A Fuel-Energy System Based on Mining Preparation and Underground Burning of Coal Layers

Gennadij Gayko The Donbass Mining and Metallurgical Institute. Alchevsk, Ukraine International Mining Forum 2004, Kicki & Sobczyk (eds) ©2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: In the work an analysis of underground coal conversion to energy is given. A new concept of fuel-energy system was proposed, suggesting preparation of coalfield and controlled burning of separate blocks of coal layers. Technical and economical comparison with traditional technologies was done. KEYWORDS: Preparation works, underground burning of coal, heatcarrier, electricity

1. INTRODUCTION World energy manufacturing is substantially formed by the influence of two factors. One is a continuous growth of need of energy sources in a combination with prices of traditional sources rising, the second—a problem of environmental protection. One of the main particularities of mining industry is the fact that new mining areas are poorer and (or) less accessible, than the ones worked before. So in the mining branch not only improvement, but also maintenance of current level of technical and economic parameters is impossible, if one sticks to known technological, technical and organizational solutions. In spite of significant progress of new heading and mining techniques, there is no economically effective technology for development of thin layers, which comprise more than 70 percent of world’s coal reserves. The main principle of using nature’s wealth— maximum extraction of mineral resources—is being broken when due to economical reasons the layers with thickness of 1,5 m and less are not mined. Because of this attitude favorable conditions for developing new reserves are limited to 20–30 years, whereupon


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(if the utilized technologies are not changed) coal (i.e. electricity) price will rise dramatically. The trend to extend the sphere of using open-cast mining has been already exhausted because geological perspectives point to the necessity of the consecutive exploitation of poor, deep-laying coal deposits, in complex mining and geological conditions. Besides, the cost of rehabilitation of opencast mines sites makes them unprofitable in many cases. In this context one of the most prospective trends of development of fuel-energy system is using the geotechnologies, based on underground coal gasification. 2. RETROSPECTIVE ANALYSIS OF NON-TRADITIONAL METHODS OF COAL SEAM DEVELOPMENT The idea and technological principles of underground coal gasification was formulated in 1888 by great Russian chemist D.I.Mendeleyev, who studied fire accidents on the mines of Donetsk basin (Gayko 2001). “Probably, when the time passed an epoch would begin—Mendeleyev wrote—when the coal would not be extracted out of the Earth, but there in the Earth, it could be changed into flammable gases, which in their turn would be transported on long distances” (Mendeleev 1888). And then: “…having drilled some holes into the layer, one of them have to be used for input or blowing in the air, other ones for output or drawing off flammable gases, which would be easily transported then to the furnaces on long distances”. The designs of underground coal gasification utilizing holes (W.Ramsey 1910–1915) and the mining method (B.I.Boykij 1925) proposed afterwards weren’t accepted for industrial use because of a number of unsolved problems. However in 1934 in Lisichansk (Donetsk region) the first experimental station of underground coal gasification was built. By 1941 in the former USSR artificial gas was produced by 10 stations. The technology of work consisted of drilling pumping and gas-removing holes, connecting them by gasification channel and burning up the coal layer. The exploitation experience of the first enterprises of underground coal gasification has revealed the great disadvantages of this technology: big energy losses in the Earth (60–70% of coal energy was spent for useless heating of rocks); difficulties in controlling the process of coal burning (as a result—unstable gas characteristics, uncontrolled spread of surface deformation); low heat value of the produced gas (3–5 MJ/m3—as compared to 8–10 MJ/m3 for timber, or 30–35 for natural gas). These reasons, and also opening of rich oil and natural gas deposits lead to phasing out of research on industrial coal gasification. The experimental stations working now limit themselves only to search for technical solutions that could make these methods competitive. In view of this, the achievements of O.V.Kolokolov’s scientific school (National Mining University of Ukraine) are of great interest, in particular using hightemperature combined blowing, reversing of blown and productive flows; using of flexible pipelines to deliver the blow into the center of burning place; gasification under pressure; using the injection input of out-gased space and so on (Kolokolov et al. 2000). New developments greatly improve the perspectives of underground coal gasification,

A fuel-energy system based on mining preparation

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but to realize its controlled process by means of technology utilizing holes is still problematic. Many disadvantages of underground coal gasification could be avoided by full burning of coal layers with the aim to obtain thermal and electric power. Unlike systems oriented at obtaining flammable gases, the technology of underground coal burning is directed at extraction of gaseous heat-carrier, which gives off its energy to water by means of heat exchangers. The first propositions on heating the heat-carrier (water, steam) by burning coal in the place of its bedding were brought forward by Yu.D.Dyadkin in 1980 (SaintPetersburg Mining Institute) (Arens et al. 1991). Development of technological principles of underground coal burning was carried out in the 80’s of the 20th century with V.V.Rzhevskij at the head (Moscow Mining Institute) (Rzhevsky, Selivanov 1989). Mine experiments on underground coal burning undertaken in Ukraine in 1983–1991 (Kolokolov et al. 2000) have proved the stability of burning process. At this the maximum burning temperature of 900–1000°C was achieved, the average burning speed was nearly 1 meter per day; heat losses on heating the surrounding rock were less than 40%; temperature of the gaseous heat-carrier was insufficient for energy production and was varying from 130 to 250°C. Comparing the hole and mining technologies allows considering the mining method as more promising. 3. THE CONTROLLED METHOD OF UNDERGROUND BURNING OF COAL LAYER High efficiency of underground coal burning could be achieved with producing electric power directly in a mine. For this it is necessary to have the output temperature of gas (or heat-carrier) of 600–800°C that in its turn requires providing activity and stability of the burning process and also decreasing heat losses during moving the heat-carrier from the fire work place to the turbines of electric power station. To achieve this goal the author has developed a new control method for the burning process and heat exchange, based on preparation of power blocks by mining and using a pipe collector for heat-carrier’s circulation. A general layout and principle of operation of this method are shown on figures 1 and 2. This is how the method is realized. According to parameters of power station, which is sited on the most productive mine level the dimensions of blocks 1 of coal layer are determined. Gullies 2 are developed around blocks 1. Temporary support (for example, steel support, roof bolting), which allows access to coal, is used in the tunnels. Breakthrough channels 4 are made along the coal seam. Grooves 6 are drilled in the floor. It is best to develop channels 4 and grooves 6 by boring. In grooves 6 metal pipes are placed forming a pipe-collector 5 along the fire working face 7 the lighting devices are placed (they are not shown). After that concrete support 3 is erected in gullies 2, making power blocks hermetical and heat-insulated.


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Figure 1. Method of underground burning of coal using steam and flammable gases as heat-carriers


75

Figure 2. Sequence of burning of coal blocks In this way several complexes of blocks are prepared and coal is burned in them one by one. To do this the lighting devices are put into operation by remote-controlled method, and stream of air is delivered to fire working place 7. The output gases move through the breakthrough channels 4 to gas pipeline 9 and, using a pumping method, they are delivered to the turbines of power station or heat exchangers. During the coal burning process the temperature in the block in operation reaches 1000–1500°C. At this the caved-in rock of the goaf 8 is heated up to the same temperature and keeps it for a long time. All this time into pipe collector’s input 5 the water (power-carrier) is being delivered and gas of high temperature is obtained at the output and is transported to the power station’s turbines. By regulating the circulating speed of heat-carrier in the collector depending on the required temperature and quantity of steam one can control the process of heat exchange in the power block. The intensity and speed of coal burning


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can be influenced by the quantity of input air. Coal burning is carried out without presence of people in mine working places. Of great interest is the sequence of processing of power blocks given on figure 2. In the first stage coal burning is done in block, which is closest to the generator plant chamber 12. When working the next block steam in collector 5 and output gases in goffered pipeline 9 will be transported through high-temperature zone of the previously worked-out block, which will exclude the power loss during the transport to bowlsutilizers 10 and turbines of power station 11. It is best to equip the power station not only with turbines powered by hightemperature steam and gas burning, but also with aggregates of geo-thermal power stations. This would allow to subsequently, and for a long time, use the geo-thermal energy accumulated in goaf after burning coal in power blocks. 4. TECHNICAL AND ECONOMICAL FEATURES OF THE NEW METHOD OF UNDERGROUND COAL PROCESSING Thanks to the fact, that the layer of coal is divided into blocks by gullies with heatinsulating support, borders of spreading of the coal burning process are provided and the necessary block dimensions according to the requirements of electricity-generating machinery and collector system are formed. The use of pipe collector, in which circulation of heat-carrier is done, makes it possible to affectively transfer the energy first from coal burning and then from heated goaf and surrounding rock, to the heat-carrier (water, steam) and its controlled transport within the borders of a block and outside it. For the first time the process of underground coal burning allowed to utilize not only the products of burning, but also hightemperature steam, transported to electricity-generating turbines. The fact, that pipe collector is placed in the seam’s floor allows to protect the pipes from being destroyed by roof caving in during coal burning. Transporting of steam and gas—the product of burning, in pipes through hightemperature area of previously worked-out blocks decreases heat loss of heat-carrier to a minimum and allows producing electric power by using the regular equipment. Comparing with traditional mining development, construction of power station underground will simplify the layout and decrease the length of development works. The method presented here changes the traditional technology involving complex operations of shearing coal and transporting the output from the mine workings to power stations on surface, to machineless, operated by minimal number of people technology of burning coal layers and producing electric power from high-temperature power-carriers directly in underground conditions. It allows solving the problem of effective development of thin coal layers and greatly decreasing the cost of electric power. Besides, a number of ecological problems of mining are solved because the ash and slag remain in the places of burning. Controlled method of underground coal burning without the presence of people solves the social problem of safety of mining operations. The author is currently looking for partners to investigate and experimentally prove this theoretically developed method and will be thankful for contacts in this regard.


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REFERENCES Arens V. et al. 1991: Non-Traditional Solutions in Mine Production (NEDRA: Moscow) ISBN 5– 274–01866–4. Gayko. G. 2001: History of Mine Techniques (DMMI: Alchevsk) ISBN 966–7560–28–7. Kolokolov O. et al. 2000: The Theory and Practice of Thermo-Chemical Technology of Extracting and Processing Coal (NMU: Dnepropetrovsk) ISBN 966–7476–35–9. Mendeleev D. 1888: Future Force, Resting on Donets River Banks. In: North Herald, № 8–12. Rzhevsky V., Selivanov G. 1989: Underground Coal Burning (MMI: Moscow) ISBN 5–274– 01231–1.

Exploitation of Technologically Generated Methane Deposits by Means of Surface Wells

Alexsandr Egorovich Vorobyov Russian University of Peoples Friendship. Moscow, Russia Tatyana Vladimirovna Chekushina Research Institute of Comprehensive Exploitation of Mineral Resources, Russian Academy of Sciences. Moscow, Russia Grigoriy Artemovich Balykhin Ministry of Education of Russian Federation. Moscow, Russia Alexsandr Dmitrievich Gladush Russian University of Peoples Friendship. Moscow, Russia International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 According to the developed concept of formation of mineral reserves and the results of the authors’ own research, technologically generated deposits (i.e. deposits formed as a result of human action) can be formed on the surface (from waste products of mining and from the concentration plant—metallic corns and tailings), as well as in the lithosphere depths. Because of high wastefulness of mining technologies the first techno-gene deposits group sharply prevails in the total amount. Until now less attention was given to technogene deposits from the Earth basement and while they can be basically used in the translation of reserve deposits into balanced (as the result of a partial disclosing of the thin dispersing mineralization, as the result of the ores organic component oxidation, as a result of the technological harmful impurities dissolution and removal and as result of other similar processes). The useful mineral category also has a determining value. In the resource producing mining technologies the most studied is the behavior of the non-ferrous and precious metals. The unconditioned coal, combustible slates, bitumen, asphalt, oil, combustible gas, chemical raw material, building materials, facing stone and mineral waters also can be ennobled and transferred to the category of techno-gene mineral raw material. For example, methane techno-gene deposit in the stage of the preliminary industrially decontaminated coal layer can be formed as the result of the gas redistribution from the infringing techno-gene containing breeds (where the concentration of methane makes in


80

aleurolites and argillites—0,25–2,0 m3/m3, in sandstones—up to 7), and by its microbiological formation. Planning the redistribution of the gas it is necessary to take into account the different sizes of methane volume from the coal deposits and containing breeds. So, within the limits of the mine fields of Kuznetsk and Pechora coal basins there is only 304 billion m3 methane, and in their containing breeds (up to depth 1800 m) already 15 trillion m3. The features of the resource producing mining technologies have allowed creating a number of new ways for improving the initial quality of the natural and techno-gene mineral raw material. Now the world coalmines are allocated 29030 billion m3/year of firedamp (20–21 billion m3/year of methane and 8–10 billion m3/year carbon dioxide) a day. Only in 1991, to 3000–3100 world mines rich in methane were allocated more than 20 billion m3 of methane (in the mines of

Table 1. Characteristics of the mines with high native pressure of methane Mine

Output of gaseous substances %

RUSSIA

Pechora coal basin

Deposit’s Methane Methane’s debit depth pressure thousand m MPa m3/day

Polar

32,4

440

4,8

210

Northern

31,5

750

7,2

484

Vorkuta

32,2

685

7,4

258

Kuznetsk coal basin Abashevskaja

29,0

540

3,3

180

Chertinskaja

30,0

365

3,8

178

Komsomolets

23,6

380

4,1

145

Named After S.M.Kirov

26,6

480

4,9

125

Octyabrskoe

28,5

330

5,0

141

42,6

450

4,7

106

445

4,2

86

Uglovsky coal basin Capitalnaya

Partizanskii coal basin Nagornaja

36,6

UKRAINE

Donetsk coal basin

Cold Beam

27,3

630

6,5

104

Named After A.I.Gaevovo

27,0

740

7,6

117

Named After K.I.Pochenkova

21,0

915

7,9

180

Exploitation of technologically generated methane deposits

81

Red Profintern

21,0

865

8,0

122

Junkom

18,0

826

8,4

115

Named After M.I.Kalinina

22,8

1307

9,0

275

Komsomolets

30,9

940

9,5

100

Named After A.Skochinskogo

24,7

1270

9,6

206

Named After A.F.Zasjadko

26,5

1230

10,0

145

Named After A.F.Zasjadko

26,6

1100

11,8

139

Petrovskaya Glubokaya

25,0

1196

12,0

121

Ignatevskaja

24,0

597

13,4

205

KAZAKHSTAN

Karaganda coal basin

N 37

30,0

320

5,4

100

Nam. After 50 Years of October

29,5

600

6,2

174

FRANCE Lens CHINESE PEOPLE’S REPUBLICS

Coal basin of Holes and Pas de Calais 28,0

1100

12,0

135

Tjanfu coal basin

Jujutszjagou

22,7

490

5,8

155

Mosinpo

24,4

625

8,0

271

418

6,8

190

729

8,2

245

300

2,1

338

255

2,6

767

3,0

600

Nantong coal basin Ljtjanbjao

20,5 Bejpjao coal basin

N7 USA Lavridge

27,4 Pennsylvania coal basin 29,0 Virginia coal basin

Blue Shout

33,0

Black worrier coal basin Ouk I Rove

32,0

355

the former USSR—7 billion m), to this it is necessary to add 3,9 billion m of methane, captured from 619 mines (including the mines of the former USSR—1 billion m3), hence, the total methane debit from the world mines exceeded 23 billion m3. To the mines of the CIS, Germany, Chinese People’s Republic, the USA, Australia and other countries that where rich in methane has been allocated by 0,3–0,7 million m3/day of methane, or up to 150–200 m3 per 1 ton extracted coal.


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At low methane content (less than 5 m3/t of coal), and it is the most widespread case, its industrial extraction becomes unfavorable (for economic reasons) and then it is necessary to apply new technologies, namely, creation of techno-gene coal-bed methane deposits. The natural methane coal depends on methane native pressure from the coal and their absorption abilities. Methane absorption from coal (9 m3/kg) at the temperature 288°K and pressure 0,1 MPa linearly depends on the energy of the depressive relation Eд: (1) where ao—represents the gas absorbed molar share, per one unit of its quantity in free condition at temperature 288°K and pressure 1 Pa (ao=1,363·10–3 mol/kg·Pa), EД— represents the molecular bond energy of the sorbent, EДК—represents the critical value of the sorbent depressive energy, below which gas absorption in the given sorbent does not occur (EДК=0,5 Jol/mole). After the condition 2–3 is met, a transition of gas molecules from a free phase in a sorption condition is carried out: (2) (3) where Eдиc—represents the energy of the dispersing interaction of one gas mole, Jol/mole, R—represents the gas constant (R=8,31 Jol/mole·K). For the maintenance of a basic opportunity of methane techno-gene reproduction in coal deposits with the purpose of their subsequent industrial (trade) extraction, it is necessary to establish a communication between the highest levels of the material organization, comprising a coal layer and containing breeds, and their micro-structural properties. In the natural layering conditions of the coal beds, the prevailing process of the diffusion gas transfer on the coal and breed practically does not change the homogeneous condition of the natural system “coal (breed)—gas—natural moisture”. The quantitative characteristics of the gas-coal transfer system in homogeneous conditions are the values of the mechanical pressure and the concentration (maintenance) of methane corresponding to such transfer in various coal components. In such conditions, the coal-substance is practically stable, and in the gas evolution (gas restitution) from the coal prevail diffusion processes. If gas restitution of a layer is made without the application of appropriate stimulating methods it goes extremely slowly and, as a rule, does not have practical value. Besides, the gas restitution processes in a coal layer and its possible degree of gas exhaustion when absorbing methane is inextricably related to the coal structural transformations, which are defined by the processes of destruction and semi-condensation. The mass transfer of coal, determining its transformation and its intensity of gas restitution, and the gas exhaustion volume, depend on the natural type, and in the greater degree on the artificially created defects and the forms of the gas existence in coal and nature. Using the physical complex, the physiochemical and coal petrographic parameters in vitrite (the basic coal variety, that determines the gas restitution of a layer), the presence


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of the polymeric substances with linear and the ramified macromolecular structures of globe forms and the prevalence in them of the linear structured fragments were determined. Therefore, the determining value for the methane effective redistribution in a coal layer and in the containing breeds have the conditions of stability and homogeneous disintegration of the solid coal-gas solution (TUGR) with gas liberation. The methane pressure (table 2) in macro-pores, in the coal layers’ natural fractures and in the containing breeds is one of the major factors (parameters), determining the volume of the absorbed and free methane. Methane in coal layers is basically in an absorbed condition (a solid coal-gas solution) and adsorbed (on macromolecules and micro-cracks surface), also in rather small volumes in a free condition. With depth, the quantity of free gas grows due to pressure increase practically in all conditions and it reaches 10–12% of the total contained gas volume.

Table 2. Dynamics of the maximal natural gas pressure (MPa) in the coal layers on CIS basic basins Layer depth m

Coal basin Kuznetsk

Pechora

Donetsk

Karaganda

200–300

1–2

2–3

N.A.

1–3

301–400

3–4

4–4, 5

1–3

2–5

401–500

4–5

5–6

3–5

4–6

501–600

5–6

6–6,5

4–6

5–6,5

601–700

6–7

6,5–7

6–8

6–7

701–800

–

7–7,5

7–9

7–8

801–900

–

7–8

8–10

–

1001–1100

–

8–9

9–11

–

1101–1200

–

–

10–12

–

>1201

–

–

12–14

–

In natural conditions in coal layers (and in containing breeds) there is a dynamic balance between the absorbed methane and methane in a free phase, which is characterized by this that at any moment of the layer condition the methane number of molecules passing from the free condition in the absorbed phase, is practically adequate to the number of methane molecules passing from the absorbed into the free condition. Free coal-bed and mountain massive methane occurs in natural cavities volume, within the limits of which methane molecules interaction with coal or rock surface is rather small; therefore methane volume in the natural emptiness free phase is insignificant. In such conditions the free methane plays the role of the “support” for the absorbed methane, substantially impeding the methane adsorption from the coal. The infringement of the dynamic balance


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in the natural system “coal (breed)—methane—natural moisture” due to the influence of redistribution of stress due to mining works demonstrates itself in excess of the number of adsorbed methane molecules above the number of the absorbed ones. The methane volume in the free phase at the depth of 700–1200 m makes up: in coal of average degree of metamorphism 5–10%, in coal of high degree of metamorphism 4–6% and in coal of little degree of metamorphism 10–14%. At the infringement in the system “coal (breed)—methane—natural moisture” the established balance of methane molecules movement from the natural coal pores occurs as by their moving on the pores walls, and by methane molecules from an absorbed condition into adsorbed condition. Then they are allocated into free phase in a zone methane decrease pressure, before existing in a free phase, then starts the disintegration of the coal-gas solid solution, accompanied by the decrease of native gas pressure. With other similar conditions (natural methane, metamorphism degree, porosity, humidity, etc.), the time of the coal layer degassing is approximately 4 times greater than the degassing time of the collector non-absorbing pores (sandstones, sand, etc.). THE COAL LAYERS AND CONTAINING BREEDS’ METHANE CONTENT The natural distribution volumes of methane content in carboniferous thickness up to depth of 1800 m have the following appearance: – coal layers (the general capacity of 1–6% depending on the deposit methane content) contain 20–50% of all methane, – breeds containing coal layers (up to the capacity 94–99% of the carboniferous thickness)—50–80%. The methane content in the carboniferous thickness and in the containing massive is defined by the existing thermodynamic conditions: temperature, gas pressure and physiochemical properties of the coal type weights—absorption ability and metamorphism degree, porosity, permeability, humidity. Besides these, essential influence on the content of methane render methane deposits’ preservation conditions— the presence, the structure and the capacity of the quaternary adjournment; the deposit’s depth from the surface; the presence and capacity of permafrost breeds or saliferous thickness; the coal deposits; the presence, the capacity and the temperature of intrusives, etc. The major factors determining coal-bed methane content are also metamorphism, the material structure and tectonic processes. In the majority of coal basins (deposits), the quantity of natural coal methane content depends on the depth of the deposit and on the degree of its metamorphism, in regular intervals increasing from brown and long flaming coal to lean and to anthracites: the layers of the high metamorphosed coal (grades A12A14) are characterized by rather low natural methane content (up to 2 m3/t). The increase of the coal layers methane content with the depth corresponds to the growth of the sorption ability and to change of the coal porosity; the volume of the coal sorption of methane increases with the gas pressure growth and reduces when increasing the coal temperature and humidity. The isotherms of coals with methane content are


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practically stabilized at the gas pressure equal to 6–8 MPa. With increasing depth of the coal layers deposits, the temperature of the containing breeds raises due to the fact that the coal sorption ability is reduced, despite of the fact that the gas pressure raises. The intensive (practically linear) growth of the natural coal methane content occurs depending on the coal metamorphism degree up to a certain depth. Further, the increase rate of the methane content growth is slowed down and even stops: for the CIS anthracites deposits—from the depth of 600–650 m, for the coal of average degrees of metamorphism—continues to increase up to the depth of 1200–1300 m. Within the deposit limits of 300-meter depth, the natural methane content of coal grows most intensively. COAL BASINS AND DEPOSITS METHANE RESOURCES The coal layers methane content depends on the general gas dynamic processes, proceeding in the carboniferous layer thickness, and on the local geological conditions of methane preservation in a deposit strata—the degree of coal metamorphism, the natural gas permeability and the dip of the strata, the character and the intensity of formation of geological disturbances, etc. For the majority of coal deposits, the distribution of the gas ash value differs with a character of gradual change in the gas structure from the surface to the deposit’s depth of occurrence from air (CO2, nitrogen) to nitric (N2, CO2), nitrogen-methane (N2, CO2, CH4) and hydro carbonic gases (CH4). The reduction of the stratifications within the limits of the methane deposit leads to the growth of natural coal methane content. In natural conditions, outside the zones of influence of geological disturbances and intrusives, the basic coal methane volume (up to 85–96%) is in a sorption condition with prevalence of the absorbed methane in the form of coal-methane solid solution. The total volume of methane reserves within the limits of the main, controlled by the Russian coal industry, basins is estimated at 22–26 trillion m3, but only in the coal deposits up to depth of 1800 m from a surface (under the estimation) contain 72–79 billion m3 of methane (table 3). The general expected methane reserves in the largest basins of Russia (Kuznetsky, East Donbass, Pechorsky, Sakhalin and Partizansky) up to depth of 1800 m from the surface make 17–20 billion m3, or 17% of the world methane reserves in coal layers and about 10% of the natural gas reserves of Russia. The total expected methane reserves in the largest coal basins of the world—Tungus and Lensk make about 50 billion m3. The average specific methane Russian reserves reach 300–900 million in m3/km2 in Kuznetsk basin and on Apsatskom deposit. The maximum specific methane reserves of Kuzbass (average basin 720 million m3/km2)—1600–2100 million m3/km2 in Erunakovskom, Mrasskom and Tom’-Usinsk areas. In Kuzbass, the general reserves of methane make 265–280 billion m3. The methane reserves (trillion m3) from the coal layers of the main areas of the Pechora basin are distributed as follows: – Vorkutinsky 0,1–0,13, – Halmer-Jusky 0,3–0,6, – Vorgashorsky 0,2–0,3,


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– Usino-Sejdsky 1,3–1,7.

Table 3. Expected methane reserves in the CIS countries coal basins Average methane content of coal layers, m3/t, at depths:

Methane resources Specific, average, million m3/km2

Specific, maximum, million m3/km2

Up to 1000 m

1200 m

Average, billion m3

Lensky

5–10

15–18

8410

230

300

Partizansky

5–10

14–18

20

6

10

Chelyabinsky

6–9

7–11

5–6

4

15

Uglovsky

6–10

25–30

5

10

35

Sakhalinsky

6–12

20–30

96

31

48

Minusinsky

7–10

12–13

12

56

80

Arkagalinsky

8–9

24–26

6

2

5

Bureinsky

8–13

18–20

132

27

50

Ulughemsky

9–10

12–13

142

57

75

Zyrjanovsky

10

15

296

32

45

Kizelovsky

10–12

26–32

10

0, 5

3

SouthernYakutsky

10–13

15–19

190

10

15

Eastern Donbass

12–16

20–22

–

90

N.A.

Apsatskoe deposit

12–18

20–21

18

250

320

Golovsky

16–18

20–25

122

230

N.A.

Tungusky

16–19

24–35

4140

40

N.A.

Kuznetsky

18–23

31–35

13200

720

3400

Pechorsky

18–24

22–38

1950

60

500

LvovskoVolynsky

10–12

20–25

5

0,7

2

Donetsky

17–22

23–32

2500

24

410

Coal basin

RUSSIA

UKRAINE

KAZAKHSTAN


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Ekibastuzsky

12–15

20–24

90

320

430

Karagandisky

17–20

24–26

2000

350

410

The general expected methane reserves from the coal layers of the Pechorskogo basin make 1,9–2,63 trillion m3/km2, specific average—60 million m3/km2 and maximal—500 million m3/km2 (including the coal layers of the Vorkuta deposit 340 million m3/km2, Usinsk 200 million m3/km2 and Vorgashorskom up to 300 million m3/km2). The maximum specifics content of methane reserves within the limits of the Pechora basin are: the Vorkuta deposit 2650 million m3/km2, Usinsk—1460 and Vorgashorskom—840 million m3/km2. So, within the area of “Vorkuta” mine up to depth of 720 m the general methane reserves of the coal layers and of the containing breeds are approximately equal and exceed 40 billion m3, and the “podsvitah” methane resources are practically identical and make in Rudnitskoj “podsvite” 10, 7 billion and in Intinskoj—9,6 billion m3. The expected world of congestions volume (table 4) makes 6–10% of the total methane reserves and is distributed among the basins as follows (billion m3): Kuznetsky – ~100, Donetsky—30–35, Pechorsky—up to 28, Karaganda—5–8, Chelyabinsky—1–2; the USA basins—280; People’s Republic of China—about 300; Japan—80; Poland— 300. In total, the world free congestions contain up to 6–9 billion m3 of methane.

Table 4. World total methane reserves in coal basins and deposits Country

Total reserves CH4 trillion m3

Country

Total reserves CH4 trillion m3

CIS countries

72–79

Netherlands

0,4

People’s Republic of China

30–40

Japan

0,2–0,3

USA

11–12

Columbia

0,2–0,3

Australia

6

Belgium

0,2–0,3

Germany

3–4

Romania

0,15

Republic of South Africa

2,0–2,3

Mexico

0,1–0,3

Great Britain

1,9–2,8

Spain

0,1

Poland

1,6–2,0

KNDR

0,05

India

1,4–1,5

Bulgaria

0,05

Czechoslovakia

1,1–1,5

Turkey

0,02

France

1,0–1,5

Indonesia

0,01

Hungary

0,8–10

Yugoslavia

0,01

Canada

0,6–1,6

Zimbabwe

0,09

Austria

0,5


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THE TECHNOLOGICAL PRINCIPLES OF CREATION OF METHANE-CONTAINING TECHNO-GENE DEPOSITS Currently, ahead of the Russian Federation mining industry raises the problem to increase the efficiency of methane extraction from coal layers. This can be achieved by reproduction of methane by redistribution of the gas accumulated in the containing breeds, into the coal deposit. The objective can be achieved as follows. After exploitation of methane in a traditional way is complete (drilling the wells, hydro-fracturing the coal layer, initiation of methane sorption and extraction until methane is exhausted) it is necessary to plug the wells, and to form infringements of integrity in a file underlying methane-containing breeds. Then by the techno-gene fractures and other infringements, methane, which is in a free form in a file of containing breeds, will be redistributed in a coal layer where it will pass into the coal-sorption form. After methane redistribution up to its former volumes (or at their excess) in a coal layer, its extraction with traditional technology is continued. We shall note, that if in coal layers up to 90% of methane are in the sorption condition and only 10% in free, in sandstones containing a coal layer there is an inverse relationship. It is possible to redistribute methane from the containing breeds only by forming gas channels—infringement of the breeds’ integrity (without carrying out other operation—adsorptions). To reach the necessary methane deposits for techno-gene formation of significant volumes, it is necessary to take into account the methane content of the containing breeds. So, the average value of the natural methane content of the sandstones is 5 m3/m3 that is 4–5 times less than coal methane content. Therefore at the reproduction of a methane deposit it is necessary to involve the corresponding volumes of the containing breeds. Hence, the formation of a methane deposit is possible due to the unused earlier parts of methane, which are in containing breeds. The method works as follows. The methane-containing coal layer is opened with a well. Then the well is sealed, after that a hole is drilled and a coalface cavity water delivery is formed. After that, the layer is hydro-fractured by water pumped under the pressure of 10–12 MPa. When the fracture system is formed, it is made the methane initiation contained in the coal layer. For this purpose, through the well compressed air is passed. The extraction is done after the end of the initiation cycles, which gradually exits through the well to surface. After about 5–10 years of the coal layer degassing, the basic methane quantity is taken on the surface and the well’s output is sharply reduced. The methane content of the containing breeds located under the coal layer is released by detonating an explosive charge. Through the channels, which have opened into them and through the fractures there is a migration of the, contained before, methane being mainly in a free condition from the broken zone. But because these breeds occur under the coal layer, the migration is directed to the layer in which coal is a natural collector for methane. As result, there is a techno-gene formation of methane deposit in a coal layer. And only after that methane extraction through a well starts. The positive effect of the offered technical process lies in the increasing methane extraction efficiency from a coal layer due to its reproduction using redistribution of the gas contained in the containing breeds of the coal deposit.


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This technology can be used in the industrial (preliminary) methane extraction from the coal layers, which have not been touched by mining works. Except for the considered methane extraction technology efficiency raises its microbiological reproduction. We shall consider a specific case of the process of formation of techno-gene methane deposits in a layer. The methane-containing coal layer 1 (figure 1) is opened with the well 2. Then the space 3 is sealed, a punching 4 is made by encircle columns (it is not shown on the plan) and a coalface of water 5 cavity deliveries is formed. Hydro-fracturing of the layer 1 is conducted by the hydraulic influence of water, which is pumped under the pressure of 10–12 MPa. When a system of fractures 6 is formed, the methane initiation is done. For this purpose into the layer 1, through the well 2 carbon dioxide is pumped, which replaces methane from the coal and subsequently, is not extracted to the surface. The methane extraction is done at the end of the initiation cycles—it gradually exits through the well 2 to the surface. After 5–10 years of the coal layer 1 degassing the basic methane quantity is extracted to the surface and the wells’ 2 methane output is sharply reduced. The worked site of the coal layer 1 is shielded from the not exploited massive with screens 7, then it is consistently submitted (together with life products providing ability products) with microorganisms oxidizing the free oxygen, and after its oxidation—water bacteria. It is necessary to take into account the known in the natural systems, but unused until the present moment techniques and technologies of process of methane renewal. So, the microbiological research has shown, that the bacterial methane restoration occurs exclusively due to the carbonic gas restoration with hydrogen. The conditions of the many mines (temperature 15–20°C and pH=6,4–7,5) promote the course of similar micro biochemical processes for the purposeful methane techno-gene formation. And because methane is formed in an anaerobic environment, then for the development of these processes it is necessary to reduce the oxygen concentration in the coal layer. Thus, the techno-gene process of micro-biochemical methane producing includes the following stages of development: – oxidation of the free oxygen with special bacteria, – hydrogen microbiological formation, – the formation of methane as the result of the reaction: CO2+4H2=CH4+2H2O (Vorobyov, Hotchenkov & Chekushina 2002).


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Figure 1. Formation of techno-gene methane deposits in a coal layer: 1— the coal layer, 2—the well, 3—sealed space, 4—well punching, 5—the cavity, 6—the fracture, 7—the screen As the sorption result of the synthesized gas a techno-gene methane deposit in the coal layer is formed. Then its extraction through well 2 starts. Other performance variants of the proposed method are presented below. After opening of the coal layer with a well and after end of the works at the sealed spaces, there are punched encircled columns, and hydro-fracturing of the layer is done. The pressure of forcing the working fluid (water) into the layer, the volume and the pumping rate are defined using standard techniques. The pump units 4AH-700 are used. After the set volume of working fluid is pumped-in, the well is closed for 2–3 months. Then, four compressors KPU-16/250 are connected to the well, which pump carbon dioxide into the layer with the rate of 64 m3/minute and pressure of 12 MPa. After pumping of the set volume of the gas, the well is closed for 10–15 days, and then methane extraction from the layer in a self-expiration mode is conducted. With years (5–10 years) the basic methane quantity is brought to the surface and the wells output is sharply reduced. Then the worked-out site of the coal layer separated from other file by a screen (for example, an injection of cement grout) then together with the substances life providing ability, solutions of microorganisms (from the aerobic class) oxidizing the free oxygen. The well is closed and maintained before oxidation of 90% of oxygen. Then in the layer fall the microorganisms’ solutions, forming hydrogen (the main producers of hydrogen—hemitrope bacteria which make an anaerobic decomposition as fermentation of the various organic substrata, for example, the representatives of the sort Clostridium, Rhodospirillaceae or zinc bacteria). The well is again closed before the formation, as the result of the interaction of hydrogen with carbon dioxide according to


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the formula 3.4, techno-gene methane in industrial concentration (for example, up to 10– 15 m3/t). This is the method how techno-gene deposits of methane are created in a coal layer. After that methane extraction through a well starts again. The specified technology of microbiological methane reproduction is not unique, but because it is possible to use microorganisms oxidizing free oxygen and forming CH3COOH compound, and then bacteria translating CH3COOH to methane. As result there also happens a methane renewal up to the industrial values. After that its extraction with the use of traditional methods continues.

Figure 2. Formation of techno-gene methane deposits in a coal layer Layers: 1—coal layer, 2—the well, 3—sealed space, 4—well punching, 5—fracture system Thus, the process of techno-gene micro biochemical methane producing includes the following development stages: – Oxidation of the free oxygen and cell with special bacteria before the formation of CH3COOH compound; – Microbiological methane formation as the result of the reaction: CH3COOH→CH4+CO2 (Nikitin, Potapov, Gnoevyh & Krylov 2000). The developed method works as follows (figure 2). After the methane extraction in the worked site of the coal layer 1 consistently fall (together with the products assuring life) microorganisms oxidizing the free oxygen and forming CH3COOH and then methane forming bacteria. Microbiological research has shown, that the bacterial methane restoration occurs due to the transformation of CH3COOH compound. The conditions of the many coal layers (temperature 15–20°C, pH=6, 4–7, 5) assure the course of similar micro-biochemical processes having the purpose to form techno-gene methane. As the result of sorption synthesized gas coal occurs the formation of the techno-gene methane deposits in a coal layer. After that extraction through the well 2 starts. After 5–


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10 years of exploitation in the worked site together with the other substances providing bacteria life (with waters containing cellulose) solutions of microorganisms, which oxidize the free oxygen and form CH3COOH, then the microorganisms of the group Methanosarcina or Methanothrix. The well is again closed before the formation (as a result of bacterial interaction) of the techno-gene methane of the industrial concentration (for example, 10–15 m3/t). REFERENCES Vorobyov A.E.: Kompleksnoe ispol’zovanie uglemetanovych mestorozhdenii. Sovremennye problemy shahtnogo metana. M.: MGGU, 1999. S. 257–262. Vorobyov A.E., Balyhin G.A. & Gladush A.D.: Technogennoe vosproizvodstvo uglevodorodnogo syr’ya v litosfere: faktory, mechanizmy i perspektivy. M.: Izdvo “Ucheba” MISiS, 2003. 417 s. Vorobyov A.E., Gladush A.D.: Sposob degazacii uglemetanovych mestorozhdenii: Patent 2198297 RF, 2003, BI № 4. Vorobyov A.E., Hotchenkov E.V. & Chekushina E.V.: Technogennoe resursovosproizvodstvo uglevodorodnogo syr’ya v glubinach litosfery. GIAB. № 2,2002. M.: MGGU. S. 119–124. Nikitin B.A., Potapov A.G., Gnoevyh A.N. & Krylov V.I.: Sowemennoe sostoyanie i perspektivy razvitiya techniki i technologii stroitel’stva neftyanych i gazovych skvazhin. Nauka i technologiya uglevodorodov, № 6, 2000. S. 71–83. Molchanov V.I. & Goncov A.A.: Modelirovanie neftegazoobrazovaniya. Novosibirsk, Nauka, 1992. Orientirovochno dopustimye urovni (ODU) chimicheskich veshestv v vode vodnych ob’ektov hozyaistvennopit’evogo i kul’turno-bytovogo vodopol’zovaniya: Gigienicheskie normativy. GN 2.1.5.690–98. M.: Rossiiskii registr potencial’no opasnych chimicheskich i biologicheskich veshestv Minzdrav Rossii, 1998, 45 s. Trubeckoi K.N. & Vorobyov A.E.: Klassifikaciya metodov vosproizvodstva mineral′nogo syr’ya. Gornyi zhurnal, № 1,1998. S. 30–34. Fuks I.G. & Evdokimov A.Yu.: Topliva i smazochnye masla na osnove rastitel’nych soedinenii. M.: CNIITENeftehim, 1992,24 Koptyug V.A.Zhurn. Ros. him obsh im. D.I.Mendeleeva, 1993. T 37, № 4. S. 4. Shahnovskii I.M.: Sovremennye predstavleniya o genezise neftyanych i gazovych mestorozhdenii. Geologiya, geofizika i razrabotka neftyanych mestorozhdenii, № 7.1999. S. 17–22.

Methane as a Source of Energy in an AirConditioning System in “Pniowek” Coal Mine

Nikodem Szlązak AGH—University of Science and Technology. Cracow, Poland Andrzej Tor, Antoni Jakubów, Kazimierz Gatnar Jastrzebie Coal Company. Jastrzebie Zdroj, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: The initial part of the paper characterises “Pniowek” Coal Mine as far as methane and temperature hazards, which make mining operations more difficult, are concerned. Methane as a source of hazard as well as precious fuel to be used in energy-generating systems is characterised. Central air-conditioning system in “Pniowek” Coal Mine, in which chillers of 5 MWch combined power are driven by two TBG 632 V16 gas engines powered by methane coming from drainage, is presented in this paper. The obtained effects of the system and the work done up till now are estimated on the basis of the data for 2000–2002. The estimation of the energy-generating system includes: the way of building in into mining electricity and heat systems, cost of electricity, heat and chill production and the method of their utilisation, level of power demand and energy effectiveness both of the whole system and of their underground and surface parts. This work also emphasises economic effect of chill production from methane drained from coal deposits, simultaneously estimating the influence of the project on overcoming temperature hazard in headings down to level 1000 m. In the conclusion, high effectiveness of usage of methane drained from coal deposits in systems comprising airconditioning system, their high power effectiveness, utilisation of this energy source and reduction in expenditure on methane drainage thanks to profits from gas sales are discussed. Table 2 presents methane (CH4) utilisation and value of production obtained in combined electricity— cooling system and table 3 the level in which production obtained in an electricity—cooling system provided for the needs of the mine.


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KEYWORDS: Methane, temperature hazard, modern technology, airconditioning of mines, central cooling system, combined energy-cooling system

1. INTRODUCTION “Pniowek” Coal Mine is one of the five mines belonging to Jastrzebie Coal Company. As an administrative unit the mine operates in Pszczyna district in Pawlowice area (fig. 1). Coal deposit in “Pniowek” Coal Mine is geologically situated in the southwestern part of Upper-Silesian Basin in the southern inclination of the Main Basin (to the east from Jastrzebie anticline). The deposit is broken up by a series of faults of both meridian and parallel NW-SE strikes and throws from a few to several meters. The faults mentioned above are usually accompanied by smaller disturbances. Particularly intensive tectonics can be observed in a tectonic trough formed by Warszowice and Pawlowicki I faults. The whole mining area of the mine is situated between two big parallel fault zones of a regional range.

Figure 1. Administrative localisation of “Pniówek” Coal Mine “Pniowek” Coal Mine mines coal deposits located at great depths underground, characterized by a very high methane hazard and high virgin rock temperature. The necessity of improving work conditions underground and the prospect of mining deposits at 1000 m level enforced the decision of installing central air-conditioning in “Pniowek”

Methane as a source of energy in an air-conditioning system

95

Coal Mine—the first of its kind in Poland. On the basis of a study in which calculations of environmental conditions when mining deposits of “Pniowek” Coal Mine from 1999 to 2005, it was concluded that in order to keep the air temperature at the return from longwalls below permissible 28°C, it is necessary to cool the intake air. The power of air coolers that were to be installed at the longwalls’ intake sides should amount to approximately 5 MW. As a result of the research conducted at the Technical University of Science and Technology in Cracow into different kinds of central air-conditioning systems, a combined electricity-cooling system was selected (Praca zbiorowa 1998). Such a solution is unique worldwide and is based on combining an electricitygenerating system and a cooling one. Moreover, it is the first installation of a central airconditioning system in Poland, which creates new possibilities for mining at great depths below 1000 m, where the virgin rock temperature frequently exceeds 45 °C, for Polish both coal and non-coal mining. 2. CHARACTERISTICS OF METHANE AND TEMPERATURE HAZARDS IN “PNIOWEK” COAL MINE 2.1. Methane hazard Methane and explosion hazard associated with it are two of the most dangerous phenomena accompanying coal mining and most common hazards in coal deposits. Methane emission from broken coal and goaf creates a very serious hazard as far as work safety is concerned, both due to the risk of explosion and oxygen deficiency, which may result from its outflow. Many disasters resulting from the reasons mentioned above prove how serious these hazards are. Methane drainage in mines is a very important and effective method for methane prevention and is conducted in 18 Polish mines. The reduction of methane hazard thanks to this method helps to improve both work safety and uninterruptedness of machinery operation, thanks to limiting the number of machinery stoppages caused by power cut-offs resulting from exceeded permissible values of methane concentration. Effective methane drainage systems also create the possibility of obtaining methane as a natural source of energy as well as limiting the negative influence on natural environment due to methane emission to the atmosphere. Mining of coal deposits in a gassy mine requires taking special technical steps in order not to allow permissible values of gas concentration in the mine air to be exceeded. The basic method consists of using appropriate ventilation systems together with large volumes of air. Absolute methane bearing capacity in “Pniowek” Coal Mine (as of September 2003) amounts to 234,8 m3 CH4/min, out of which 88,7 m3 CH4/min, that is about 37,4% is drained to a drainage system and about 146,1 m3 CH4/min is drained together with ventilation air to the atmosphere. The volume of methane that can be removed from underground by means of ventilation is strictly restricted by regulations for combating methane hazard, which do not allow for CH4 concentration in ventilating air in underground workings to be greater than 2% by volume.


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2.2. Temperature hazard Geological structure of the minable deposits consists of Quaternary (6–80 m), Tertiary (210–850 m) and Carbon layers. “Pniowek” Coal Mine mines coal deposits in Orzesze layers (group 300) and Ruda layers (group 400) located at a considerable depth underground characterized by a very high methane hazard and high virgin temperature of rock-mass resulting from a geothermic anomaly in this area. “Pniowek” Coal Mine belongs to a group of mines exposed to a great environmental hazard. The character and intensity of heat exchange between air and rock-mass surrounding the mined deposit depends on many factors, among which the depth below surface and the virgin temperature of rock play vital roles. The environmental conditions at particular levels are as follows: at level 705 m: – temperature at shaft-bottom is 25°C, – environmental factor *K=0,91, virgin rock temperature tp=31–35°C; at level 830 m: – temperature at shaft-bottom is 27°C, – environmental factor *K=2,69, virgin rock temperature tp=36–40°C; at level 1000 m (being constructed, with deposit of 93, 8 mln tons): – temperature at shaft-bottom is 29°C, – environmental factor *K>3, virgin rock temperature tp=41–45°C. *K—environmental factor, whose value is the basis for the choice of necessary preventive measures to be taken in order to sustain proper environmental conditions in underground workings, can be generally defined as K=(tpg−td)/(td−tp), where tpg—virgin rock temperature, °C, tp—air temperature at shaft bottom of a downcast shaft, °C, td=28°C. Going deeper with mining from level 705 m to level 830 m led to an increased temperature hazard in those working places that were ventilated by air coming from level 830 m. The main method for overcoming temperature hazard is pushing large volumes of air into the areas and only when this is not sufficient, local cooling of air is used. In “Pniowek” Coal Mine in the first half of 2000 (before installing a central airconditioning system) the following local cooling equipment was used: – 7 pieces of equipment with total cooling power of 1520 KW, – 7 pieces of equipment in deposits with separate ventilation system with total cooling power of 890 KW. The total installed cooling power amounted to 2410 KW. As there was an increase in environmental hazard it was necessary either to implement local air-conditioning more widely or to use central air-conditioning. The analysis conducted showed that in order to keep production level at 14000–15000 ton/a day in


97

“Pniowek” Coal Mine, it would be necessary to use equipment with the following cooling power: – at level 830 m—5 MW of cooling power, – at level 1000 m—10 MW of cooling power.

3. METHANE AS FUEL Great volumes of methane are drained during the mining process in “Pniowek” Coal Mine. Mixtures containing 50–60% of CH4, a valuable, cheap and ecological fuel for steam, water and gas boilers, are economically utilized in installations of “Jastrzebie” Power Stock Corporation (SEJ S.A.). Methane mixtures are delivered to “Moszczenica” and “Zofiowka” Thermal-Electric Power Plants grouped in SEJ S.A., and to two TBG 632 V16 gas engines operating in a combined electricity-cooling system of a central airconditioning system. Methane mixture obtained from drainage is used for firing up boilers and is burned together with coal dust. Moreover, cyclone boilers OCG-64 in “Moszczenica” Thermal-Electric Power Plant operate only on gas, with a simultaneous considerable reduction in nominal parameters. Methane combustion, which means replacing some part of coal with cheap local fuel, is both economically viable due to the gas price as well as ecological with regard to atmospheric pollution. Before the implementation of a central air-conditioning system, methane utilization in “Pniowek” Coal Mine in 1999 was as follows:

Table 1. Methane utilization in 1999 Utilization of methane [103×m3 CH4]

Mine

“Pniówek”

Total volume of methane drained [×103 m3 CH4 per year]

52.730,1

Total volume of methane [×103 m3 CH4 per year]

33.961,8

Utilization Volume of methane drained [×103 m3 CH4 per year]

Place of utilization

2.325,6

Boilers in “JasMos” Mine

8.216,0

EC “Zofiówka”

14.181,8

EC “Moszczenica”

9.238,4

CHP “Pniówek”

Utilization ratio %

64

Failure to fully utilize all the drained methane (utilization level—64%) resulted in financial losses both due to charges for emission (about 11 thousand PLN) and lost income from its sale (about 1, 7 million PLN).


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4. DESCRIPTION OF CENTRAL AIR-CONDITIONING SYSTEM IN “PNIOWEK” COAL MINE 4.1. Background of the project When analysing the problem of air-conditioning “Pniowek” Coal Mine, the following factors were taken into consideration: availability of methane drained in the mine, its considerable surplus not fully utilized up till then and the experience gained during the implementation of the first combined power system based on methane in “Krupinski” Coal Mine in 1997. That system was based on TBG 632 V16 gas-powered engine with the nominal power of 2, 7 MWel and 3, 1 MWt made by MWM Deutz Germany Company. JCC made the decision to the implement the “Central Air-Conditioning in “Pniowek” Coal Mine” on the basis of a combined electricity-cooling system in 1998, with the following provision: – “Jastrzebie” Power Stock Corporation became responsible for realization of the part located on surface, that is gas engines, wiring systems, chillers and pipelines on surface, – Jastrzebie Coal Company “Pniowek” Coal Mine became responsible for the underground part, that is pipelines in the shaft, three-chamber feeder SIEMAG to level 853 m, network of underground distribution pipelines and air coolers. A number of different contractors were employed on the whole project, whereas the tender for the installation of gas engines and refrigerators was won by a German company—Saarberg Fernwärme GmbH (the same as in “Krupinski” Coal Mine). According to the timetable the realization was to be conducted in two stages: I stage—with realization time for June 2000: – the first TBG 632 V16 gas engine with the nominal power of 3, 2 MWel and 3, 7 MWt and cooling system with nominal power of 2, 5 MWc, II stage—with realization time for October 2000: – the second TBG 632 V16 gas engine with nominal parameters identical to the first one and cooling system with power of 2,5 MWc analogical to the one in stage I. 4.2. Description of the installation The installation located on the surface consists of two identical blocks: a gas engine, warm-water absorption chiller, hot-water absorption chiller and compression chiller, with nominal cooling power 2.5 MWch for each block. The installation underground consists of: pipelines in the shaft, three-chamber feeder of SIEMAG type and a network of pipelines together with chillers installed in working places (fig. 2). TBG 632 V16 engines TBG 632 V16 engine is a four-stroke, turbo-charged engine, with two-level mixture cooling and 16 cylinders working in V-configuration, with spark ignition. It has cubic capacity of 271, 7 dm3 and operates in Otto system during combustion with a considerable excess of air λ=1,8–2,0. Methane and air mixture from drainage in


99

“Pniowek” Coal Mine with CH4 content of 50–60% is used as fuel. The engine drives a A.Van Kaick generator which has nominal power of 3,993 kVA (actual power at cos =0,8 is equal to 3, 194), voltage 6, 3 kV and frequency 50 Hz. Conversion efficiency for electric power amounts to 38%, heat power 42%, resulting in general efficiency of the combined system of 80%. A gas engine, which is a driving unit for a cooling block, has two levels of heat recovery.

Figure 2. Simplified scheme of combined engineering-cooling system in “Pniowek” Coal Mine – warm-water cycle of nominal temperature range 86°C/72°C for heat recovery from cooling the body ofan engine, oil and mixture after turbine-charging, – hot-water cycle of nominal temperature range 125°C/100°C for heat recovery from exhaust gases. Chillers Heat recovered from an engine is used, first of all, for driving absorption chillers and during smaller demand it is directed to the heat distribution network of the mine. A cooling system co-operating with each engine: – absorption chiller in a warm-water cycle with cooling power 600 kW, using heat from cooling the engine (800 kW), cooling water from 18°C to 14,5°C, – absorption chiller in a hot-water cycle with cooling power 1730 kW, using heat from exhaust gasses (2449 kW), cooling water from 14,5°C to 4,5°C, – ammonia compression chiller using some part of the power from the generator (570 kW), cooling water from 4,5°C to 1,5°C.


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The cooling power from one installed block sums up to 2,92 MWch ensuring an adequate surplus over guranteed capacity of 2,5 MWch. Absorption chillers are lithium-bromide chillers (LiBr=H2O), where lithium bromide solution is an absorbent and water is a cooling medium, which at low pressure reaches boiling temperature at level +3,5°C in a hot-water chiller. YORK company produces both YIA HW 3B3 and YIA HW 6C4 chillers. Compression chillers are ammonia chillers with screw compressors and ammonia (NH3) as a working medium. They work in a closed-circuit system in separate chambers with a system for ammonia detection and emergency ventilation. Such a solution ensures work safety and allows for early detection and elimination of possible leakages.

Figure 3. Pipe-lines and air-coolers positioning at level 830 m Underground installation Cooling water at the temperature of 1,5°C-2°C and nominal flow of 300 m3/h is directed through a 300 mm diameter pipeline through a shaft to level 853 m and to a threechamber SIEMAG DRK 200 liquid feeder, where pressure is reduced from 9, 5 MPa to 2,0 MPa in an underground cycle and cold water pushes warm water, returning from underground at about 18°C, to the surface towards the chillers. To ensure the continuous flow of cooling water in the primary and secondary cycle, the feeder consists of three pipe chambers whose working cycles are shifted in phases by 120°. Heat loss in a feeder amounts to approximately 0,5°C and its work is monitored by a control system, which sets off an alarm in case of emergency. From SIEMAG feeder cold water runs through insulated collecting pipes 200 mm and 150 mm in diameter, and later through insulated pipes 100 in diameter, to air coolers located at longwall faces. The number of working coolers ranges from 8 to 10 and the demand for cooling power—from 360 kW to


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600 kW. Coolers are constantly moved forward with the advancing faces. An example of positioning of pipelines and air coolers is shown on figure 3. 5. EFFECTS OBTAINED BY A CENTRAL COOLING SYSTEM 5.1. Evaluation of technical results of the electricity generating of the system The first stage of the project was implemented in June 2000 and the second one in October 2000. Therefore the first year of its operation was 2001; so all kinds of analysis can be done on the basis of the results that have been obtained since then. While planning the cooling system, the following three basic criteria for the selection of the installation’s size were taken into consideration: – the mine’s demand for electricity, – gas supply from methane drainage in the mine, – possibility of the local utilization of all generated electricity. It was necessary to build in the new system into the already existing electricity and heat system in “Pniowek” Coal Mine in such a way that the obtained electricity was not a “competition” to the already existing supply, but merely economically justified supplement. In turn, after completing the installation, it was necessary to decide on the principles of mutual settlements between PSC and JCC so that to safeguard the interests of both parties to the maximum (taking into consideration external settlements with Upper Silesian Power Corporation S.A.—GZE S.A.). Two principles were therefore adopted for settling electricity, heat and cooling power supply: – settling a combined electricity-cooling system as an integral part, – cost savings thanks to the purchase of electricity and heat by a mine. Figures 4 and 5 show gas consumption and heat utilization in absorption chillers in 2002.

Figure 4. Volume of gas consumption (100% CH4) in gas engines in 2002


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Table 2 presents methane consumption and value of production in a combined electricitycooling system and table 3 presents the extent to which production in a electricity-cooling system covers the demand of the mine.

Table 2. Methane utilisation and value of production of combined energy-cooling system Year No. 1.

Specification Working hours*

Unit h

2000 and 2001

2002

21,663

15,975

th. m

19,224

12,906

%

57

56

[MWh]

65,950

45,135

3

2.

Consumption of fuel (CH4)

3.

Average concentration of CH4

4.

Electricity production

5.

Heat production

[GJ]

215,213

139,971

6.

Cool production

[MWh]

74,998

29,280

Figure 5. Power obtained from gas engines in 2002


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Table 2. Methane utilisation and value of production of combined energy-cooling system Year No. 1.

Specification

Unit

Working hours*

2000 and 2001

2002

21,663

15,975

th. m

19,224

12,906

%

57

56

[MWh]

65,950

45,135

h 3

2.

Consumption of fuel (CH4)

3.

Average concentration of CH4

4.

Electricity production

5.

Heat production

[GJ]

215,213

139,971

6.

Cool production

[MWh]

74,998

29,280

* total time of work of engines no. 1 and no. 2.

Table 3. Supplying mining needs by production from combined system 2000 No.

Specification

Value

2001

% of demand

Value

2002

% of demand

Value

% of demand

199,120

17,7

Electric Power [MWh] 1.

—Total consumption —roduction by combined system

195,485 11,865

196,397 6,1

40,374

20,6

35,160

Heat [GJ] 2.

3.

—Total consumption

148,154

—production by combined system

3,644

Cooling Energy [MWh]

36,580

178,475 2,5 100

42,957 38,418

166,441 24,1 100

23,434 29,280

14,1 100

For a fuller understanding of the gains obtained by “Pniowek” Coal Mine due to electricity purchase from the system, which is cheaper than electricity from GZE S.A. network, the following conditions must be remembered: – electric power is introduced into the system of “Pniowek” Coal Mine at a voltage of 6 kV and is distributed according to A23 GZE S.A. tariff (two-unit, three-zone tariff), – electricity price from SEJ S.A. is the maximum price (one-unit, all year round tariff), – unit price of cooling power is decided upon after negotiations between JSW S.A. and SEJ S.A. and is based on its actual technical production costs.


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Table 4 show prices in the first half of 2003.

Table 4. Quantity & Unit price (including VAT) of electricity purchase in “Pniowek” Coal Mine Year/Month

2003

GZE S.A.

Combined system

Total

Quantity [MWh]

Price [zł/MWh]

Quantity [MWh]

Price [zł/MWh]

Quantity [MWh]

Price [zł/MWh]

I

15,573

242

3,091

178

19,248

231

II

12,003

246

2,909

178

16,170

231

III

13,201

249

2,938

178

16,592

235

IV

12,130

234

2,799

175

17,228

219

V

12,094

232

2,867

175

15,796

220

VI

12,060

235

2,545

176

15,447

224

77,061

240

17,149

177

100,481

226

I–VI 2003

The method of settlement used allowed to obtain a relatively low price of cooling power (157, 10 PLN/MWh excl. VAT in 2003) and at the same time to generate funds enabling the quick process of paying back the loans taken by SEJ S.A. for the project’s implementation. Gas engines with generators and heat recovery Gas engines are supplied with methane and air mixture from drainage, at a rate of approximately 50 m3/min, which at average concentration CH4 of 55% means 27, 5 m3/min of methane and about 8442 kW of chemical power in fuel. Total efficiency equals to about 80%, out of which 38% is for electric power and 42% for heat. Taking into consideration the fact that for modern condensation-type power plants’ efficiency averages approximately 36–40% and where gas-steam systems are used—maximum 60%, efficiency of about 80% obtained in a combined power system decisively surpasses other solutions. Full utilization of heat obtained in a combined system both in summer and winter as well as the possibility of working without a drop in efficiency during partial only heat utilization in absorption chillers (reduced cooling power) are of great importance. In such a case excess heat is directed to a heat system in a mine. In other combined electricity and heat generating systems the lack of heat supply in summer leads to the necessity of its release into environment, and at the same time to a drop in its efficiency. 5.2. Evaluation of technical results of a cooling part of the system Chillers Effective and efficient combining absorption and compression chillers in a system to produce cooling power are a technical solution, which has great advantages.


105

Absorption chillers utilize low-temperature, the so-called “waste” heat obtained from cooling an engine and exhaust gases throughout the whole year, while the maximum demand for power is in summer, when cooling power is close to its nominal value, that is 5 MWc. Thanks to such a solution about 80% of the total cooling power is generated thermally, which greatly affects a unit cost of generating 1 MWch of cooling power. The availability of soft heat and its low cost (thanks to a combined system) were the basis for applying this both technically and economically effective solution. The lowest temperature of a cooling medium that can be obtained by using absorption chillers is about +4,5°C, therefore in the system used in “Pniowek” Coal Mine it was necessary to apply a final unit for compression chillers in order to reach the final temperature of +1,5°C. Compression chillers utilize some part of the electricity produced by a generator (1140 kW), limiting the possibility of its sale to “Pniowek” Coal Mine. However, in the cost balance of generating cooling power, this solution is far cheaper than selling the generated electricity and then using electricity bought from Upper-Silesian Power Station to run the compressors. The indicators of cooling efficiency of particular elements of a cooling system are as follows: – warm-water chiller ε=0,848, – hot-water chiller ε=0,709, – compression chiller ε=3,56. Flux of heat equal to 100 kW, supplied to the desorber of a warm-water absorption chiller, allows to obtain 84,8 kW of cooling power. Similarly, a flux of 100 kW in heat supplied to a hot-water absorption chiller enables reaching cooling power of 70,9 kW. Electric drive of a compression chiller of 100 kW power results in cooling power of 356 kW. If constant voltage and hot water temperature in the first absorption chiller are maintained as well as constant voltage and cold water temperature in the second, the obtained cooling power of the system depends only on temperature of return water from the secondary cycle in a mine. The change of cooling power of a chiller is dependent on the graph of cooling efficiency as a function of evaporation temperature of a working medium. Underground cooling installation In “Pniowek” Coal Mine the main aim is to cool water to the required parameters and to supply it further to air coolers in longwalls to level 830 m (and also level 1000 m). Water temperature, when entering a cooling unit, ranges from 14°C to 18°C (on average 17°C) due to heat pickup on the way. Therefore in summer the cooling power obtained ranges from 4 to 5 MWc. The obtained up till now experience from using a central airconditioning system shows that the maximum cooling of a cooling medium takes place during the movement of both blocks but during the movement of one block cooling power is reduced with regard to nominal power. Power balance of an underground cooling system is based on heat exchange efficiency of installed air coolers. Energy balance in heat exchangers through which humid air flows requires taking into consideration thermodynamic properties of air. Efficiency of an air cooler depends, first of all, both on the construction and size of a heat exchanger. Moreover it depends on the


106

volume of cooling water, rate of airflow, differences in their temperatures and air humidity. An increase in the difference of temperature between feed and return water as well as an increase in the volume of water results in a better utilization of cooling power of air coolers. Figure 6 illustrates this dependence with regard to the total cooling power of an underground part of the installation (based on the results of measurements done in 2001). The dry-bulb air temperature at the entry to a cooler, relative humidity of air entering the cooler and the volume of air flowing along the longwall face where the cooler is located have significant influence on its cooling power. It increases together with an increase in dry-bulb temperature and relative humidity of the air (figures 7 and 8). In the period under consideration, that is in 2001, actual cooling power of installed coolers ranged from 200 kW for longwall S—3 to 400 kW for longwall B—1. Nevertheless, most coolers reached the power ranging from 200 to 300 kW. Average cooling power obtained during the use of SIEMAG exchanger in 2001 amounted to 2965 kW. In summer cooling power peaked at 5000 kW (with the mass of water circulated of 256 Mg/h). Heat losses, mainly in the return, not insulated, pipelines amounted to 1095 kW. Therefore, the efficiency of the underground cooling system amounted to 63%, which resulted mainly from a smaller than forecasted volume of cooling water as well as air temperature at coolers’ intake side. This temperature was frequently lower than 28°C.

Figure 6. Effect of the difference between feed and return water temperatures on cooling power


Figure 7. Effect of dry-bulb temperature of air entering coolers on cooling power

Figure 8. Effect of relative humidity of air entering coolers on cooling power

107


108

Figure 9. Effect of airflow rate on cooling power However when there is an increase in volume of air the effect of cooling is smaller (fig. 9). In the period under consideration, that is in 2001, actual cooling power of installed coolers ranged from 200 kW for longwall S-3 to 400 kW for longwall B-1. Nevertheless, most coolers reached the power ranging from 200 to 300 kW. Average cooling power obtained during the use of SIEMAG exchanger in 2001 amounted to 2965 kW. In summer cooling power peaked at 5000 kW (with the mass of water circulated of 256 Mg/h). Heat losses, mainly in the return, not insulated, pipelines amounted to 1095 kW. Therefore, the efficiency of the underground cooling system amounted to 63%, which resulted mainly from a smaller than forecasted volume of cooling water as well as air temperature at coolers’ intake side. This temperature was frequently lower than 28°C. 5.3. Evaluation of economic effectiveness of a central cooling system The concept and plan of a central cooling system in “Pniowek” was developed at the Technical University of Science and Technology in Cracow. Jastrzebie Coal Company S.A. covered the cost of installation of the underground part of a combined electricitycooling system in “Pniowek” Coal Mine. “Jastrzebie” Power Company S.A. financed the installation, and became the owner of the surface part of the infrastructure. The capital money brought forward by “Pniowek” Coal Mine and SEJ S.A, subsidies and bank loans were the source of financing of this undertaking. The costs incurred during the installation of the central air-conditioning system in “Pniowek” Coal Mine from 1998 to 2000 amounted to:


109

– Jastrzebie Coal Company; “Pniowek” Coal Mine

– 20728,3 thousand PLN

– Jastrzebie Power Company

– 33063,5 thousand PLN Total


The analysis of economic effectiveness of the investment showed its high profitability. The economic indicators characterizing effectiveness are as follows: NPV>0 +202.399,4 thousand PLN IRR (NPV=0) 40% The time of return on investment calculated from the time of implementation of both stages that is from 01-01-2001 is 3 and half years. In 2002 in “Pniowek” Coal Mine one ton of coal sold was encumbered by the cost ofcentral air-conditioning to the amount of 2,40 PLN, which was equal to 1,55% of unit cost of coal production. Taking into consideration the revenues from methane sale and benefits resulting from the purchase of cheaper electric power and heat, the profit from implementing central air-conditioning amounted to 1,47 PLN/ton. 6. CONCLUSIONS 1. Central air-conditioning system in “Pniowek” Coal Mine, realized on the basis of a combined electricity-cooling system, using methane drained from a mined deposit as fuel, is a unique solution in the world. 2. Two gas-powered engines serve as a driving unit of the system. In a combined system they generate 6,4 MWel and 7,4 MWt of heat, which is the main power source for absorption chillers, although compression chillers also form a part of the cooling system with the total power output of 5 MWc. The cooling power obtained is used for cooling air in longwall face at level 830 m (as well as at level 1000 m). Water at a of temperature 1,5°C is fed through an insulated pipeline to level 830 m, where in a three-chamber SIEMAG feeder reduction in water pressure and heat exchange with water from underground coolers take place. Pressure reduction in a shaft cold-water pipeline by means of a three-chamber liquid feeder guarantees high percentage of potential-energy power recovery during transport of water in a shaft. Such a solution is more efficient than the application of a heat exchanger or a system with Pelton turbine. Air-cooling in a longwall face is realized by placing a set of coolers connected in parallel before the entry to the longwall at a distance not greater than 200 m. These coolers are moved forward as longwalls advance. 3. Total efficiency of the power system is 80% (38% for electric power and 42% for heat) and is decisively higher than the efficiency of modern condensation-type power plant (36–40%) and gas-steam systems (up to 60%), which makes it one of the most efficient power systems. Full utilization of heat generated in a combined system both in summer and winter as well as the possibility of work without a drop in the system’s efficiency, even when absorption chillers are not loaded to their full capacity, are of great importance.


110

4. Implementation of an air-conditioning system in longwalls of “Pniowek” Coal Mine resulted in a considerable improvement in environmental conditions in these longwalls, but with the exception of summertime, the efficiency of air coolers is lower than nominal because of the different than anticipated temperature and humidity of intake air. In some air-conditioned longwalls temperature of 28°C is exceeded; therefore it is necessary to undertake some steps to implement longwall air coolers located in those places. 5. The costs incurred for the installation of the central air-conditioning system in “Pniowek” Coal Mine from 1998 to 2000 amounted to: – Jastrzebie Coal Company; “Pniowek” Coal Mine


– Jastrzebie Power Company

– 33063,5 thousand PLN Total


The analysis of economic effectiveness of the investment showed its high profitability. The economic indicators characterizing effectiveness are as follows: NPV>0 +202.399,4 thousand PLN (NPV=0) 40% The time of return on investment calculated from the time of implementation of both stages that is from 01–01–2001 is 3 and a half years. 6. In 2002 in “Pniowek” Coal Mine one ton of coal sold was encumbered by the cost of central air-conditioning to the amount of 2,40 PLN, which was equal to 1,55% of unit cost of coal production. Taking into consideration the revenues from methane sale and benefits resulting from the purchase of cheaper electric power and heat, the profit from implementing central air-conditioning amounts to 1,47 PLN/ ton. 7. The mine’s power requirements are satisfied to a considerable extent by the power generated in a combined power system based on gas engines and amount to: – electric power 18–21%, – heat 14–24%. which means a great reduction in expenses on purchase of power by the mine. 8. Implementing a central air-conditioning system will enable the mine to operate for 20 more years because it makes it possible to obtain at longwall faces such environmental conditions that comply with the legal regulations. Moreover, improvement in environmental conditions leads to an increase in efficiency and improvement in work safety during mining operations. REFERENCES Berger J., Nowak E. 1999: Pozyskiwanie metanu metodami wiertniczymi z wyrobisk podziemnych i z powierzchni. Wiadomości Górnicze, nr 2. Kozłowski B., Grębski Z. 1982: Odmetanowanie górotworu w kopalniach. Wyd. Śląsk, Katowice. Nawrat St, Gatnar K. 1999: Ujęcie i gospodarcze wykorzystanie metanu z pokładów węgla z obszaru górniczego Jastrzębskiej Spółki Węglowej S.A. Wiadomości Górnicze, nr 1.


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Praca zbiorowa 1998: Badania nad doborem technologii poprawy warunków klimatycznych w wyrobiskach górniczych w oparciu o klimatyzację centralną, AGH, Krakow, 1998. Roszkowski J., Szlązak N. 1999: Wybrane problemy odmetanowania kopalń węgla kamiennego. Uczelniane Wydawnictwa Naukowo-Dydaktyczne, Krakow. Skorek J., Kalina J. 2000: Skojarzona produkcja ciepła i energii elektrycznej zintegrowana ze zgazowaniem biomasy. Nowoczesne gazownictwo nr 1/2002. Szlązak N., Nawrat S., Jakubów A. 2000: Pierwsza klimatyzacja centralna w KWK “Pniówek” Jastrzębskiej Spółki Węglowej S.A.. Przegląd Górniczy, nr 10, 2000. Szlązak N., Tor A., Jakubów A. 2001: Możliwości ograniczenia emisji metanu do atmosfery w kopalniach Jastrzębskiej Spółki Węglowej S.A. Materiały konferencyjne “Geosfera”—Krakow. Szlązak J., Szlązak N., Obracaj D., Borowski M. 2001: Wykorzystanie ciepła odpadowego w skojarzonym układzie energetyczno-chłodniczym. Kwartalnik AGH Górnictwo, rok 25, z. 2, Krakow 2001. Szląjzak N., Obracaj D., Borowski M. 2001: Przykład wykorzystania chłodnic absorpcyjnych w skojarzonym układzie energetyczno-chłodniczym. Technika Chłodnicza i Klimatyzacyjna, Wydawnictwo “Masta”, 6–7/2001.

New Technologies of Coal Bed Methane Preliminary Recovery

S.V.Slastounov Chair Engineering Protection of the Environment, Moscow State Mining University, Russia International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 For about 40 years the main coal mining regions of the former Soviet Union have been engaged in this scientific research. The basic technology is to hydraulically fracture coal beds with water and following this recovers gas and water. The depth of coal bed occurrence of more than 500–600 m requires a new technology of methane recovery from coal beds that are not distressed, based on active physical, chemical, thermodynamic and other effects, which are protected by Russian patents and have no analogies in the world practice. The paper presents information about the theory and mechanisms of the new methods and summarizes results of their tests in Karaganda Coal Basin. Their economical efficiency has been proved. Experimental investigations of a coal bed as a subject of active effects were carried out. Some experimental data was obtained, related to basic properties of coal bed. Parameters such as natural porosity and fissility as well as their change under active effects were explored. Thermodynamic properties of the system “coal—methane—working agent” were investigated. Optimal working agents, the criteria being their thermodynamic stability, were chosen. New active methods allow eliminating the main disadvantages of the traditional technology of hydraulic fracture by water and to: – increase uniformity of treatment, – increase permeability of coal bed for methane, – increase rate of water and gas recovery, etc.

As a result of implementation of the new technologies the following were achieved: – 50–60 per cent decrease in gas-bearing capacity of treated coal beds,

New technologies of coal bed methane preliminary recovery

113

– 70–80 per cent decrease of volume of gas in mines, – 30–40 per cent decrease in volume of development work and increase of safety of miner’s work in areas of intensive gas drainage. Coal mines are kind of gas traps. When the inflow of deep-earth gas exceeds its outflow to the surface a rather high gas-bearing capacity of the stratum is observed. This may lead to formation of abnormally high pressure of gas in the coal seam, similar to the one observed in oil and gas deposits. The first priority of forced gas recovery from coal deposits by engineering facilities was to enable normal operations in gassy mines. Taking into consideration high purity of methane in coal deposits it seems to be reasonable and beneficial to use methane to manufacture, for instance, protein, high-quality dyes or motor fuel. Such utilization of methane is hindered or impossible in case of its contamination by nitrogen, oxygen or carbon dioxide, unavoidable with conventional gas recovery from mine openings. Moreover, gas extraction operations may hinder mining operations and be a source of additional hazard in mines. Under the direction of academician A.A.Skochinsky the principles of separating mining and gas extraction operations, preliminary drainage of coal-bearing strata by boreholes from the surface, and increase of methane recovery by means of controlled processes were developed in Moscow State Mining University between 1957 and 1961. In the past years these principles in different respects were confirmed in practice. In most cases high efficiency of gas recovery, decrease of gas-bearing capacity and dustiness, outbursts and blowouts were achieved. However, preliminary drainage with following it methane recovery is not widely applied. The reasons are as follows: – the time of preliminary work required to reach practical and perceptible results is three to seven years, – low control of the process, – myth about high cost of work, vigorously developed and maintained by the advocates of “old” traditional method of gas recovery from underground openings. In recent years this myth is gradually disappearing because of realization of what Russian scientists predicted would happen—importing of our procedures from abroad. This paper describes some results we obtained in the last 8–10/eight-ten/years. The following postulates are of fundamental importance to the analysis of borehole yields as they are manifestation of the law of phase permeability for water and gas filtration in coal bed and replacement of methane by water in sorption volume of coal. While boreholes were drilled from the surface it was discovered that after the hydrofracture of coal beds, gas yield change in time could be characterized by two functions (fig. 1). An increase in borehole’s productivity during the process of drilling (curve 1) is explained by gradual release of main joints and filtrational volume by water and creation of favorable conditions for gas outflow. It is important to put injection wells quickly into operation for maximum gas yield. It is possible to shorten the time of holding the liquid in bed by pneumatically pressing the working fluid away immediately after hydro-fracture. This will increase the rate of fluid transition from fractured joints to filtrational pore volume. Besides, when the volumes of water being pumped-in are small, the necessary time of holding the liquid in bed decreases.


114

Distribution of water in coal bed around the injection well is not uniform. Full saturation takes place in the center of the treatment zone and natural on the perimeter. Consequently an additional action to the coal bed in the form of pneumatic pressing away seems quite logical. This allows some portion of fluid to be moved to peripheral section of the treatment zone. On the other hand, considering available data and the model of mass transfer it was noted, that the longer water was held in bed the less fuel was removed from the borehole, therefore the more quantity of water was absorbed into volume of coal and, accordingly, the more useful work it did for methane replacement. It makes possible indirect evaluation of hydro-fracture efficiency: the greater the volume of liquid that self-outflows the less gas is recovered from the borehole. Prevention of the process of self-out-flowing and boreholes exploitation with the liquid remaining in the coal bed is the way to increase hydro-fracture efficiency. Currently, the problems of coal bed methane could be in general divided into three: – safety problems, – ecological problems, – energy problems. These problems are interconnected. Thus, mine methane utilization i.e. in boilers solves not only energy problem but also ecological one as well by reduction of CO2 emissions, elimination of dust emission, etc. The last stage of research is devoted to improvement of the ways of preliminary methane recovery from coal beds, which goes in the following basic directions: – improvement of the technology of exposure in regime of cavitations, which includes selection and development of working substances and regimes of exposure in order to put into realization the mechanism of self-supported destruction of coal in a wide range of mining and geological conditions, – hydraulic action by foaming agents. The essence of the method is to pump certain salts and acids into the bed in definite regime and provision of thermo-dynamical conditions which favor rapid emission of carbon dioxide, – step-by-step hydro-dynamic treatment of coal beds which result in coal swelling during the hydro-action, – coal bed treatment by hydraulic impact.


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Figure 1. Yields (q) of bore holes, drilled from the day surface while hydro-fracturing of coal beds. 1— development of bore holes pumpingout of water at short time of holding (2–4 month), 2—development of bore holes after holding the liquid in bed during 9 month and over Sufficiently high power-intensiveness of the process of fracture is defined according to the postulate, that fracture opening in a stressed gas-filled seam is possible only by supplying pressure overcoming the force of mining stress and deforming process during fracture formation. In these conditions the use of the energy of the gas and mining stress forces contained in coal bed for coal destruction is of great interest and can provide conditions for the fractures staying open for the exposure of partial de-stressing of the coal bed. Technologically the process of initiation of self-supported coal destruction is realized by steadily pumping working liquid into the seam and then following with a sharp release of pressure. As a result destruction and coal and gas outburst from the seam occurs. Regime of power action and regime of coal dust output are repeated over and over again. The created extended slot-like space forms around itself in a coal bed a destressed zone of high jointing and gas permeability, where the fracture opening caused by de-stressing provides a sharp increase in gas permeability of coal and partly destructed coal because formation of new degassing spaces increases a rate of gas emission desorbed from the coal. It is necessary to accentuate, that the zone of low stresses develops with time. Development of the technology of pneumo-hydro-fracture of a coal bed in regime of cavitations and optimisation of the method parameters have been carried out in the following directions:


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– choice of the volume and rate of working agent (water and air) pumping, – elaboration of parameters of working agents employed for realization of the effect of coal outburst into the well, – elaboration of sequence of technological operations during the effect on a coal bed. Combination of the process of hydro-fracture with pneumatic treatment of a man-made collector by compressed air allows providing: – intensification of absorption of water in coal and substitution of methane in the coal bed, – cleaning of main system of fractures from coal dust with a small range of particle dimensions, – coal bed warming up. Investigations of the process of initiation of coal outburst from the well during pumping in of the working agent and its setting on (regime of cavitations) have been carried out on a series of wells in the mining field named after “V.I.Lenin” and at “Kazakhstanskaya” mine. In some cases dozens of coal outbursts into one well were initiated. The weight of coal extracted after the outburst reached 3–5 t. It was established that coal outburst from the nearby area into a man-made collector in the zone of preliminary degassing preparation is provoked by pumping in a portion of liquid at a rate of 60–80 l/s in the volume larger than 800 m3 and applying the pressure onto the well collar of 9,0–11,5 MPa. It is necessary to drop the pressure after a portion of water is pumped in. Intensification of the methane recovery during this method is 2 times higher than an average statistical intensification of methane recovery from the wells treated in accordance with traditional technology. Fracture by foaming agents allows to: – intensify by 1,3–1,5 times an active effect of periphery part of the zone of hydraulic fracturing on account of real velocities of putting into practice of the foaming agents, – increase uniformity and quality of treatment on account of greater penetrating ability of the foaming agent with low viscosity into tiny pores and fractures of a coal bed, – intensify removal of the working agent from fractures induced by hydro-fracturing by gas. Working agents compositions were defined during laboratory tests. The main criteria were the volume and intensity of gas emission during the chemical reaction between the salt and acid and opportunity of their use on the exposure of further coal reserves exploitation. The result of the tests was a technological process that provoked the rise of rate of pumping in to 12–15 l/s and pressure of fracturing to 2,0–3,0 MPa. Now development of experimental wells has begun, the results of which will allow evaluating effectiveness of the technology. Further plans concern the next experiment with parameters, provided the increase of pumping-in rate to 30–40 l/s, which allows to extent the radius of action. Multi-stage treatment of a coal bed is based on coal swelling while increasing its moisture content in conditions of free volume. To extent radius of the action it is necessary to increase viscosity and pumping-in rate. Pumping-in rate is limited by parameters of the pumping plant. It is also necessary to take into account the expediency


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to minimize the volumes of liquid pumped-in. One of the directions of active action technology development is the use of multi-stage treatment of a coal bed. As a result of coal swelling when working agent remains in situ temporary reduction of permeability occurs. Further treatment after the process of swelling but before working liquid absorption in fact leads to extension of specific rate of pumping-in. In the laboratory conditions an intensive swelling of the samples lasts 2–3 days. In conditions of a coal bed the duration of this process should be estimated as 10–15 days. In time working agent gets absorbed from cleats into the volume of coal and partly into surrounding rock. According to recommendations and a project while treatment of the well No. 28 a following sequence of operations was produced: – cavity formation at the weak part of a seam, – pumping of 200 m3 of water at a rate of 72 l/s, – intermediate working agent left for 15 days for realization of the process of seam swelling, – realization of the second stage with pumping of 2950 m3 of water at a rate of 80 l/s, – working agent left for 8–9 months for realization of the effect of methane substitution, – exploitation of the well with working agent recovery from man-made collector by machine bell-crank. Preliminary results show the increase of gas emission by 20–30% in comparison with traditional technology. The technologies of degassing coal seams using water hammer effect and redistribution of rock stress in a borehole area close to a working face have been developed and tested in situ. Basic technology includes coal seams being exposed to water injection under 100–150 atm. pressures. Under the technology suggested, at the hole water discharge stage in predesigned conditions, water hammer uses cyclic water flow closures to do its job. It causes impulses of pressure of 170–200 atm. amplitude. Another technological variation has been developed to change rock stress in hole surrounding area due to the slots forming a helical line. Coal block developed tangential stress, which increases its permeability. The above technologies ensure 1,5–2 times increase in coal methane discharge. Elaboration of methods of predicting zones of heightened gas emission is a subject of great interest both from the point of view of intensification of gas liberation through holes as the safety of mining operations. Predicting the zones of heightened gas emission is realized by the following methods: 1) geodynamic zoning, 2) predicting zones of heightened jointing based on analysis of structure and composition of surrounding rock layers, 3) analysis of stress deformation of coal-and-rock-mass. At present methane recovered during preliminary degassing at mine “Khazakhstanskaya” is burnt up in flares. This helps reducing emission of greenhouse gasses into the atmosphere. 512 m3 of CH4 is an equivalent of 7,16 t of CO2 should it be discharged into the atmosphere, but produces only 1 ton of CO2 during burning. Therefore burning of each 1000 m3 of methane is equivalent to the reduction in CO2 emissions by 12 t, and burning in flares 20 m3/min of methane gives an annual reduction of effluents of 126


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thousand ton of CO2. In the past it was utilized in boilers. Experience of exploitation of coal reserves at Karaganda Coal Basin shows that zones of preliminary degassing reduction of gas bearing capacity correspond to outputs of preliminary recovery. This allows considering methane recovered at the fields of working mining enterprises as mining methane. Appropriate works are recommended to include in the rank of prospective during elaboration of mechanism of trade of quotes of greenhouse gases.

Geomechanical Problems in Simultaneous Exploitation Both Open Pit and Underground at Minera Michilla, II Region Antofagsta— Chile

Alfonso R.Carvajal Mining Department of La Serena University. Chile Claudio G.Fernández & Jaime M.Carmona Rock Mechanic Staff Minera Michilla S.A.Chile International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 1. INTRODUCTION In Chile, the most important copper ore deposits have been exploited under both open pit and underground methods. A number of them (Chuquicamata, El Romeral, Carmen de Andacollo, etc.) are in the final step of exploitation by any of the methods above mentioned. On the other hand, new deeper geological resources have required a shift from open pit to underground mining because stripping ratio have raised to uneconomical level, involving different mining problems. Other deposits have evolved in the opposite way, that is, from underground mining to open pit due to new shallow geological resources, which were not previously recognized for different reasons. Other deposits are exploited simultaneously by open pit and underground mining. This work presents problems related to this situation in Minera Michilla S.A. with an emphasis on geotechnical management as an important part of the mining business. 2. BACKGROUND The Michilla mining district is located 112 km north of Antofagasta in the Second Region (Chile), at 850 m above sea level. The mining holding involves about 450 km2 stretching from 22°39” to 23°00” south latitude and 70°04” to 70°17” west longitude (figure 2.1). The origin of the Minera Michilla S.A. is the Compañia Minera Carolina de Michilla, founded in 1959 with the purpose of exploiting the mines located in the Michilla Mining

Geomechanical problems in simultaneous exploitation

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district, in the Antofagasta province. In 1980, the major part of the mining holding was acquired by CORFO, the State Production-Promoting Corporation. In 1989 the enterprise participated in the formation of Compania Minera Lince Ltd., a joint venture with the Outokumpu and Chemical Bank group with the purpose of exploiting the Lince Project. Two years later, the stockholders linked to the Michilla mining holding acquired the Outokumpu and Chemical Bank participation. After this business operation, Compania Minera Michilla and Compania Minera Lince merged into Minera Michilla S.A.

Figure 2.1. Location of Minera Michilla S.A. 3. GEOLOGY The geology of the Michilla Mining District includes volcanic rocks of La Negra formation, intrusive rocks of the Coastal Batholite and sedimentary rocks. The volcanic rocks belong to the Jurasic Period (185–140 millions years ago); this geological unit with 60% of the outcropping is the most important in the district, presents an monoclinal attitude and is formed by two stratigraphical levels: the lower level, which stretches from Quebrada de Mejillones in the south to Quebrada de Guala Guala in the north. Its 2,000 m thickness presents aphanitic and porphydic andesite levels with shades outwardly black, green and red. The volcanic breccias with reddish brown vesicular clasts appear in an aphanitic and porphydic andesite matrix. Tuffs and tobaseous sandstone also outcrop. The upper level, spreads from Quebrada de Guala Guala to the northern border of the district. Its 9,500 m thickness contains andesitic volcanic levels with in minor tuffs. The intrusives, which belong to the Upper Jurasic and Lower Cretacic Periods (140–100 million years ago), are batholitic bodies and stocks of gabro-dioritic to monzonitic andesitic composition. The intrusive stocks are semicircular with restricted outcrop and with marked structural control, generating haloes of local metamorphism around them, which are related to the copper mineralization. The crops are tabular bodies located throughout the structures, usually cutting the largest stocks.


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The unconsolidated sedimentary deposits represent 30% of the district area and correspond to continental gravel from the Oligocen-Miocen Period, sea deposit of terrace from the Upper Pliocen, deposits of the Mejillones formation coming from the PleitocenHolocen Period and aluvial deposits and debris from the Pleistocen-Holocen Period. The Michilla district is characterized by its copper deposits that correspond to a group of deposits with similar characteristics located in the coastal zone (Mantos Blancos, Buena Esperanza, Ivan Zar and the Michilla District). The mineralization of the Michilla district corresponds to stratabound orebodies and, in a minor proportion, to vein deposits related to regional structures and stock-type intrusive bodies. These deposits had been grouped in two belts: the western belt, which most frequently includes vein deposits (Carolina, California, Puerto Arturo, etc) and irregular breccious pockets (Graebe, Cerro Pardo, etc.). The eastern belt is mainly formed by stratabound orebodies (Lince, Buena Vista, Polos), with evidence of hydrothermal breccias genetically connected with the strata (Estafania). The copper mineralization is present in sulfides such as chalcolcite most of the time, and to a lesser extent, in covelline, bornite, chalcopirite and pirite. The oxidised ores are composed of atacamite, crisocole and small amounts of malaquite and cuprite. Figure 3 shows the general geology of the Minera Michilla holding.

Figure 3.1. Geology of the Minera Michilla


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4. STRUCTURAL CHARACTERIZATION Applying traditional methods in the gathering of field data in both open pit and underground areas, it is possible to say that the Lince open pit is constituted by a stratigraphic sequence of volcanic breccias, interspersed by volcanic to subvolcanic andesites. Probably many of these bodies correspond to strata, some of these strata are very narrow (1 to 2 m) to be classified as lava flows and, on other hand, it is difficult to find typical surface flow of the volcanic lava, however, tabular jointing typical of dykes and veins can be observed. The minor structures were defined as those that cut at least a bench in the open pit whereas intermediate structures are those that cut at least a section between ramps or have been defined because of their importance. The major structures correspond to faults or faulty dykes with respect to their own identification. On the structural map of the district it is possible to see the structural system with the Mititus Fault prevailing with a subvertical NS strike, which is believed to be directly associated to the district mineralization. The open pit mine as well as the underground one present structural sets corresponding to this orientation. 4.1. Major and Intermediate Structures A series of major and intermediate structures were recognised in the Lince open pit and the underground mines during the surveying process. 4.1.1. Carmen Fault This fault appears in the Lince open pit as well as in the underground mine. In the open pit, this structure is continuous and in the west wall is formed by a reddish gouge zone of approximately 1 m thickness as shown in figure 4.1. In the underground mine, this fault is cut by structures with an approximately N-S strike and subvertical dip with throws of less than 5 m. This structure is constituted by fractured rock zones up to 5 m thick, reddish brown fault gouge 20 to 100 cm thick, and parallel intermediate structures appearing in some sectors. In the westernmost outcrup, the Carmen fault presents highly fractured rocks, 5 m thick, and saalband in the boundaries. 4.1.2. Muelle Fault This structure, with a NEE strike and 50° to 70° dip to the south, is observed in the northern wall of the open pit. It shows a crop of 50 to 100 cm thick associated to it and formed by fault breccias and fault saalband. Other structures associated to Muelle fault correspond to Frontera and Continuidad Fault, which will be analysed farther on.


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Figure 4.1. Carmen Fault view in west wall in open pit, phase 7 4.1.3. Ana Rosa Dyke This structure, 2 to 3 m thick, runs through the whole open pit in the central zone. In the western and central sections, this dyke has a NEE orientation with 50° to 65° dip and returns running parallel the to Carmen Fault. Figure 4.3 shows the crop in the eastern sector of the open pit.

Figure 4.3. Ana Rosa Dike


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4.1.4. Faulty Sandra Dike This crop is located in the western wall of the open pit, central section of phase 7 of exploitation, finishing at the Ana Rosa Dike. This structure has a NNE strike and presents a 50° to 70° dip to the south, with a thickness varying from 3 to 5 m in sheared rocks. 4.1.5. Stratification The planes of stratification appear throughout the open pit but only the sections to the south of Carmen Fault are classified as important. This is so because the orientation of these structures is more favourable in the southern part of the pit, while in the southeast and southwest, wedges with other structures are formed. Weak evidence of failures in these joints are observed, usually cut and displaced by other structures, which turn them stepped. These planes are very rough with a NEE orientation and 30° to 40° dip to the NW. 4.1.6. Continuidad or Terminal Fault This fault presents a NE strike with a 50°–60° dip. It is found in phase 7 of Lince open pit and the Susana pit. 4.1.7. Frontera Fault This structure is parallel to Muelle Fault and is a part of the same system. See figure 4.4.

Figure 4.4. Frontera, Muelle and Carmen faults in open pit, phase 7


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4.2. Structural Domains Eight structural domains were defined in the open pit. These sections were defined based on major and intermediate structures. In the northern part, a more detailed analysis was carried out because of the larger amount of outcrops available. In table 4.1 the summary of the structural sets per domain is presented.

Table 4.1. Structural System Set 1

Structure S.Muelle

Domain Domain Domain Domain Domain Domain Domain Domain 1 2 3 4a 4b 5 6 7 52/166

43/171

58/169

11 S.Carmen 2 3

S.Muelle

S.Muelle

5

S.Carmen

54/137

7

56/143

59/136

80/178 66/144 78/206

62/146 63/220

56/142 79/212

50/043

71/022

68/190

6 61

57/162

80/168

S.Juana

4

55/161

54/182 72/088

56/186 75/085

75/100

51/184 75/091

85/272

S.Mititus

82/122

79/115

71

84/298

8

S.Continuidad

9

Stratification

55/134 38/331

35/335

10 F.Sandra 12 S.NW

80/253

75/064

86/061

74/235

78/062

79/241

5. GEOTECHNICAL CHARACTERIZATION A general geotechnical classification of the rocks in the area and a description of the geotechnical features of the structures are carried out below. 5.1. Rock-Mass This volcanic sequence is constituted by breccias and andesites with good to bad quality, which establish an insertion of different qualities. The GSI system was used to define the quality of rocks. Table 5.1 shows a summary of the classification. The breccias are generally of good quality, presenting a few rough wavy structures with low mechanical alteration and with a strong rock matrix. These rocks are stable in the slopes and are only affected by major structures. They are limited by stratifications.


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Table 5.1. GSI values of Rajo Lince rocks GSI

Lithology

Structure

Max. Middle Min.

Structural Condition

Intrusive

54

46

37

Very Blocky

Fair

Breccia

77

66

54

Blocky

Good

Porphidic Andesite

54

46

37

Very Blocky

Fair

Breccia Mineral

77

66

54

Blocky

Good

Aphanitic tectonised andesite

37

30

22

Disturbed

Bad

Amigdolidal Andesite

65

55

44

Blocky

Fair

Autobreccia

77

66

44

Blocky

Good

Andesite Cornea

37

30

22

Disturbed

Bad

5.2. Features of the structures The summary of the geotechnical features of the structures is presented in table 5.2 and is based on the mapping of the Lince open pit.

Table 5.2. Characteristic of Geotechnical major and intermediated structures Fill Estructure Carmen Fault

Thinckness Strength cm

Roghness

Type

Wall Moisture alteration E.Max E.Middle

S6

20–50

Ar-Salb.

S.O.

3–4

A.

S.

Faulty Dike

S6-R3

300

Ar-SalbrxBx-Rx

S.O.

5

S.A.

S.

S.Muelle Fault/Dike

S6-R2

10–30

ArRxBxCarb-Rx

S.O.

5

S.A.

S.

Sandra Faulty Dike

S6-R2

400

Ar-SalbRxBx-Rx

S.O.

4

A.

S.

Silvana Faulty Dike

S6-R3

80

ArRxBxSales

S.O.

5

A.

S.

Ana Rosa Dike

S6-R3

150

Rx

S.O.

5

S.A.

S.

Dacitic Dike

S6-R3

600

Rx

S.O.

5

S.A.

S.

Juana Fault

S6-R2

80

Ar-Salb.

S.O.

4

A.

S.


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Continuidad Fault

S6-R6

10

Ar-Salb.

S.O.

4

S.A.

S.

S.NS.Fault

S6-R1

5

Salb-Lim

S.O.

5

S.A.

S.

Stratification

S6-R2

0, 5

RxBx

O.

5

S.A.

S.

6. METHOD OF EXPLOITATION The deposits are exploited both by open pit and underground mining with a total production of 486,000 t/month. The Lince open pit contributes with 300,000 t, Nucleo X open pit with 50,000 t, and the underground mine exploited by sublevel stoping and cutand-fill methods with 63,000 t, whereas the small mines exploited by contractors produce 73,000 t. 7. GENERAL PRESENTATION OF THE PROBLEM The new design of the Lince open pit incorporates new geological resources which are concentrated in the old Susana underground Breccia mine, a subvertical structure mainly exploited by sublevel stoping method where chambers of large dimensions connected to the surface were abandoned. From the point of view of safety, these chambers represent a problem to the open pit operations, as it means a loss and/or dilution of the reserves. The exploitation of the new reserves requires a special modified open pit method that should consider the filling of the chambers in the conflict area. The topographies of six principal chambers using of the I-SITE system were obtained and their characteristics are shown in table 7.1. The interaction of these underground chambers with the open pit is show in figure 7.1; they were obtained by using the VULCAN software. Pillars separate these chambers about 10 m thick.


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Figure 7.1. Open pit and underground overlap Table 7.1. General information on old chambers Chamber

Method of exploitation

Volume

Floor level

Roof Level

Pillar

m3

m

m

m

1

156,500

740

780

10 (780–790)

SLS

2

281,300

695

740

without

SLS

3

53,300

655

685

10 (685–695)

SLS

4

100,000

595

645

10 (645–655)

SLS

5

60,500

585

635

without

C&F

6

247,900

495

585

10 (585–595)

C&F

Total

899,500

SLS=Sublevel Stoping, C&F=Cut and Fill

The main problems associated to the simultaneous open pit and underground exploitation at Michilla mine are listed below.


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7.1. Dilution This problem will appear in the drilling and blasting operations. Consequently, open pit blasting against filling of the chambers will be avoided. Instead, a recovery of the filling will be carried out first to avoid dilution that may jeopardize the economy of the project. 7.2. Recovery of Reserves The recovery of reserves includes ore located in the Susana area. In turn, the recovery of pillars will depend on the technique of the fill operation of the chambers where the most important aspect is the correct confinement to get the best contact between fill and pillars. However, these reserves are considered an addition to the original project. The spalling from the mineralised walls of the Susana pit is not considered because this is subject to high dilution. 7.3. Extraction Planning The current ramp, principal accesses to the underground mine, is located in phase 7 of the open pit; this means that open pit work is restricted in that phase to avoid interference with the underground operation. This access will have to be relocated depending on the sequencing of both the open pit and underground exploitation. The above demands a careful analysis in both the open pit and underground planning in order to maintain the continuity and feasibility of the access to the underground mine. 7.4. Additional Filling Operations The abandoned chambers left by the underground exploitation will be filled imposing an additional cost in the global operation costs of the mine. Michilla considers safety a number one priority and has designed a practical work methodology to carry out the filling operation. 7.5. Ventilation System The present underground production of 65,000 t/month requires about 475,000 cfm of fresh air. The presence of old chambers and the starting of the Lince open pit project have led to the existence of a connection between the open pit and the underground mines. This connection exerts a notorious influence on the ventilation system. On the one hand, it produces an additional supply of fresh air but, on the other hand, the loading and haulage operations as well as drilling and blasting in the open pit produce considerable pollution which is taken into the underground mine through this connection. Figure 7.2 shows the overlapping of chambers and open pit.


131

Figure 7.2. Open pit and underground overlapping 8. MAIN SPECIFIC GEOTECHNICAL PROBLEMS The geotechnical approach to dealing with these mining problems involves defining the stability problem and the infrastructure involved, critical parameters and the method of analysis, acceptability criteria, and the proposed monitoring program. 8.1. Stope Stability The visual inspections of some chambers or stopes mentioned in section 3 of this paper show good stability conditions from the static point of view. The critical parameters in this problem are the strength and quality of the rocks, in-situ and induced stresses in the environment of the excavation, and damage on the stope wall induced by open pit blasting. There are a few empirical rules based on rock-mass classifications, which will be useful in verifying the stable dimensions of the stopes and may be used to analyse this problem. Numerical analysis of stope and mining sequences using a three-dimensional analysis for a number of orebody shapes could give indication of possible problems and the estimates of support requirements. 8.2. Pillars and Their Role in the Global Stability of the Mine The pillars left by the underground exploitation are stable but it is necessary to carry out an analysis of the following critical parameters: strength of the pillar-forming rock, joint condition (strike and dip) which can produce slide, underground water conditions (although it seems to be favourable), in-situ stress and geometry of the pillars. The methods of analysis in this case may be based on rock-mass classifications including adjustment by dynamic events (such as blasting of rocks in the open pit) that will change the original stable rock-mass condition of the chamber walls and pillars, and the analysis of limit equilibrium that can provide useful guidance when verifying the dimensions of the pillars in different rock-mass qualities to ensure very low probability of failure. Moreover, numerical analysis of discrete elements can provide approximated stress levels and data on zones of potential failure.


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8.3. Open Pit Blasting Effect on Underground Mine An important area of the underground mine is located below the Lince open pit where big-scale blasting is carried out every day. Minera Michilla and ENAEX (National Explosives Company) carried out a monitoring of blasting to evaluate the damage produced in the tunnels of the underground mine within the influence area. This analysis is based on Hook’s law through which it is possible to define the maximum tension (σmax.) induced by a deformation (ε), produced by the level of vibrations from a blasting in a rock-mass with Young’s modulus (E).

Figure 8.1. Vibration analysis information σmax·=Eε being: ε=PPV/Vp


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where PPV=Peak Particule Velocity, Vp=“P” wave velocity in the rock-mass. ENAEX carried out a monitoring of vibrations at Susana ramp, which is a strategic area in the access to the underground mine. The QED software was used to do the analysis of the ignition sequence. The Blendex 950 explosive was utilized as a bottom charge (3 m) ANFO as a column charge (3 m) and a 5 m stemming in a pattern of 6×7 m. The initiation time for all pits was 164,9 min. and the first detonation in a pit occurred at 202,3 min., the difference being 37,4 min., which suggests that everything is initiated when the detonation of the first pit occurs. The analysis shows that the worst case presents a detonation of three pits at time intervals of 793–801 min., with a total of 399 Kg of explosives. On the other hand, the probability of having 100% of the pits exploding at the nominal time is 95%. Five monitoring points were taken into consideration in the Susana ramp between surveying points 790 and 696, the distance from the blasting should not be greater than 250 m. Eleven monitoring events were recorded. The highest vibration level reached is 16 mm/sec. with an associated frequency of 64 Hz for 132 kg of explosive at a distance of 37 m. See figure 8.1. 8.4. Dynamic Stability of Chambers and Benches The evolution of the geometry according to the sequence of exploitation will generate new conditions of stability, which are mainly originated by the variation of the open pit geometry at its intersection with old chambers. The critical parameters correspond to the accepted planning and to the extraction sequence in the geotechnical conflict zone. The variation in geometry will occur both in the open pit and in the underground mine. Wedges and blocks, which would not normally occur under normal exploitation conditions, could be generated in the former. It would be worth noting that, in certain cases, the wall of the open pit could remain unfavourably inclined, so it will be necessary to consider stabilization systems for the slopes. The critical parameters depending on the open pit failure mechanisms are: height and angle of the well-formed bench, structure slope and strike, underground waters, potential dynamic charges (earthquakes and blasting), and shearing strength along the surface of circular failures in the case of highlyfractured rock-masses. The proposed methods of analysis are: (1) analysis of limited equilibrium in relation to the failure surface, whether it is surface alone or the intersection of two planes in competent rocks or circular on highly-fractured masses, (2) failure probability analysis based on the distribution of the structural strike or the numerical analysis of stresses, respectively. The recommended safety factors for temporary slopes are>1,3 whereas for permanent slopes they are>1,5 (Hoek 1996). The same methodology set forth in 4.1 is proposed here in the case of chambers. On the other hand, the open pit design, both in its expansions and in the final slope, will have to be continuously validated mainly due to the variation of the geometry and its influence on the local and global stability of the mine. The chamber’s filling material and geometrical shape play and important role in its stability. The filling techniques will have to be highly effective to leave the material in contact with the chamber’s exposed irregular surfaces, thus avoiding the generation of important voids that could produce settlements, cavings and subsidence in the open pit. The most adequate technique to avoid this type of problems is hydraulic filling but, due


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to its high cost, it will not be considered in this case. Consequently, filling with ordinary material must be carried out using mechanical compacters and shovels to adequately spread the material. The critical parameters considered in this case are related to the sieve test of the filling material, density of spongy material, technique of filling placement, and the area exposed to landslides (Hydraulic Radius), taking into account the possible distances to the chamber’s roof and the floor of the filling material capable of settling. The methods of analysis can be the rock-mass classifications, specially that of Laubscher, that define the minimum hydraulic ratio needed to produce cavings; however, due to the complexity of the problem, it would be good to carry out a 3-dimensional numerical modelling with the purpose of getting to know the behaviour of the fill as a supporting material under new rock-mass conditions of the chambers.

Figure 8.1. Summary of geomechanical problems during simultaneous exploitation by both open


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pit and underground methods at Michilla Applying cables, which will also improve dilution, may control local instability. Macrostability is controlled by the open pit geometry, the sequencing and quality of filling the chambers and its following recovery. It is recommended to use Hoek and Brown’s fracture criteria to define the main major and minor stress relationship. The former will also be used to establish a relationship with Mohr-Coulomb’s criteria through the use of the ROCKLAB software that will define the parameters of shear strength, friction angle, and cohesion. The figure 8.1 shows a summary of geomechanical problems in simultaneous exploitation both open pit and underground at Michilla. 9. CONCLUSION AND SUGGESTIONS The operations overlapping problem between the open pit and the underground mine include technical/economical components. This brings forth an additional endeavour on the part of engineers and workers to attain an efficient and coordinated work among the different areas involved with the purpose of securing the technical/economical sustainability of the project. The geology of the deposit shows strata-linked bodies, which means that it will be necessary to diversify the exploitation systems according to each particular geological condition. The specific geotechnical problems described here have been discussed according to a logical sequencing; however, if it is technically and economically possible, priority must be given to the definition and estimation of a few critical parameters involved with each problem. The effect of vibrations caused by blasting in the open pit is under control; however, it should be necessary to continue monitoring to define maximum volume of rock to be removed per each geotechnical unit in order to avoid strong, rough distressing of the rock-mass which may threaten the stability of the geotechnical conflicting zone. Finally, it is necessary to carry out 3-demensional simulations of local and global behaviour in order to have well-founded approaches; however, if it is economically possible, they must be validated with a deformation-monitoring program. REFERENCES Carvajal 2002: Fundamentos y Principios Geomecánicos en la Aplicación del Cableado. Curso Cyted. Universidad Internacional, Andalucía, España. Hoek, E. 1996: Practical Rock Engineering. Lecture Notes. University of Chile. Santiago, Chile. Minera Michilla S.A. 2003: Procedimiento Explotación de Sectores con Laboreos Subterráneos en Mina Rajo. Internal Report. Operations Department, Minera Michilla S.A. Antofagasta, Chile. Minera Michilla S.A. 2003: Metodología de Extracción de las Reservas del Sector Susana, Fases 11, 12, 13 y 14. Internal Report.Operations Department, Minera Michilla S.A. Antofagasta, Chile.


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Ingerock 2003: Actualización Modelo Estructural Rajo Lince. Internal Report.

Underground Geotechnology of Polymetallic Ores (Based on the Example of North Caucasian Region Deposits)

Alexsandr Egorovich Vorobyov, K.G.Karginov Russian University of Peoples Friendship. Moscow, Russia Tatyana Vladimirovna Chekushina Research Institute of Comprehensive Exploitation of Mineral Resources, Russian Academy of Sciences. Moscow, Russia International Mining Forum 2004, Kicki & Sobczyk (eds) ©2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: The paper describes development of scientific—methodical bases of perfecting the methods of underground mining of polymetals by leaching ensuring increase of efficiency of development of poor ores by optimization of technological parameters. The basic idea—effective development of the not in details explored reserves of poor polymetal ores on the basis of flexible designing of systems of underground leaching. Since 1991 the nonferrous metallurgy of Russia has experienced significant difficulties connected to a general crisis of reforming the national economy. From 1990 up to 1998 the extraction of the nonferrous metals’ ores has decreased as a whole by 42%, the manufacture of lead concentrate—by 3,9%, and zinc—by 23%. The available polymetallic ore reserves reduced in time (figure 1), which was due to the fact of their occurring in favorable mining—geological and economic—geographical conditions, on the one hand, and the constant increase of the industry need for lead and zinc—on the other hand, assume in the future an organization of highly profitable feature methods available also in the earlier not featured reserves of poor and balanced polymetallic ores and rational use of bowels in the working polymetallic mines.


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Figure 1. Change dynamics of Russian lead, copper and zinc reconnoitered reserves The raw-material base of Sadonskogo lead-zinc works (SSTK) is marked up by the reserves of eleven polymetallic deposits. However their long period of exploitation has led to a significant exhaustion of the initial reserves and to an essential deterioration of the polymetallic ores quality indicators (simultaneously with the increasing cost of their processing). Currently, the basic industrial activity of Sadonskogo STSK is concentrated at Zgidskom, Sadonskom and Arhonskom deposits. Within their limits, the profitable reserves that can be mined without big investments remained for only 6–7 years. Only commissioning of new deposits to support and expand mineral raw-material base of Sadonskogo industrial complex, can extend its life. The best in this sense could be Dzhimidonskoe deposit where work in the Bozang field is done on a skilled-operational blocks by traditional underground methods, and the further intensification of the prospecting works would allow to create a compact mining enterprise and to increase the volumes of polymetals extraction, due to the use of underground leaching. This deposit has been classified as being of the 4th group of complexity, and its reserves are classified to C2 category, reserves that represent: ores—2 287,5 thousand tons, lead—20,2 thousand tons (with grade of 0,88%) and zinc—66,8 thousand tons (2,94%) in it. The total estimated ore reserves from the Bozang field classified to C1category make 457 thousand tons (which contain Pb—10 thousand tons and Zn—21, 7 thousand tons). The ore zone represents ribbed, ribbed-interspersed, interspersed and massive deposits. The total ore zone length of the horizontal well N3 makes 670 m (including the prime working field—160m). The prodeleting of the meridian ore zone steeply dipping (70–85°) to the East and West. Within the limits of SHPV skilled blocks, the Bozang ore zone wedge on the revolt and on the horizontal well N8 quartz ribs and a sulphide ore are present basically. The capacity of an ore body from the SPV skilled working field changes from 1 up to 6 m, with 2, 83 m average.

Underground geotechnology of polymetallic ores

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The content of the useful components is extremely variable: from 0,01 and 0,1 percent soars up to 15% of lead and up to 11,37% of zinc. In the total reserves classified by the B category, content of lead make 3,41%, zinc—6,17%. When calculating the excavation cycle parameters the initial data are the mininggeological conditions, the section of development and the field works organization. The choice of the mining excavation equipment is done based on these data, and also these data define the optimum borehole depth, and the duration of the excavation operations and also the number of cycles per day. First the drilling time of the borehole set to have depth l0 is determined: (1) where Nx—represents the number of the boreholes from the set; l0—represents the optimum boreholes depth, m, vcp—represents average drilling speed, m/min, nx— represents the boreholes specific consumption, m/m3, S—represents the excavation section area, m2, η—KISH. After that the time of breed shipment at the borehole depth l0 is determined, after the time of shipment at the depth l1 (without taking into account the increase of productivity): under condition that vydx=vydv

(2)

It is then necessary to estimate the loading and blasting time with increasing the drilling boreholes depth: (3) where ψ3.B—represents the factor of time spent for loading and blasting of the boreholes while increasing their depths. Time for ventilation while increasing their depth of drilling and including the relative factor of spent time for ventilation ψ′yб is defined by the expression: (4) The relative time of preparatory-final operations in the cycle (for 1 m of production) will be equal: (5) The duration of fastening operations, of flooring ways, of fixing pipelines etc will be expressed by a constant value because the time of these operations will proportionally increase with the boreholes depth and it will be expressed by the formula: (6)


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The numbers of boreholes from a face depends on the section of work and can be expressed by the specific drilling consumption: (7) where nx—represents the boreholes specific consumption at the increased borehole depth, m/m3, n1—represents the boreholes specific consumption at the initial boreholes depth, m/m3, a—represents the factor of relative boreholes specific consumption change, equal to 1,1 m/m3, practically constant for all breeds categories. The borehole rational depth is defined by the formula: (8) where vcp—represents the boreholes drilling speed by the part of drifters, m/min, tΠ.3— represents the time of mine preparatory-final operations; a=1,1—represents the proportionality factor. After that we had investigated the influence of various geological factors and parameters of the drilling and blasting operations on the efficiency of the outlined excavations. The fraction parameter of Vmax rocks, which changes for over a wide range (from 0,8 up to 26 cm3), has been chosen as a criterion for the estimation of the rock strength when estimating an optimum size of VV charge in the perimeter shotholes. The basic parameters of the jointing which influence the quality of the outlined exposures are: the joint width r; the angle of intersection of the joint surface and of the exposure surface γ; the angle formed by the joint trace direction having boreholes outline direction φ and the character of joint-gauge fill. The degree of the cracks opening is characterized by the opening depth and by the КB opening factor: (9) where КB—represents the crack opening factor (changing from 0 up to 1), LB—represents the opened field length, cm, L—represents the general crack length within the limits of the opened field, cm. In the figure 2 there are represented the experimental dependances on the change of the КB opening factor and the opening depths hT of the angle formed by crack with the face exposed surface at the moments: usual (1) outline (2) detonation. It is visible from the schedules, that the detonation outlined method allows to reduce considerably the size of the opening and the fracture depth.


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Figure 2. The influence of the angle of intersection of cracks and the exposed surface with: (a) opening factor; (b) their depth, utilizing: (1)—standard, and (2)—smoothwall blasting methods Researching the influences of the type of VV on the quality of the excavation outlining that is determined by the characteristic the VV brisance has been undertaken. The experiments have been made at constant burden between perimeter shotholes ax, LNS, W and the boreholes diameter, but with various VV types at a specific loading qo equal to 0,3, 0,45 and 0,1 kg/p.m. At the research of the influences of the VV charge diameters and of the perimeter boreholes on linear deviations size of the actual contour of development from design it has been established, that with increasing VV charge diameter the linear deviations will grow sharply. All these factors and parameters have allowed optimizing the 7th borehole drilling volumes from Dzhimidonskom deposit (figure 3).

Figure 3. N 7 Borehole drilling volumes With the purpose to define the rational parameters and using the method of splitting explosive, that provide the least section deviations of the mining excavation from the design values, we have made some industrial tests of various BVR passports with the elements of the smoothwall blasting, in unstable, semi-stable and stable rock. For tests the following initial parameters have been accepted: the burden between perimeter shotholes 40–60 cm (borehole diameter 40–42 mm); the proximity factor of


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perimeter charges m=0,6–0,8, VV charge weight for 1 l.m. of perimeter shothole of 0,2– 0,3 kg. 7 variants of different types of VV charges designs (continuous and dispersed) have been tested. For each variant 3 skilled explosions have been made. The estimation of the outlining quality and the drilling efficiency have been done according to KISH, using the choice of breed, safety of “traces” (percentage relation of “trace” length to initial borehole length) and the size of the actual excavation section. The skilled works were carried out in two directions: working on BVR parameters with the purpose of achievement KISH of size 0,9–0,95 and researching for the optimum parameters of outlined detonation. The skilled explosions have allowed revealing the optimum pattern of shothole arrangement in a face for which the W specific charge has decreased from 2,98 to 2,45 kg/m3 (by 21%), and KISH has increased from 0,8 to 0,98. Also the optimum parameters of the perimeter blasting have been determined: the distance between the perimeter shot-holes of 0,45 m arch, the distance between the perimeter shot-holes of 0,5 m boards, LNS charges perimeter shot-holes of 0,55 m, the proximity factor—0,5. The size of the perimeter shot-holes charge at the application of 6 GV ammonite made 0,8 kg, and detonate 10A—0,6 kg, having the cartridges diameter of 28 mm. As the result of carrying out a series of trial explosions the BVR passport has been processed, which has allowed lowering the linear searches on excavation boards by 40%, on flooring—by 60%, and on the excavated section area—by 60–70%. It has been established, that if maintain the parallelism of the perimeter shot-holes and corners of their inclination to the face of 85–87° then the breed searches do not exceed 2–3%. Thus the volume of the drilling has increased by 14%, and the VV specific charge has decreased by 22% in comparison with usual blasting method. The calculation has shown, that employing smoothwall blasting and carrying out support of the excavation by shotcrete or by monolithic concrete support, the labour expenses per 1 l.m. of excavation have decreased, accordingly by 0,2 and 8 manhour. In this way, smoothwall blasting allows to lower breeds’ searches to 3–4% and to reduce considerably the total expenses of carrying out the excavations. Decreasing the seismic influence of explosions on the surrounding excavations we propose carrying out of the following actions: – reduction of the shot-hole number on a face with the purpose of bringing auxiliary and spitted boreholes LNS to 600–800 mm, and the burdens between the top perimeter shot-holes to 500 mm, – reduction of quantity of VV, which is blasted in one step of delay, to 2–4 kg, – use in the top perimeter shot-holes of 6 GV ammonite charge dispersed by air intervals, initiated DSH. The size of the charge of these shotholes depends on the stability and on strength of breeds and should be within the limits of 0,2–0,6 kg, – the excavation drilling with an advancing face and its subsequent expansion to design section—the size of advancing of a face is equal to two exits, – application of the combined detonation of VV charges using the EDK3–153N and EDK3–153N electro detonators at the maximum number of delay steps, – detonation of VV charges in the top perimeter shot-holes in the last turn, – increasing the efficiency of the boreholes placed in the drilling.


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The domestic and the foreign experiences show, that one of the ways of increasing the efficiency of drilling work in the underground mines is the reduction of the boreholes diameter. Industrial tests of different makes of drill crowns have been done with the purpose to estimate the efficiency of the application of crowns of reduced diameter. The results of the tests have shown, that the speed of drilling with crowns having the diameter 36 mm is higher by 20–30% in ore and by 30–40% in breed than the speed of drilling using crowns having the diameter 40 mm. During the processes of drilling with BI-36–22 crowns in ore, increase in speed of drilling by 10–12% was observed, with the reduction of the stretch angle made between the plates of a firm alloy from 110 up to 90– 95°. We had revealed a certain dependence of drilling speed change on the number of holes drilled per one crown (figure 4).

Figure 4. Change of drilling speed depending on the number of holes drilled per one crown: 1—BI-36–22 crown, 2—KKV-40 crown, 3—KKA40 crown Also we have made skilled explosions with reduced diameter boreholes. In the quality of VV boreholes with diameter 40–42 mm we have used 10A detonator and 6 GV ammonite. The analysis of these data shows, that the reduction of the borehole diameter from 40 to 36 mm allows improving considerably the BVR economic parameters at the moment of mining excavations. Besides, when reducing the boreholes diameter, the time of preparing the face is considerably reduced which is very important for the organization of carrying out of high-speed excavating. The research of BVR passports and of the trial explosions results, executed when drilling the mining excavations, shows, that the specific VV charge depends, basically, on the excavated section, on the ores and breeds strength, on their structural and textural features, on the borehole depth, VV type and its loading method.


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Table 1 shows average actual BVR parameters as the results from the drilling (75 cycles in excavations section of 3,2–14,4 m2, in ores and breeds with f=4–13 various jointing and stability). The specific charge of the AS-8 granulate in the section developments up to 6 m2 by 1,26 times, and in the section developments >6m2—it is by more than 1,3 times, than VV cartridge (accordingly 4,72 and 3,8 kg/m3 against 3,94 and 3,02 kg/m3).

Table 1. Basic BVR factors when drilling excavation Drilling excavations

Layered dredge

Section till 6 m2

Section >6 m2

Basic Clearing layer

Average area per one chink, m2

0,25

0,29

0,69

0, 63

3,2

Passport average of the VV specific charge, kg/m3

3,8

3,02

1,24

1, 27

0, 95

Increasing factor of VV charge of the boreholes capacity, КB

1,38

1,38

1,52

1,52

1,46

Power drilling factor in view of the thermodynamic losses, КЭ

1,3

1,3

1,4

1,4

1,3

The recommended specific charge of the AC-8 granulate, kg/m3

4,1

3,25

1,35

1,38

1,07

AC-8 granulate carrying out while pneumatic loading, %

5

5

8

5

10

Temporary normative of the AC-8 granulate specific charge, kg/m3

4,3

3,4

1,46

1,46

1,21

Average specific factor of the AC-8 granulate charge based on the industrialskilled explosions, kg/m3

4,72

3,94

1,83

1,5

1,23

Type of mining works

Chink split

Note: the specific charge and the normative are given in view of insurgents of VV cartridge.

Using the pneumatic method of loading the granulated VV, the boreholes are filling on full section with average loading density 1,1 g/cm3, thus the capacity of 1 l.m. of the borehole is direct ratio to the square of its diameter. At the moment of the borehole loading having diameter from 36 up to 42 mm with VV cartridges having diameter of 32 mm, their capacity remains practically constant and equal to 1 kg/m.


145

Figure 5. Comparative efficiency characteristics of the SHPV polymetals different actions Except this, a significant influence on the efficiency of the underground mine leach technology renders such factor, as the quality of the explosive crushing of the polymetallic ores, which is the basic technical parameter (figure 5). In particular, for the subsequent technological leach processes it is rather expedient to provide explosive crushing destruction of rocks on inter grain (inter mineral) surfaces (figure 6). This process will be defined, including such an important parameter, as the grains average size of separate minerals.

Figure 6. Revealing of the optimum effective leach dependence on the polymineral breeds selective strengthening As a whole, the quality of the explosive crushing is defined by the steady properties of the destroyed breeds, as shows the research results that we have developed in the corresponding passports. Besides that, the explosive destruction of the polymetallic ores will be also defined by type of bond in the minerals. Rather important parameter, that influences the quality of blasting, is also BB type. As VV researched examples have been chosen granutol, ammonite N1 and aquakvatol M-15,


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as the most distinguished in their properties. The specific VV charge has been accepted as equal to 1 kg/m3, that is more often used in BVR manufacture. Ore minerals (with the exception of chalcocite and molybdenite) strengthen by less degree explosion, than the other breeds (except quartz, the last of other breeds that are more steady to explosive dispersion). That is, at the detonations of mining weight there is a primary micro-crack destruction and an opening of ore minerals for the subsequent influence on them of leaching geotechnological solutions, penetrating into micro-cracks. At the explosion emission strength and concentration of the micro-cracks decreases. The quality of the mined massif crushing using blasting for the extraction of the polymetallic ores, the underground leach is one of determining factors of the efficiency of the applying this method on rocky deposits. In connection to this substantiation and this choice of BVR parameters—of the specific VV charge, to the distance between the chinks ends on fans (a) and the lines of the smallest opposition (W) have been given a special attention at the moments of designing and preparation of ore for the leaching in the SPV skilled-industrial block (table 2). The efficiency and competitiveness of the minerals extraction using the method of underground leaching in comparison with the traditional development methods of the rocky deposits (figure 7) substantially depends on the extraction completeness of extracting a useful component. The results of the technical and economic calculations and design studies show, that on the polymetallic deposits which have balanced reserves, despite of obvious advantages, the polymetal underground leach method cannot compete with traditional underground mining rocky ores having an extraction factor below 72– 70%, because of the big losses of the useful component.

Table 2. BBP indicators at the moment of leaching SPV block Indicators

Chinks fans I

II

III

IV

V

Block average importance

The quantity of broken ore, m3

788 780 570 778 618

The distance between the ends of the chinks, m

3,0

3,0

3,0

3,0

3,0

3,0

The line of the 3,0 least resistance, m

3,0

3,0

3,2

3,0

3,04

The rapprochement of the chinks factor

1,0

1,0

1,0 0,93 1,0

1,0

VV specific charge, kg/m3

1,0

0,9 1,58 1,4 1,58

1,27

Loosening factor, 2,36 1,8 Кp

1,6 1,46 1,34

1,34

Variation Relative Trust factor of error % interval quartz %


Output of ore from 1 m of chink, m3/m

5,02 5,21 3,23 3,67 3,34

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4,1

Fraction output, % К1 (0–50 mm)

66,6 52,4 64,1 70,6 66,9

64,1

3,0

2,2

1,5

К2 (+50–100 mm) 12,0 13,0 14,3 17,4 24,4

16,2

3,2

23,8

3,8

К3 (+100– 150mm)

5,5

4,6

8,8

6,6

3,8

6,5

52

38,6

2,5

К4 (+50–200 mm) 10,3 10,4 3,2

2,3

2,6

5,8

69,1

51,3

2,8

К5 (+200 mm)

2,6

3,1

2,3

7,4

80,0

59,4

2,6

Σ К1−К4

97,4 80,4 90,4 96,9 97,7

92,6

3,8

2,8

2,6

Piece average diameter, cm

6,0 11,5 7,7

5,4

5,2

7,2

Piece average diameter determined by the polymetal losses, cm

9,1

6,7

8,7

8,8

9,6

8,2

9,6

7,3

Figure 7. Completeness of the polymetals extraction from ores. Extraction: 1—specific, 2—through The increase of the polymetals extraction completeness could be probably achieved only using a finer rock-mass crushing. But then the other formidable factor inures: at the moment of using the infiltration leach system (the only real variant for rocky ores is to be introduced commercially), the thinnest crushed product is more easily to press, practically turning in a water-emphasis that is not studied with a reagent that leads to the polymetals loss. If in a traditional technology to have pieces of the rock weight having a size of more than 400–600 mm it is simply undesirable, then the presence of the broken ore in SPV rocky polymetallic ores of more than 15–20% of the fraction of 200–150 mm can in some


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cases worsen the extraction quality indicators and raise the cost price of final product up to the limits of its unacceptability. In case of the deposits with thin interspersed mineralization the useful component in the broken ore is, at crushing, distributed in regular intervals under the maintenance in pieces of the different sizes, therefore the extraction factor of leaching process is proportionally reduced with output of large fractions. In the same time in the interspersed and thin interspersed mineralizations a redistribution of the mineral substance in broken ore is observed—fine fractions have higher share. The structure of the metals losses and their influencing character on SHPV method parameters change. Despite of rather low metal extraction from pieces that have size of +150 mm, the specific losses, falling into their share, are insignificant, and in some cases it is less, than from a trifle. The increase in the size of the average piece of broken ore due to the increase of an average output and due to the large fractions and reduction an average weight, the trifles not only worsen the completeness parameter of leaching, but renders a positive influence on the technological parameters. The quantity of the lost useful component, falling per one surface unit: (10) where SHi—represents the surface of i-fraction, Q—represents the general losses of the useful component, come per one share of particles of i-fraction, Ni—represents the number of pieces from i-fraction. The extraction completeness of the useful component from ore pieces is proportional to the general active surface of the environment including (besides the external surface of the particles) also the planes formed by the cracks open under the influence of the work of blasting in those pieces. It is necessary to notice, that the designing parameters of the extraction completeness are practically provided only to the due of 0–50 mm fractions. The metals additional gain is expected to come from the fraction of 50–150 mm, on which share come 8,6% more of metals and ~23% of rock weight. The other efficiency fractions of the essential process will not much render (though they can play a positive role). Due to the fact that leach process is mostly intensively proceeding at the beginning of the 4th–6th months, which are enough for the processing of 0–150 mm fractions, then half in comparison with the project should diminish the duration of leaching at the achieved crushing size. REFERENCES Vorobev A.E., Karginov K.G. et al.: Designing of the Geotechnological Enterprises. The Mining Information—Analytical Bulletin. No 4, 2002. P. 170–173. Vorobev A.E., Karginov K.G.: The Concept of Mineral Resources Reproduction in Lithosphere. Russian Ecology and Industry. February 2001. P. 33–38. Karginov K.G., Chekushina T.V.: The Management Concept of the Natural Leaching Processes. Proceedings of The 1st International Conference: “Resource Producers, the Less Polluting and Nature Protection Technologies of Ore Formation”. M.: RUDN Publishing House, 2002. P. 63– 65.


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Chekushina T.V., Karginov K.G.: Designing of SHPV Mines on Insufficiently Reconnoitred Deposits of Minerals. Proceedings of The 1st International Conference: “Resource Producers, the Less Polluting and Nature Protection Technologies of Ore Formation”. M.: RUDN Publishing House, 2002. P. 166–168.

Technological Advances in Underground Mining of a Stratified Copper Deposit in Poland

Jan Butra CBPM “Cuprum” Spółka z o.o. Wrocław, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 MINING CONDITIONS IN LIGHT OF GEOLOGICAL SETTING OF COPPER DEPOSIT OF THE FORE-SUDETIC MONOCLINE Since its discovery, economical and safe exploitation of the copper deposit in the LubinGlogow district (Fore-Sudetic Monocline, Poland), has been largely dependant on the choice of a mining technique. Due to the overall geological complexity of the deposit, the list of critical characteristics that have to be taken into account while selecting an optimal technique inccludes: – substantial depth of occurrence of the deposit that ranges from 600 m to 1200 m below surface, – diverse lithology, – substantial aerial extent combined with a low (4° to 6°) dip angle of mineralized strata, – rock-mass competency differences (the strength of main roof strata is 7 to 10 times greater than that of the floor), – the ability of ore and surrounding rock-mass to accumulate, and violently release, elastic strain energy, – complex tectonic conditions, – diffused nature of mineralization and the absence of macroscopically identifiable mineralization markers. The stratiform sulfide copper deposit of the Lubin-Glogow district is hosted by discoloured sandstones of the Red Beds and/or sandstones of the White Beds, copperbearing shales, and carbonate rocks with dolomite as a dominant rock type. Significant concentrations of copper occur within a 40 meter thick stratum. Six classes of ore have been identified depending on their location within the lithological profile:


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– dolomites only, – shales and sandstones, – sandstone roof strata only, – exclusively sandstone strata located below the sandstones of anhydrite binder, – sandstones, shales and dolomites typical to the entire mining district, – shales and dolomites. Both the extent and intensity of copper mineralization reflect the natural extent of sulfides deposition. However, none of these features correlates strictly with lithological divisions of the host rocks. Thickness of the deposit ranges from 0,4 m to 26,0 m. Individual rock layers dip NE at low angles of 2° to 5°. Brittle tectonics is the major factor contributing to the complex geometry of the deposit that makes exploitation of the latter a challenging task. Individual faults run parallel to one another and they either converge or diverge at various angles. Throws of the largest faults range from 50 m to 60 m. The diverse lithology furnishes diverse distribution, intensity and dip of joints, as well as characteristics (roughness) of fractures’ surfaces. The presence of ore-forming minerals in three different lithological units (shale, sandstone, dolomite) combined with uneven thickness and intensity of copper mineralization produce the complicated picture of the deposit. The cut-off grade is set at 0,7% Cu whereas deposit thickness and copper content variability in individual mining blocks have been estimated to be 81,5% and 78,3% respectively. Rock-mass parameters and the depth of mining works below surface have the greatest impact on the safety of mining. Spatial distribution and physical properties (both across and along the stratigraphic profile) of individual lithological units represent basic criteria for estimating geo-engineering characteristics of copper ore-bearing rock-masses of the Fore-Sudetic Monocline. Due to the lithological diversity and the corresponding physical inhomogeneity, determining these estimates become particularly difficult both for the roof and floor strata of the deposit. Of particular importance is the 200-meter thick series of roof anhydrites and dolomites whose strength is 7 to 10 times greater than that of the floor. Due to variability of strength parameters and tectonic disturbances with the depth, exploitation of the deposit is carried out under severe stress conditions. As a rule that applies to the entire mining district, the copper bearing strata rest on a series of low or very low-strength sandstones that respond to deformation forces in an elasto-plastic manner and that are not prone to rock bursts. As to the main roof strata, the geographic distribution of mechanical properties of individual lithological units has led to the following rock-mass classification (figure 1).

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Figure 1. Distribution of the copper deposit as a result of different types of main roof strata A. Central Region represented by a 160 m to 220 m thick suite of roof strata lithologies, namely anhydrites and dolomites that cover most of the district’s area. The compact, high strength rocks of this suite form a stiff layer prone to rock bursts. It rests directly on the top of the ore-bearing series and is a source of dynamic phenomena while subjected to mining. B. Northern Region characterized by the presence of a rheological layer of salt that is up to 180 m thick and rests on the rigid layer of dolomite and anhydrite. The latter is barely 32 m to 90 m thick and, as in the Central Region, represents the roof strata of the copper-bearing series. By redistributing stress, the salt has a significant impact on stress conditions of the undisturbed rocks subjected to mining. C. Southern Region with a unique geology. As a result of the specific sedimentary and tectonic environment, rocks of the Tertiary and Quarternary age lie directly on sediments of the Zechstein age, namely anhydrites and, occasionally, on carbonaceous rocks.

MINING METHODS USED DURING THE FIRST PERIOD OF EXPLOITATION The concept of the copper ore mining in the “Lubin” and “Polkowice” mines from 1964 assumed the extraction by longwall method known from the “old” basin (“Konrad”,


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“Lena”, “Nowy Kosciol”). In May 1965 mining operations in two faces (one 40 m long and second about 50 m long) in an experimental panel of the “Lubin” mine were commenced. Both walls were run in-line, what gave in result one longwall. The faces were supported by SCG-51 steel extension roof bars 1,2 m long, SW-30a early supporting props and Vallent immediate supporting props with 1,2 m spacing along the strike and 1,0 spacing along the dip. The ore winning was carried out using blasting (1,4 long blasting holes, 5G1 or 5A dynamite, ZE delay action, one second fuses). At first the ore was manually loaded on the chain conveyor Slask 59. Then the WŁE-50S cutter was used to load the ore on the conveyor. At the end 1967 only scrapper loader with 0,75 m3 capacity was used for ore loading and haulage. Transport of the ore from the wall was carried out by PTG-800 and then PTG-1000 belt conveyor to the point where it was poured onto the cars of 0,86 and 1,36 m3 capacity and then along the rails to the shaft and onto the surface. Appropriate roof control was initially provided by using the dry backfill (barren carbonate and sandstone rock crushed underground) until about 1,5 ha of mined area was obtained. Then till finishing the experimental longwall mining (1969) the roof control was based on self—caving or roof fall induced by blasting (about 3–3,5 m high). In 1967 a mining field for the room-and-pillar system was prepared. The field was cut into large dimension (25×35 m or 25×25 m) pillars by 5 m wide rooms. Room-and-pillar mining based on cutting large pillars into small technological (supporting) pillars having dimensions of about 5 by 5 meters. The ore was mined using blasting technology basically. The blast holes were drilled by hand-operated rockdrills (WUP—22) and later by drilling cars (SBU—USSR made, Serpent—import from the west-European countries and SWW—domestic production). The length of machines drilled blasting holes was 2,8–3,8 m with diameter of 36–46 mm. The blasting holes drilled by hand-operated equipment had 2,4 m length and their diameter was 36 mm. Dynamite, ammonite and newly developed saletrol were used as the explosive materials. The fleet of heavy, self-propelled underground equipment, which was prepared at that time as a result of wide co-operation and import, allowed for full mechanization of underground mining works. 18HR4 or PNB-3k loaders and haulage cars of Expadump14D2 type had already carried out the mined-out ore from the test field of room-andpillar method. Later, the prototype of Polish ŁK-1 loader was used. This loader could also work as an ore haulage car on short distances, for example to the loading site. Only roof bolting, mainly expansive bolts were used in order to obtain the proper roof support. Roof control in the first experimental field of two-phase, room-and-pillar mining was based on carrying the bending roof on the supporting pillars and then inducing roof caving through blasting the remnant pillars (having cross-sectional area of approx. 12 m2) together with blasting shotholes drilled in the roof. The roof caving in the room-and-pillar method was widely used during seventies and eighties. It was used until the early nineties thus over 20 years. This technique of blasting was continuously adjusted to the mining methods, being permanently modified, and changed very often depending on the geology and rock-mass conditions. The characteristic of the caving blasting was simultaneous blasting in roof and residual supporting pillars. The holes in the roof over the blasted pillars were relatively short (length 2–3,5 m and diameters up to 51 mm) or much longer (5–10 m, rarely longer), 51 mm of diameter. The holes were drilled under an angle of about 45°. At the end of the period when inducing of roof caving by blasting was


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practiced, the remnant pillars were not being blasted. However, different variants of blasting patterns with groups of holes drilled in the fan shape in one or two rows were used broadly. Two-phase room-and-pillar methods with roof caving had been used till 1974 both in “Lubin” and “Polkowice” mines. At the time when those methods were applied, high seismic hazard was noticed because longer pillars were very stiff and behaved as a quasielastic material. During the second phase of extraction, in the course of cutting those pillars into smaller ones (post-failure) their sudden destruction due to transformation of solid rock from the elastic to the post-failure state, in the working area frequently took place. In the early seventies, the first methods of seismicity and tremors control such as for example so called “limited advance” of the development in front of the face were introduced for the first time (figure 2).

Figure 2. Diagram of two-phase roofand-pillar method with limited advance of the first phase mining However, this technical modification had not resulted in seismic hazard limitation. This in turn caused rejecting the mining methods, which produced stiff pillars in the first place. Presence of single or grouped headings in the solid rock in front of the face created


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also very dangerous geomechanical condition. Therefore two-phase methods with roof caving, where the solid rock-mass was cut into stiff pillars, had been gradually abandoned. One-phase room-and-pillar methods had been favored instead (figure 3). In this method the ore is mined out by one face consisting of numerous room faces connected by workings with technological pillars securing roof stability within the working space. Such systems avoid inconveniences which result from initial cut of the deposit into blocks, like in two-phase methods, and permit obtaining much more favorable distribution of mining stress in the vicinity of the face. With the growing depth of exploitation and higher variability of rock-mass and roof characteristics, the difficulties in using those methods caused by increasing stress, had occurred. Problems with proper adjustment of system’s parameters (pillar dimensions, rate of mining face progress etc.) to the given geology and rock-mass conditions, were encountered. The most important issue was technological pillars’ size selection providing the roof strata stability without the large amount of elastic energy stored within the rockmass.

Figure 3. Diagram of one-phase roomand-pillar mining method Moreover at the advent of one-phase room-and-pillar methods, technological pillars were usually located with long axis in parallel to the line of mining face. During next years, due to the necessity of concentration the possible largest number of room faces along the specific mining face (in order to induce the dynamic effects by mining blasting) the technological pillars were cut perpendicularly to the face line (figure 4). Increase of the distance rooms are advanced into the solid to anything between ten to twenty, and even 30 or more meters currently allows to conduct blasting in the zone of exploitation stress and obtaining better conditions for rock-mass controlled tremor


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occurrence. Placing the pillars with their longer axis perpendicularly to the mining face is also favorable from heading roof stability point of view.

Figure 4. Diagram of two-phase roomand-pillar mining method with pillars located with longer axis perpendicularly to the mining face The essential modification concerning mining methods in the early eighties was the development of the method exploiting the technological pillars whose slenderness was calculated based on the deposit’s thickness. The J-S mining system is one of that employing the above-mentioned solution (figure 5). The mining of thick deposit has been carried out using room-and-pillar method with two layers of hydraulic backfilling. Due to easy roof control in the dolomite-limestone rocks, the upper layer galleries were usually excavated in the first stage of mining. During this phase the stiff, large-dimension pillars (25–30×30–40 m) were cut. During the second phase the pillars were split before the lower layers could be mined. First mining methods using hydraulic backfilling for 7 m thick deposit UZG and D-P (figure 6) were developed for the “Lubin” mine. In early eighties the two-phase, twolayer, room-and-pillar method “Rudna 1” was introduced in the “Rudna” mine (figure 7). It enabled mining deposit even up to 10 meters thick. The mining face in this method is divided into three blocks with the following operations: – mining the upper layer, – mining the lower layer, – backfilling the mined-out space. For the deposit with thickness bigger than 10 meters two-phase, two-layers methods “Rudna 4”, “Rudna 5” and “Lubin 1” were used.


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Figure 5. Diagram of J-S room-andpillar method with roof caving

Figure 6. D-P room-and-pillar method with hydraulic backfill. Way of driving the workings at the mining face


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Figure 7. Diagram of two-phase, roomand-pillar method NEW SOLUTIONS IN MINING TECHNOLOGY In seventies and eighties the technological roof caving was utilized when mining deposit with thickness up to 5 meters located out of protection pillars. Together with inducing the caving, the remnant pillars were blasted out. This method of technological caving had not met the requirements because the blasted roof plate with low loosing factor of the fallen rock had not ensured the appropriate support for upper roof strata. This resulted in periodical stress increase in the solid in front of the mining face and numerous tremors and stress-relieve events caused by emission of elastic energy gathered in the rock-mass due to exceeding the compressive strength. The tremors and stress-release events were often of global nature and covered major part of the mining area. Therefore, intensive investigations on improving the technology of caving and a search for solutions ensuring the required roof support in the mined-out space were carried out. The J-Z mining method with limited (with regard to place and height) roof caving and J-3S method with periodical caving and intensive blasting were introduced. The idea of that solution was giving up blasting of the remnant pillars. At the same time experimental exploitation with roof control by its deflection on the remnant pillars and spontaneous caving of the roof rocks, without its induction by blasting were introduced in the “Rudna” mine. On the basis of this experiment the experimental exploitation with roof deflection was initiated in the G-7 mining panel of “Rudna” mine. It was the first panel in the mines located in the Fore-Sudetic Monocline using the new method of roof control at the entire length of the mining face. Within comparable periods before and after the mining method with protected roof was implemented, the following were observed: – reduction of the number of seismic events classified as bumps, by 66,2%, – decrease in amount of energy emitted from the rock-mass by 90,3%.


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This method of mining essentially limited the number of spontaneous seismic events occurring in the solid in front of the face. The relatively quick increase of the mine headings convergence was observed. It confirmed that the nature of the roof strata was different than before. During one year of mining associated with simultaneous visual and acoustic observations, no increase of the tremors hazard was noticed in G-7 panel while the reduction of the roof fall hazard level had shown to be evident. Caving of the roof in goaf occurred spontaneously and the fallen rock filled up all voids already at the forth line of remnant pillars behind the line of final extraction works.

Figure 8. Diagram of mining method with roof deflection

Figure 9. Diagram of R-UO room-andpillar mining method In 1991 and 1992 the extent of mining without technological roof caving was increased. The investigations made in 1993 revealed that the ore mining without technological (forced) caving, with giving up blasting the remnant pillars and using the natural


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susceptibility of the roof to deflection and to the spontaneous caving, did not caused any increase of the rock-mass stress hazards (relieves, tremors). It was also stated that it had positive impact on the stability of headings in the mining space at the face area. Since 1994 one-phase room-and-pillar method with roof deflection (figures 8 and 9) is the most often used mining technique. The idea of mining with roof deflection consists in cutting the ore by the system of rooms and creating technological pillars of the dimensions allowing for their yielding due to excessive load (figure 10).

Figure 10. Characteristics of the technological pillars behavior in the mining fields Yielding pillars working within decreasing slope of stress-strain characteristics allow for roof deflection in similar way as timber packs or dry backfill do in classical mining methods. Roof caving takes place far from the mining face front, after partial extraction of the technological pillars. Up-to-date experience confirmed that the bolted roof has the tendency to yield without causing caving on the mining face. Planar dimensions of pillars and their slenderness are being adjusted depending on the characteristics of pillar rocks and local rock-mass conditions. Positive results of experiments carried out in the deposit of intermediate thickness led to widening the usage of this method of roof control also to thin deposit conditions. A new variation of the room-and-pillar mining method is one with so called “closing operational” pillar. In this method gob formation is not induced over the whole length of the workings immediately behind the progressing face. In the mining method with closing operational pillar (figure 11) during the mining advance, the full mining (from


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development to goaf creation) process is carried out with some delay, except the separated zone (40–120 m wide) which is treated as barrier pillars, i.e. they are mined-out after completing the development in the entire field.

Figure 11. J-UGR-PS room-and-pillar mining method with deflection and closing operational pillar Mine workings in the separated operational pillar play transport and ventilation roles. Starting the mining operations depends only on small-scale development mining creating for instance a group of headings from which the mining face will be commenced. The copper ore along the whole face line is mined through 7 meters wide rooms and headings with separating them technological pillars (5–9×6–16 m), located with longer axis perpendicularly to the face line. With the progress of mining face, the technological pillars at the line of goaf zone are mined mechanically using loading and hauling machines or using explosives, depending on the degree of disintegration. The resulting smaller pillars, after proper scaling are left as remnant pillars in the gobs, where they act as supports moderating the curvature of bending roof layers. When barren rock is


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extracted, during selective mining, heading re-arrangement or dinting, for example, it is placed in gobs, especially in the vicinity of the residual pillars being formed. The rock, which is delivered there, gives additional roof support or side spragging to the remnant pillars at the gob formation line and in the gob zone. When the volume of waste rock exceeds the capacity of mined out voids it is allowed to place it in the operating area at the gob side. Filling only every second room with rock allows free access to each plot in the worked-out area without necessity for access ways to be opened up. When the mining face is in progress robbing of technological pillars does not cover this part of field where the pillars are left intentionally. They gradually form elongated closing operational pillar. In one entry of that pillar the belt conveyor is installed, which is successively elongated along with progress of solid cut. In order to improve the working’s stability within the operational pillar, it is allowed to fill it with waste rock. Robbing of technological pillars within the area supported by a closing operational pillar is carried out after completing the mining operations in the section. In relation to the previous direction of the face, direction of pillar robbing is reversed or it is done from the middle of operational pillar towards its flanks.

Figure 12. Diagram of one-phase, onestage, room-and-pillar with roof deflection mining method Mining of thin deposit using self-propelled vehicles higher than the deposit thickness results in ore dilution i.e. lower level of copper-bearing mineral content. Using the


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separate mining of ore and barren rock limits the dilution only in the first phase of mining i.e. driving rooms and workings. Since technological pillars are crushed mining them separately is not possible during the subsequent phase. Therefore the one-phase method is used for mining the deposits whose thickness is lower then minimal height of entries where self-propelled machines can be used. The solid is cut into the technological pillars with separate mining the deposit and barren rock. This phase of the ore mining is the final one because the separated pillars, having the dimensions of residual pillars, stabilized by waste rock, pass with the face progress to the goaf zone (figure 12). While mining the thick deposit (thicker than 7 m) two-phase methods are not used at present. Currently the most modern solution with to regard one-phase mining (with hydraulic backfill) of thick deposit are as follows: “Rudna-7” for deposits in rock prone to tremors (figure 13) and R-C/PH for mining the thick deposit outside the protective pillars. R-C/PH method is one-phase, two-layer with partial gob-creation in the minedout space with the use of hydraulic backfill. It uses the experience obtained during exploitation with roof deflection and hydraulic backfill and combines both in its method of gob formation.

Figure 13. Diagram of RG-7 one-phase room-and-pillar method with hydraulic backfill


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MINING OF BARRIER PILLARS AND ORE RESIDUES In the eighties experiments were undertaken to extract the chain pillars left for protection of main transport and ventilation galleries. The following factors have influence on difficult conditions encountered during mining of each pillar: – lithology and rocks strength, – width of the barrier pillar or width of solids blocks in the pillar, – width of gobs on the both sides of pillar and methods of roof control, – depth of mining. When the first pillars in the “Lubin” and “Rudna” mines were mined using already known room-and-pillar methods with gobs created behind the face line, tremors took place. Therefore those methods were replaced with the two-phase method characterized as follows: – during the first phase, the pillar solid on its entire length are yielding through cutting the solid by the mining face moving in the one direction towards the technological pillars being in the post-failure state (figure 14), – during the second phase the technological pillars were mined out and the gob creation was carried out along the face moving towards the opposite direction.

Figure 14. Diagram of ore mining using R-UO/FO room-and-pillar method with the roof deflection when the resistant pillars are yield and liquidated In 1989 this method was used to mine the chain pillars in “Polkowice” and “Rudna” mines. This method allowed to reduce the hazard caused by dynamic symptoms of the rock-mass stress without increasing the roof fall hazard in the transport and ventilation galleries driven in yield zone either in mine workings during the second phase of exploitation. Using the gained experience new one-phase method with roof deflection for yielding and robbing of chain pillars—R-UO/FO was developed. The presence of zones with barren rocks or with small mineralization within the mining field is a very important problem when the deposit of small and medium thickness is under exploitation. To advance production faces through these zones was not


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economical. Attempts to fracture rock-mass in the zones by blasting were not successful despite of high expenditures. Investigations helped to establish the impact of the pillars formed by barren zones left in-situ on the seismic hazard in the mining field and surrounding area. It allowed establishing the width of the barren zones, which may be left in the mined out area, without causing any hazard of tremors. The width of the zone cannot be less than 20 meters for deposits up to 5 meters high. In complex geological and rock-mass conditions occurring in the copper mines, the deposit lot may have such location (residual block, from ten to several hundred meters wide), that the only possible direction of mining is towards the gobs or area cut by workings (yield zone). Exploitation of the deposit residues towards the gobs or yield area may be carried out in the copper mines only under specific conditions. Tests indicated the critical values of the residual solid width, at which the risk of strong seismic events is the biggest. Those are solid widths between 75 and 85 m and from 50 to 55 m. While mining the residues the R-UO/H one-phase method with roof deflection designated for regions with extremely difficult rock-mass conditions is used. The method assumes ore mining towards the gobs or yield zones (figure 15).

Figure 15. Diagram of R-UO/H mining method


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CONCLUSIONS In one-phase room-and-pillar methods used in copper ore mines, technological pillars separated on the mining face by headings and rooms, transfer into the post-failure condition (they are crushed by the exploitation stress of the rock-mass). These pillars of high deformability operate in the mining field revealing post-failure strength. Due to the tremor hazard occurring in the copper mines, the behavior of technological pillars (postfailure strength) must be utilized. In order to obtain the proper control of stiff roof in the room-and-pillar mining methods, its support during each phase of ore mining is very important. The decisive for the proper support is the cross-section area of the technological and remnant pillars, their slenderness and post-failure strength of the rocks being mined. Those parameters are selected for the specific conditions of each mining field. When the ore is mined under high hazard caused by dynamic symptoms of the rock-mass stress there is a possibility to reduce the hazard by leaving barrier pillars, stiff or yielding. Other factor important for room-and-pillar methods’ development in the copper mining industry is conditions of maintaining of stability of mine workings. However it is not possible to find any revolutionary solutions at present. Roof bolting support (resingrouted and expansive) wills still the basic one. When this type of support is used there is a barrier for widening the workings (span of uncovered roof). It appears that this technological boundary has already been reached in the Polish copper ore mines. In turn it limits the dimension of the underground equipment and machinery used. Other important issue concerns new technological solutions with regard to mining methods and possibility of considerable reduction of the rock-mass deformations by wide usage of flotation tailings and other waste in the form of hydraulic, solidified or paste backfill placed in the mined-out voids. REFERENCES Butra J., Kicki J. 2003: Evolution of Mining Technology for Copper Ore Deposit in Poland. School of Underground Mining Library, Krakow. Butra J., Kicki J., Siewierski St. 1994: The Solution of the Ore Deposit Mining Methods Technology in Poland. World Mining Congress. Session: “The Mining Industry on the Threshold of the XXI Century”. 10–14 September 1994. Sofia, Bulgaria. Kłeczek Z. 1994: Mine Geomechanics. Sląskie Wydawnictwo Techniczne, Katowice.

Possibility of Employing Mechanical Extraction in Thin Copper-Ore Deposit Conditions

Lubomir Horoszczak KGHM Polska Miedź S.A. O/ZG “Polkowice-Sieroszowice”. Lubin, Poland Leszek Ziętkowski CBPM “Cuprum” Sp. z o.o., Ośrodek Badawczo-Rozwojowy. Wrocław, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: The exploitation of copper ore deposits having a thickness of below 1, 5 m will require the use of mining machines capable of obtaining output as little diluted as possible, with production cost significantly lower than it currently is. Machines capable of mechanical excavating of relatively strong rock, in faces 1,1 m high are already known in the world. KGHM Polska Miedź S.A. “PolkowiceSieroszowice” mine has undertaken exploitation tests using a mechanical miner designed for winning of copper ores. KEYWORDS: Mechanical extraction, thin deposit, new techniques & technologies

1. INTRODUCTION Currently applied solutions in the field of excavating competent ores of widths lower that the height of machines currently used in traditional methods, have been extensively discussed and presented in various publications. For example in paper (Butra, Dębkowski, Kosior 2003) the authors have quite comprehensively presented techniques & technologies applied in the process of extracting copper ore deposits of low thickness. At the current state of knowledge, not one of the proposed and applied solutions eliminates the problem of output dilution, which is inherent to any particular exploitation method.

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The technical development in mechanical mining may indicate a significant turn around in the field of limiting dilution level and an effective decrease of useful metal ore winning costs in underground mines characterized by specific deposit and temperature conditions. 2. TECHNICAL SOLUTIONS POSSIBLE FOR IMPLEMENTATION For many years KGHM Polska Miedź S.A. “Polkowice-Sieroszowice” mine has been active in carrying out studies and searching for machines and technologies which would enable applying mechanical winning techniques to copper-bearing rock. Currently at least two companies in the world offer their solutions, giving hope for an effective implementation of mechanical winning in the specific conditions of thin deposit. These are: – DBT, – Sandvik—Voest Alpine Bergtechnik. 2.1. Hard rock cutting machine and mined ore transport system by DBT Hard rock exploitation tests with the use of a mechanical miner aimed at achieving the expected technical & economic parameters are carried out in one of the mines located in the Republic of South Africa. A view of the machine is presented in figure #1 and a diagram of the working mechanism and broken ore transport unit is presented in figure #2. In the author’s opinion (Walter von der Linden 2003) the solution being tested will allow mining of rocks with compressive strength of up to 300 MPa.

Figure 1. Hard rock cutting machine


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Comparison table of planned & obtained extraction results Current results with 400mm diameter drum, with 1 line of bits. – Cutting depth per one bits line: 20–25 mm. – Speed of advance: 2 m/min. – Performance: 0,06 Mg/min=3,6 Mg/h. – Bit cost (wear): 5 EUR/Mg. Projected production performance in 0,9 m thick reef based on the results of the trial. – Cutting depth with 3 bits lines: 60–75 mm. – Speed of advance: 4–5 m/min. – Performance: 0,09 Mg/min=54 Mg/h. – Bit cost (wear): 2–3 EUR/Mg. A characteristic feature of this solution is applying mining tool activation system and running the continuous miner on a conveyor, analogically as in coal mining. The authors have named this type of exploitation a “short wall”. The proposed application of this technique is presented in figure #3.

Figure 2. Layout of work of the continuous miner and a mined ore transport system

Figure 3. “Short wall” mining and view of the cutting head


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This is a technologically mature solution, which has been proved in real mining conditions of winning of ore-bearing rocks. Constant continuation of work aiming at the improvement of the efficiency, bit durability, and possible modifications to the concept or construction of the tools, with which the cutting organ will be armed in the future, may in a significant way contribute to an economic success of this method. As mentioned before, the main criteria having a decisive role in the success of this solution are its effectiveness and cost of tools. The application of standard cutting bits on the head and their relatively quick wear may be the reason for difficulties in achieving the main aim, which is a lower unit cost of mining 1 Mg of ore while using the new winning technique in comparison to the one currently applied. An essential matter in this case is also the effective winning time. With rapid wear of tools, the time necessary for their replacement may have a meaningfully impact on proportion of disposable time spent on effective work. 2.2. Narrow Reef ARM 1100 ore cutting miner by Voest Alpine The machine presented in figure #4 is the result of cooperation between two companies: “Sandvik—Voest Alpine Bergtechnik” from Austria and “Lonmin Platinum” from the Republic of South Africa. For the specific deposit conditions present in the Bushveld complex a head equipped with 6 cutting tools shown in figure #5 was applied. Testing has consistently been carried out for two years. In the author’s opinion (Jagiełło 2003) the continuous miner has been quite effective in excavating rock having width of 1,1 m and compressive strength of 180 MPa. According to the information provided by the machine manufacturer, another improved model of the miner will be delivered to the mining facility in the current year.

Figure 4. General view of ARM 1100 miner


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Figure 5. Cutting equipment The significant difference between the solutions offered by “DBT” and “Sandvik—Voest Alpine Bergtechnik” apart from the main idea, is the type of tool applied. In case of the ARM 1100 asymmetrical discs have been applied. This way the cutting technique has been replaced by a technique of undercutting by discs. In figure #6 the diagram shows the commonly known differences between symmetrical and asymmetrical discs (Hinterschneidttechnik. Wirth maschinen und bohrgeäte—Fabrick GmbH). This winning method differs quite significantly from the standard one based on the use of cutting tools.

Figure 6. Differences in impact of symmetric and asymmetric disk on excavated material


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3. MECHANICAL EXTRACTION OF THIN COPPER DEPOSIT TESTS KGHM Polska Miedź S.A. “Polkowice-Sieroszowice” mine prepares to undertake testing activities of winning copper ore deposits with the use of the ARM 1100 miner, equipped with tools adapted to the specific conditions of thin ore beds present in the mine (Collective work 2003). Analysis of the copper ore reserves in the mining field of this section show that deposits having thickness in the range of 0–1,5 m occupy around 37% of the area for which the reserves are calculated and make up for about 19% of deposits in total. Simultaneously, around 60% of Cu in ore deposits is tied up in thin beds of “Polkowice-Sieroszowice” mine, which means that to increasingly greater extent, it will be necessary to exploit exactly those deposits. Most of them are characterized by thickness of 0,4–2 m. In order to carry out field tests a mining plot located in the SW-1 area, between the A4 and A-1/6 inclines, and A-106/6 and W-109 galleries, has been chosen. The deposit in this area is characterized by thickness of around 0,4–2,4 m and Cu content of around 4,37%. The exploitation will be carried out using a short-wall system, with the total face length of about 50 m, individual panel length of about 4 m, and mining width at about 1,1 m. A two-stage exploitation test is planned: Stage I Preliminary, the layout is presented in figure #7—will serve as a verification of capability to win the ore seam with the aim to determine technical and cost parameters of the process. Exploitation will take place at the A-1/6 incline floor level and a BOA feeder will remove output.

Figure 7. Stage I—initial tests with continuous miner


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Figure 8. Winning test—Stage II Variant I Stage II The experience gathered during the experiment conducted in Stage I will allow making a decision on its continuation to stage II, according to one of the following variants: – Stage II—Variant I—on the floor of the A-1/6 incline (figure #8). – Stage II—Variant II—with short-wall location behind the safety pillar on the A-1/6 incline floor level (figure #9). – Stage II—Variant III—under the roof of an earlier excavated haulage road, with leaving a safety pillar (figure #10).

Figure 9. Winning test—Stage II Variant II


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Observations of the winning and output haulage processes and rock-mass behaviour are planned to be conducted during the experiment. Simultaneously, the operational capabilities of the miner will be verified, with the focus on the miner’s ability to change its direction of movement. Observations will also include rock-mass behaviour and any potential hazards inherent to the work of machines and other equipment during the process of mechanical winning. Hazards prevention systems and procedures in case of emergency situations caused by mining and mechanical conditions are also foreseen.

Figure 10. Winning test—Sage II Variant III 4. CONCLUSIONS Having in mind the information presented above it must be stated that a possibility of applying mechanical extraction for exploitation of thin copper ore deposits most surely exists. The technical success, in the field of effective excavating of competent rock by a miner equipped with correctly chosen cutting tools, may lead the way to a fundamental change in the current technology of extracting ore-bearing rocks. Implementation of this technology in thin deposits (face height-H≤1,1 m) should allow achieving the following: a) Eliminate the necessity for waste rock to be extracted, and subsequently located in gob area, to make room for machinery in thin deposits with widths smaller than the height of currently used machines. b) Eliminate the use of explosives. c) Eliminate the use of roof bolts at the mining face. The machine is remotely operated. It is possible to implement the solution proposed by DBT.


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d) Considerable improvement in the quality of output ore (mining width as close to deposit width as possible). e) Broken ore size distribution allowing to practically eliminating the process of preliminary ore crushing. f) Possibility to more effectively organizes daily disposable time (no blasting fumes). g) Possibility to meaningfully reduce the number of machinery running on rubber tires. h) Optimisation of workforce employed at machine operation. i) Meaningful reduction of fuel consumption, oil-derivative materials as well as machine spare parts (exploitation materials & others). j) Reduction of expenses on environment protection (reduction in exhaust and blasting fumes emissions). k) Possibility of resolving various difficulties of mining in hot environment. The most important is possibility of using the miner while exploitation descends to deeper ore deposits. l) Reduction in demand for ventilating air. Currently for engines to keep them running, an adequate amount of air must be supplied. m) Better conditions for ventilating the mining section. n) Evident improvement of working conditions (improved air quality). o) Possibility of machine construction solutions allowing mining at greater widths. All of the above listed arguments connected with the new winning system should allow the new technology to be economically competitive in comparison to the technology currently in use, but most particularly for the reduction of copper ore mining cost. A constant search for possibilities of more effective work is the fundamental condition under which most companies in our business, including KGHM Polska Miedź S.A. may survive and develop further. The degree of commitment and engagement in the search for new solutions allows saying that the intended aim will be reached. Short-wall mechanical mining system carries unsolved till now questions as to the rock-mass behaviour, which will require adequate elaborations, expertises, and examinations. This so very important problem, from the safety point of view, of roof control and rock mechanics, should be taken into consideration in the existing ore bed exploitation technology using mechanical mining. REFERENCES Butra J., Dębkowski R., Kosior A. 2003: The Way of Low Thickness Ore Deposit Exploitation in the Aspect of Impoverishment Limitation. Proceedings of School of Underground Mining. Krakow 2003. Walter von der Linden 2003: New Technologies for Mining of Tabular Low Vein Hard Rock Mineral Deposits. III International Conference Mining Techniques. Krakow—Krynica 2003. Jagiełło A. 2003: ARM 1100 Continuos Miner for Low Thickness Ore Deposits. III International Conference Mining Techniques. Krakow—Krynica 2003. Hinterschneidttechnik. Wirth maschinen und bohrgeäte—Fabrick GmbH. Collective work 2003: Elaboration of Conception for Carrying Out Tests of Deposit Winning by Means of Short Wall System with Using Continuous Miner in the Conditions of “PolkowiceSieroszowice” Mine. CBPM CUPRUM. Wrocław 2003.

Experience and Practical Aspects of Utilizing a Shrinkage Method of Extraction at “Kazimierz-Juliusz” Coal Mine in Sosnowiec

Stanisław Gajos Katowice Coal Holding S.A. Katowice, Poland Marek Urbaś, Tadeusz Lamot KWK “Kazimierz-Juliusz” Sp. z o.o. Sosnowiec, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: In the paper the authors present experience gathered during exploitation of the thick and steeply dipping 510 coal seam by means of a shrinkage method using exploitation galleries for seam development at “Kazimierz-Juliusz” coal mine (KWK “Kazimierz-Juliusz” Sp. z o.o.) in Sosnowiec. This particular method of mining thick and steep-dipping seams was proposed and developed by a team of engineers from the Mine and scientists from AGH—University of Science and Technology in Krakow. The method was employed to mine this part of the 510 seam located within the boundaries of “Kazimierz-Juliusz” mine, whose thickness exceeded 20 metres and which occurred at a dip of 35 to 45 degrees. The notes presented here refer to this experience. The paper covers the topics of influence of the described mining system on neighbouring excavations as well as on the rock-mass in the area of exploitation, taking into account mining hazards involved. It presents the obtained up-to-date technical and economic parameters characterising the system. KEYWORDS: Exploitation system, shrinkage, thick seam, steeply dipping seam

1. INTRODUCTION The mining area of “Kazimierz-Juliusz” mine covers an extension of the Bytom basin, and its deposits occur in a shape of basin as well. In its northern and southwestern part

Experience and practical aspects of utilizing a shrinkage method of extraction

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the deposit occurs at a dip of 30°. To southeast, where the so-called M-3 section is located, the deposit occurs at a dip exceeding 45°. In those areas, where the thick 510 seam dips does not exceed 30° it is mined in several slices from top down with the use of a longwall system, mainly with roof-caving. Between ten and twenty million tons of proved reserves is locked up in the M-3 part of the seam, dipping at around 45°. Due to the “easier” reserves becoming depleted it became necessary to reach for these. The mining team looked for a system of exploitation of thick and steep-dipping seams that would meet the two following basic criteria: – secure safety of conducting the mining operations, and – be economically viable. Systems used in the Polish mining industry in the past, like “Miechowicki”, “Jankowicki” or “Kazimierzowski” do not meet the above criteria. Because of this several concepts of mining the steeply inclined and thick M-3 section of the 510 seam, were developed by the Mine in co-operation with the scientists from the AGH in Krakow. One of the proposed methods was shrinkage with the use of exploitation galleries, with roof caving, and field trials were started in May 2003. 2. MINING AND GEOLOGICAL CONDITIONS OF THE 510 SEAM IN THE AREA OF EXPLOITATION 2.1. Location of the M-3 section The M-3 section is located in the eastern part of the mining area of “Kazimierz-Juliusz” mine between Dorota fault, which has a throw of between 100 and 180 m, and Jakubowski fault with a throw of 270 to 5 m. 2.2. Geology The 510 seam in its M-3 part occurs in a shape of basin at a depth from about 500 m at the bottom of the basin to approx. 50 metres where it outcrops in the south. The M-3 area is divided into southern and northern parts by a fault having a throw of 25 to 40 m. The dip of the northern part changes from 0° to 30°, whereas in the southern part, where the mining system in question was introduced, it locally increases up to 50°. The 510 seam’s width in this area is variable. To the north it is 14 to 20 m thick, to the south it thins-out to a few metres at the outcrop, whereas to the east it is also thinner due to a wash-out, even to the point of becoming subeconomic. Tectonics of the deposit in the M-3 section is extraordinarily diversified. There are several faults of various throws and strikes. Their specific characteristic is high inconstancy of throw, changing with strike by a few and sometimes even more than ten metres. To the east Dorota fault runs, its throw of 100 to 180 m. There is Jakubowski fault to the west. It strikes east and its throw is between 270 and 5 m. In between there are several other faults with throws ranging from 1 to 40 metres and various strikes. Stratigraphy of the region consists of:


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– Quarternary deposits consisting of a layer of sand with mudstone interlayers, 11 m thick in total, – Carboniferous deposits consisting of: 1. “orzeskie” strata formed as shales and sandstones with coal layers, of which only 344 and 357 seams are of industrial value, 2. “rudzkie” strata, with predominantly sandstones in the bottom part and clay shales in the upper and only one coal seam (404) of industrial value, 3. synclinal strata represented by the currently mined 510 seam. The immediate roof of the M-3 part of the 510 seam is formed by intermittent, 1 to 8 m thick, layers of clay shales and arenaceous shale. Higher up, approximately 20 m above the roof of the seam, a 12,7 m thick layer of sandstone occurs. Another sandstone layer, 37 m thick, occurs approx. 45 m above the roof. Intermittently deposited layers of clay shale and arenaceous shale in general, form the foot of the seam. 2.3. Mining of the M-3 section to date No other coal seams, above or below the 510 seam in the M-3 area have been mined to date or are planned to be mined in future. Extraction by a longwall system with roof caving was started in the area in question, at its northern limit, in 1993, from the top layer of the seam. The same system is currently used in this area and in the middle part of the M-3 region for exploitation of slice IV (second from the top). In the southern wing the steep-dipping 510 seam has been divided into A, B, C, D and E mining fields (up to +50 level). Extraction by the shrinkage method with the use of exploitation galleries is currently under way in mining field A, whereas intensive development is done in order to start exploitation in mining field B. 3. DESCRIPTION OF THE SYSTEM 3.1. Method of extraction of the seam In this system panels are formed by dividing a thick and steeply dipping coal seam into diagonal layers, of thickness equal to that of the seam, by planes perpendicular to its roof and foot. Exploitation of the seam consists in mining the panels, starting from the uppermost, by extracting the coal contained in blocks above exploitation galleries by shrinkage with roof caving—figure 1.


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Figure 1. The way of dividing a seam into panels and extraction sequence 3.2. Development of the seam The mining section in which extraction of a thick and steep-dipping seam is to take place should be developed from the main level by means of a crosscut, from which a raise is to be driven on the seam’s footwall. The raise is the main development drive and should be connected to the ventilation system of the area. Slightly inclined (approx. 10° up) exploitation galleries are then driven on the footwall of the seam, from the main raise into the panels. Inclination of the tunnels stems from the necessity to facilitate drainage of any water, which could enter the workings from gob. The distance between the galleries measured on dip depends on the seam width and should approximately be equal to it. The galleries are driven of rectangular cross-section due to the shape of production face equipment. The exploitation galleries are approx. 4, 5 m wide ×3,0 m high and their support consists of steel props “Valent” and V-shaped steel beams. The sidewalls of the tunnels are supported by mesh and timber anchors. Figure 2 shows the way to support an exploitation gallery developed by a continuous miner.


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Figure 2. Support of an exploitation gallery developed by a continuous miner 3.3. Face equipment in an exploitation gallery In a gallery prepared in this way face equipment is installed. Its main components are: a Mechanised Support Set which reinforces the existing steel support of the tunnel and secures the working face, and a scraper conveyor for the removal of extracted coal— figure 3.

Figure. 3. Basic face equipment The Mechanised Support Set is made up of two units of the BMV-10 longwall mechanised support manufactured by BME Novaky Company from Slovakia, installed symmetrically along the tunnel’s axis. This type of support has caving shield constructed of units, which can be lifted, and so allows drawing coal broken during the shrinkage operations. Additionally, the units are equipped with liftable shields to support roof between the units, and with sliding plates and liftable side shields to secure the working


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area from the tunnel’s sidewalls. The liftable shields installed on the units’ bases form the scraper conveyor’s sideboards at the time of coal transport, and are used as working platform during the units’ disassembly. In view of the above the main objectives of the Mechanised Support Set are to reinforce the tunnel’s support and in doing so protect the working area, and facilitate safe drawing of coal onto the scraper conveyor, which is installed between the units of the Set. The second very important piece of equipment is the PZ-1000 scraper conveyor, which was constructed and manufactured by NOWOMAG Company specifically for this exploitation system. The conveyor must be able to meet the requirements of the very demanding working conditions, caused mainly by the necessity of the return sprocket to work under caved coal. The working face equipment set includes also two VPS-01 portable jack-supported rockdrills manufactured by a Czech Company InterCUPRO, and a blasthole charger fitted with a PVC hose, to improve the efficiency of blasting operations. 3.4. Winning of coal in the panel Mining operations in a panel consist in breaking coal by blasting shotholes drilled in a fan pattern and subsequent gradual drawing of the broken coal, together with cyclically moving the Mechanised Support Set and the scraper conveyor forward in the exploitation gallery. Simultaneously the steel support sets of the tunnel are removed and roof caving is induced by blasting if necessary—figures 4 and 5.

Figure 4. Winning of coal in a panel The moving face forms mining front along the width of the panel and extending over the whole width of the seam. As mining progresses the length of the exploitation gallery gradually decreases.


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Figure 5. The way of drawing the broken coal

Figure 6. Caving-inducing blasthole pattern


185

The void created by coal removal is gradually filled by caved roof rock. Blasting shotholes drilled ahead of the face, as shown on figure 6, induces caving. In some favourable conditions, if necessary, the void may be additionally filled with hydraulic backfill or water and rock dust mix pumped through boreholes drilled from other accessible workings. The rate of advance per cycle is 1 to 2 m, and depends on mining and geological conditions. 3.5. Removal of output Output is removed from the face by a scraper conveyor installed in between two support units, with the return sproket left in the caved area. Thereafter output is transported by a system of scraper and/or belt conveyors installed in the exploitation gallery and other haulage roads right to a loading point located in the drift on the main level. 3.6. Material transport Material is delivered to the face from the side of the relevant ventilation level (used for material transport as well) by means of overhead monorail, rail or wheel transport. In general the face does not require any material, with the exception of spare parts for the face equipment in case of repairs. It is necessary, however, to take out the component parts of the gallery prop support as they are removed with the advancing face, in order to re-use them in a gallery in course of development. 3.7. Ventilation Ventilation of exploitation galleries during mining is done by means of an independent exhaust pipe ventilation system. It is also possible to use a force or combined forceexhaust system. 3.8. The conditions and coordinating of conducting mining and development works as stipulated in the permit of the President of WUG (Safety and Health Inspectorate) Due to the fact that the 510 seam is classified as seismically hazardous, it was necessary to obtain a permit to deviate from the provisions of § 334 of the Instructions of the Minister of Industry dated 28 June 2002, dealing with the safety and health of conducting works and fire-hazard provisions in underground mines, in order to be able to conduct mining operations by means shrinkage from exploitation galleries in this area. The permit obtained from the President of WUG, which was issued based on recommendations of the Committee for Seismic Hazard in Coal Mines and the Committee for Mine Workings and Roof Control, stipulated, among others, the following provisions for mining coal with the use of shrinkage from exploitation galleries method: – This mining method was deemed to be a method undergoing field trials.


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– Extraction of the seam was divided into two stages: the first stage was to be limited to mining the least hazardous areas, as far as seismicity is concerned (I class of seismic risk), the second stage concerned mining the area belonging to the III class of seismic risk. – The second stage of exploitation may only commence after experience is gained during conducting the works of the first stage, and if favourable opinion of the abovementioned Committees is obtained. – New exploitation galleries may be driven at a vertical distance of not less than 25 m from the gallery where shrinkage extraction works are under way. – An Expert Group was called into being for current analysis of technological problems and work safety during the mining works. – Frequency and scope of inspections of the implemented safety measures and adherence to the provisions prescribed in the obtained permit, to be done by the Mine’s management and mining officials, was established.

4. THE EXPERIENCE AND PRACTICAL ASPECTS OF IMPLEMENTING THE SYSTEM The experience gained during mining of the thick and steep-dipping 510 seam in the M-3 area by means of shrinkage done from exploitation galleries, together with some practical notes, are presented below with respect to several criteria, starting from safety hazards and ending with economical results. 4.1. In respect to the influence on seismic hazard and roof control The 510 seam is classified as belonging to: – III class of seismic hazard, in its not de-stressed part at depths greater than 400 m, – I class of seismic hazard: — at depths not exceeding 400 m, — in effectively distressed areas. The area of the 510 seam where extraction by shrinkage was conducted to date belongs to I class of seismic hazard. The seam is approximately 20 m thick, occurs at a depth of approx. 300 to 330 m, dips at approx 40°, and is composed of several types of coal: laminated semi-shiny, shiny with carbonates and, locally, semi-mat. The coal from 510 seam is characterised by: – coefficient of compactness f of 0,91 to 1,13, i.e. easily extractable, – uniaxial compressive strength Rc between 13,4 and 35,7 MPa, with the mean value of approx. 20,9 MPa, – energy index of proneness to seismicity WET from 1,86 to 5,81, with the mean value of 4,73, i.e. low proneness to seismicity. In the geological structure of roof and foot, sandstones of varying grain size, mudstones, claystones and coals can be identified.


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Percentage-wise the share of the respective rock types in the strata from 30 m below the foot up to 100 m above the roof of the seam is as follows: sandstones—39%, mudstones—23%, claystones −34%, coals—4%. Stratigraphically, up to 100 m above the roof of the 510 seam, there are 3 layers of sandstones of width greater than 1 0 m, occurring approximately 20, 47 and 73 m above the roof of the seam. Average uniaxial compressive strength of these sandstones is approx. 50 MPa. Pronounced jointing occurs in some areas of each strata. The above given strata widths and their respective distances are subject to minor fluctuations within the mining area. The mean strength of the total batch of roof (100 m) and foot (30 m) strata is approximately equal to 45, 0 MPa, with individual layers’ Rc from 11,0 to 108,8 MPa. The rocks surrounding the seam have to be considered as very prone to seismicity. The “roof-foot-seam” structure is prone to seismicity. However, up to the mining depth of 400 m, probability of a seismic event occurring is small. 4.1.1. Observation and evaluation of seismic risk The evaluation of seismic hazard during mining operations was done by a composite method. The results of evaluation of seismic hazard by individual component methods were as follows: – Mining evaluation method—class “b”. – Small-diameter borehole method—class “a”, max. output of drillings—3, 5 [l/m]. Small-diameter drilling was done once every 24 hours. 10-metre-long boreholes were drilled into both sidewalls of the exploitation gallery, 5 to 30 metres ahead of the production face. The holes in the southern sidewall were drilled parallel to strata and in the opposite sidewall they were drilled horizontally. – No places with increased state of stress were discovered. – Seismo-acoustic method—class “a” and locally class “b” with limited seismic activity. Seismo-acoustic observations were done using two geophones installed in both sidewalls (one geophone in each) ahead of the face. The distance of geophones from the face was kept at 30 to 70 m. The observed low seismo-acoustic activity, also after blasting, shows that no energy was stored in the rock-mass. – Seismic method—class “a”, there were no spontaneous seismic events. The area in question is monitored by a total of 8 seismometer stations. The three closest stations arranged in a triangle are situated not farther than 300 m from the conducted mining works. This enables to register events of smallest energies (magnitude 102 J). The other 5 stations, located from 800 to 2300 m away, serve to calculate energies of any potential events of greater magnitude. The stations of the seismic network covering the lease area of the mine are located at various depths (from 200 to 700 m), which allow specifying the vertical component as well. Based on results of the composite method, the evaluated existing seismic hazard level is “a”, i.e. not hazardous. 4.1.2. Roof control


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The method of roof control utilized in the described mining system is to allow full roof caving in the worked-out area. In the first stage the caving zone encompasses coal, which is drained out as output, then it extends into the roof strata of the 510 seam. Geo-mechanical properties of coal and rocks occurring in the roof of the 510 seam in the panel where exploitation was planned (alternately arenaceous shales and claystones), indicate relative ease of inducing and managing roof caving. However, it is necessary to induce caving as soon as possible by blasting. Blasting approx. 50 metres long shotholes induces caving of roof strata. Shothole diameter is 75 mm. The blasting pattern is approx. 4×6 m. Each hole was charged with approx. 125 kg of explosives. Caving inducing blasting is conducted from exploitation galleries by means of “barbaryt” explosive. All blasts are recorded by the mine’s seismic network as tremors. Energies of these blasts—tremors depend on the volume of the explosive charge set off and were in the region of from 2 to 6×104 J. The conducted blasting works cause the rock-mass to fracture. Their effectiveness was proven by stratascope inspections. 4.2. In respect to the influence on neighbouring excavations Magnitude and extent of expected influence of the planned exploitation on other excavations in the area, i.e. the exploitation gallery from which mining was done as well as other exploitation galleries being at different stages of development, and the main ventilation-research raise, was subjected to an analysis already at the exploitation planning stage. Prof. Zorychta, in his work dealing with influence of the planned exploitation on seismic hazard, established, among others, a safe distance from a working face to the next exploitation gallery being driven below as 25 m vertically. This condition resulted in, still valid, necessity to have at least one developed and established exploitation gallery between the working face and a gallery in process of development. Additionally, to quantify influence of the mining works on neighbouring excavations, convergence-measuring stations were installed in the excavations closest to the working face. Insitu measurements were taken along with detailed visual inspections of condition of support in each working in the area. Employing the abovementioned methods to assess influence of mining on neighbouring excavations allowed coming to the following conclusions, as to the extent the shrinkage mining influenced neighbouring excavations to date: – the influence on support in production galleries is negligible, only in the zone directly adjacent to blasting operations (in front of the Mechanised Support Set), some local fracturing of roof, and resultant increase of load on support, was observed. However, this phenomena was already taken into account during the planning stage and this was one of the reasons for the tunnel’s support in this section to be strengthened, – a maximum of 3 cm of closure was observed in the gallery situated immediately below the production one, but there was no influence whatsoever on support of the tunnels situated further down, – tremors caused by blasting operations did not have any visible influence on the condition of support, only some local slabbing of sidewall (from the side of the conducted mining) in the tunnel situated below was noted, but its extent did not pose any danger to the tunnel’s stability,


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– in the sections of the exploitation galleries immediately adjacent to production faces and supported by yielding arch supports, no influence of the exploitation was observed—no yield on the bolted clamps, not even any compression of timber cribbing. Also the condition of the ventilation-research raise, which is mined towards by each shrinkage panel before they stop approx. 40 metres away, did not change. To date no influence of the conducted mining works and geological conditions on the stability of support in the neighbouring exploitation galleries as well as in the ventilationresearch raise was observed. 4.3. In respect to the influence on level of fire hazard The value of coefficient of proneness to spontaneous combustion determined for the coal occurring at the Mine, and also in the area of interest, requires classifying the coal of the 510 seam to group V of proneness, i.e. very prone to spontaneous combustion. Fire-hazard monitoring was envisaged at the planning stage to be done by: – air-sampling done by fire early detection stations located in caving gob area, exploitation gallery at the intake of ventilation duct and in the ventilation-research raise at the crosscut intersection, – monitoring by the section’s mining and ventilation officials, as well as designated personnel (methane control officers). It was initially planned, in accordance with the mine regulations, to take air samples for analysis twice a week, but the frequency was subsequently increased to three times a week. The feared existence of fire hazard in the area, envisaged during the planning stage, did not materialize. The highest values of fire hazard indicators, calculated from the results of air-samples’ analysis for exploitation galleries 1A, 2A and 3A were, correspondingly, as follows: – carbon monoxide increase index ∆CO—0,0024%, 0,0020%, 0,0018%, – carbon monoxide volume index VCO—9,2 l/min, 2,2 l/min, 3, 2 l/min, – Graham’s index G—0, 0032, 0,0020 and 0, 0013. According to the ventilation department personnel, the following factors exerted positive influence on the level of fire hazard in the area: – the use of an independent exhaust ventilation system, which greatly inhibits migration of air into gob, – setting off of large quantities of explosives at a time, which causes effective inundation of the air in the blasting zone with fumes. It has to be noted that the use of exhaust ventilation causes blasting fumes to be drawn out from gob, which may result in overrating the levels of fire hazard indicators. 4.4. In respect to the influence on methane-gas hazard


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“Kazimierz-Juliusz” mine is classified as gaseous, with the 510 coal seam belonging to class I (lowest) in the methane hazard classification over the whole lease area of the mine. Values of methane content of the 510 seam fall at the lower end of the scale of class I hazard level. Actually, as no methane is ever detected in air-samples taken in return air coming from both the mining sections and the levels, as well as in the upcast shafts, the absolute methane content may be considered as equal to zero. At the stage of planning the extraction described in this paper, methane hazard of a level approximately equal to that existing in other, previously mined, areas of the 510 seam was envisaged, i.e. class I at the most. In the area where the shrinkage method was employed, five tests of natural methane content using a direct method were done, and the results obtained were in a range from 0,0008 to 0,009 m3 CH4/Mg, which supported the above assumption. 4.5. In respect to the influence on coal dust explosion hazard All excavations in the area where the shrinkage mining method was used belong to class B in the coal dust explosion hazard classification system. Measurements of non-flammable particle content are done twice a week. Average rate at which coal dust accumulates in the workings where the shrinkage method of mining is utilized is close to the values observed in headings excavated by continuous miners. Because of the above the safety measures employed do not differ from the standard ones, prescribed by the regulations. 4.6. In respect to the achieved technical and economic results Technical and economic results achieved by using the shrinkage method do not differ from the ones projected at the conceptual planning stage, even though they are relevant only to this period of time when the exploitation is conducted as, and according to the rules of, field trial. The table presented below shows basic indicators, i.e. volume of output per production face and face and section working cost in the M-3 area. The presented data was collected from June till October, i.e. from the commencement of the exploitation until sending this paper to print.


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Table 1. Basic technical and economic indicators obtained by using the shrinkage from exploitation galleries method Economic indicator

5 544

6 216

4 536

9 901

10 210

277

283

227

450

537

Face Total, PLN working PLN/Mg cost

212 774,0

196 802,0

133 919,0

218 621,0

205 769,0

38,38

31,66

29,52

22,08

20,15

Section Total, PLN working PLN/Mg cost

331 009,0

347 894,0

253 437,0

378 527,0

401 755,0

59,71

55,97

55,87

38,23

39,35

output

tones

June 2003 July 2003 August 2003 September 2003 October 2003

Mg/day

Constant increase of output volumes together with decrease of working costs may be noted, which did and still does come as a result of gathering experience in conducting shrinkage mining operations and adjusting technology to changing working conditions. 5. CONCLUSIONS The assumptions taken when planning extraction of the thick and steeply dipping 510 coal seam at “Kazimierz-Juliusz” mine by shrinkage from exploitation galleries, were proven to be, in general, correct. Mining with the use of this method done to date allowed to gain more experience as far as extraction of coal seams occurring in difficult geological and mining conditions, especially related to width and dip, are concerned. On the other hand, the achieved technical and economic results are fully satisfactory to the Mine, particularly because ratio of coarse fractions in output coal obtained with this mining method considerably exceeds 60%. Moreover, the “Valent” props removed from production galleries as faces progress are re-used to support consecutively developed exploitation galleries, which contributes vastly to reduction of driveage costs in such a way that the costs systematically decrease as exploitation in the mining section continues.

Application of the SF6 Tracer Gas in Identifying Mine Air Flows through Abandoned Workings Sealed from the Ventilation System

Piotr Buchwald & Zbigniew Jaskólski Central Mines Rescue Stations. Bytom, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 The knowledge of the direction and velocity of flows of gases, including mine air, through abandoned workings, in particular gobs and sealed fire fields, excluded from the ventilation system as well as through rock fissures is very important in preventing spontaneous heating of coal and resultant fire hazard in collieries. A tracer test method using electronegative tracers proved to be very useful in such conditions. The tracers are electronegative gasses having the tendency to connect with electrons. Other methods, consisting in measurements of the average air flow velocity, show considerable measurement errors, are labour-intensive or prove to be technically unfeasible in the conditions of the areas in question. Sulphur hexafluoride (SF6) is currently applied in the Polish collieries as a tracer gas in the tracer tests aimed at identification of flows and leaks of gasses in mine air. In normal conditions the gas is characterised by chemical and thermal stability up to a temperature of at least 800°C and coal or other rocks do not adsorb it and its atmospheric background is negligibly low. Sulphur hexafluoride (SF6) is detected by means of chromatographic technique. The analyser is equipped with an electron capture detector (ECD) with sensitivity of the order of 1 ppb (i.e. 10−9 V/V), and in particular cases even 1 ppt (i.e. 10−12 V/V). SF6 is colourless, smell-less and not harmful to health. These properties are particularly useful when the gas is applied in areas where people are employed. The high vapour pressure of SF6 (2,4 MPa at a temperature of 25°C) enables to dose it directly from a gas cylinder. The practice of applying SF6 to measurements of the mine air and gasses in collieries shows that the volume of injections of the tracer gas should amount to 100 dm3 for each 100.000 m3 of an inspected area’s volume. The time of washing out of

Application of the SF6 tracer gas in identifying mine air flows

193

SF6 from the inspected part of the area, particularly in the case of a gob or a sealed fire field, may be relatively long. The tracer-gas tests are executed in the following stages: – Examination and preparation of the area to be inspected. – Sampling the air marked with the tracer gas. – Detecting the tracer. – Processing data resulting from the chromatographic analyses. – Interpretation of the processed data. Tests with the application of the SF6 tracer gas should be preceded by a careful examination of the area to be inspected. Because determining directions of mine airflows in an object is performed for given ventilation conditions it is necessary to make no changes in those conditions, i.e. ventilation layout, ventilation regulator adjustments or ventilation fans’ working parameters. A preliminary examination of the area to be inspected consists in an analysis of the possible directions and predicting volumes airflow. Determining directions of air leaks requires a careful analysis of the ventilation system and a detailed study of mining-geologic conditions. The examination should be performed with the help of appropriately skilled and experienced personnel from relevant departments of the colliery on the basis of: – Basic plans of the pertinent coal seams, mining districts, levels. – Potential ventilation plan or knowledge of the potential field. – A ventilation plan showing velocities and volumes of air flows. – Characteristics of an expected result. A detailed analysis allows working out a specific activity schedule that encompasses the following data: – The time of starting the test. – Places of tracer-gas injections. – Needed quantities of the gas and injection methods. – Sampling places. – Duration of the test. An injection place may be located at a point in the ventilation network where air leakage into the inspected area occurs, e.g. at a seal situated on the intake-air side. Sampling places have to be located in the part of workings into which the leakage air may migrate from the inspected area, e.g. at a return-air side seal. The frequency of taking samples depends on the distance from the place of injecting the tracer as well as the expected velocity of the tracer-gas migration through the inspected area on the route: the injection places—the sampling place. The frequency has to be determined in such a way as to capture the pocket of the air containing the tracer gas right at the sampling place. The previously mentioned conditions also determine the duration of the test. Air sampling should be stopped after the pocket of tracer gas has gone past a given sampling place. In general the time interval of sampling is scheduled to be several times longer than the one estimated from the expected velocity of the tracer in order to eliminate any possible errors that might have occurred when assessing the speed of its spreading.


194

Marking the air and gasses in a given area (i.e. a section of the ventilation network) consists in introducing the SF6 tracer gas, in a quantity determined during the preliminary examination, at a defined time at a point of air inflow into the area to be inspected. The method of dispensing (injecting) also depends on the quantity of the gas to be applied during an injection. The SF6 tracer gas is kept in steel cylinders. In the case of an injection of a volume of the order of 100 dm3 the gas may be injected directly from a steel cylinder by means of an appropriate injector with a capacity of 0,5 or 1 dm3/minute. SF6 should be prepared for an injection far away from the analyser (best of all outside the building) and far away from the area to be inspected and the downcast shaft of the mine. It is also necessary to prevent the gas sampling pipettes or squirts from getting into any possible contact with the tracer gas. During injecting the gas is necessary to assure that the injector is tightly connected to the cylinder as well as to ensure tightness of connection elements between the injector and the injection point. Right before injecting it is indispensable to make sure whether or not the point of injection is located inside a ventilation intake into the inspected object. In this case a water-filled U-tube pressure gauge is a suitable device. An injection should be performed at a strictly determined time after a previous check by the telephone that the sampling workers participating in the test are present at the determined sampling stands. Each sampling worker is obliged to take one sample at his or her sampling stand right before injecting the SF6 tracer gas into the inspected object. The samples are called “zero samples” and their purpose is to confirm that there is no SF6 content in the mine air or to affirm the existing background level. The sampling workers should have ballpoint pens or pencils and paper to write down descriptions of the samples as well as a watch to be in a position to record the times of taking samples. It is allowed to take the samples by means of polyethylene or polypropylene pipettes with a cubic capacity of 10 cm3 or 20 cm3. Those pipettes should be provided with paper labels stuck on their outside surface (to enable the sampling workers to make descriptions of the samples) as well as the inlet pipes of the squirts should be tightly closed before the test. It is very important that the gas sampling pipettes used for the test are perfectly tight. While taking a gas sample by means of a squirt pipette each sampling worker should scavenge the pipette with the air to be sampled three to five times and then take the sample and precisely plug it. Immediately after taking the sample the worker should make a description that encompasses the number, symbol or name of the pertinent sampling stand, as well as the date and time of taking it. The pipettes with samples should be transported (and stored) in openwork sacks, bags or they should be tied together so that they form bundles. It is not allowed to store them in plastic-film sacks or bags. The samples so prepared should be delivered to a laboratory at an appointed time. Detecting the SF6 tracer in the taken samples of gasses and air is performed by means of an electron capture detector (ECD) chromatograph. Such a chromatograph is usually located in a laboratory yet in emergency cases (e.g. a fire) it may be placed underground in a mine in order to speed up the transfer of information about the tracer concentrations. Based on the chromatograms of individual samples of the air and gasses, a table or a plot of changes in the tracer peak height as a function of tracer gas concentration in time from the moment of its injecting is drawn up for each sampling point. The document should also encompass basic data on the method and the course of the test. Tables or


195

plots being a result of working out output data from the chromatographic analyses must be supplemented with a description of the ventilation-geologic conditions of the inspected area. Absence of the tracer at a given point indicates that there is no migration (flow) from an injection place to a sampling place. An affirmation of the tracer gas at a given point requires finding the cause of its presence. Based on the appearance time of the tracer, its maximum concentration, and its decay it is possible to derive the velocity of the airflow from the injection place to the sampling place. Results of a tracer-gas test should be interpreted with the help of the mine’s ventilation staff members. Data obtained by means of the tracer-gas test may be: – Qualitative—determining existence or absence of flows or migration of gasses, therein air, in the inspected area. – Quantitative—determining ventilation parameters of the inspected area. – Structural—enabling reconstruction of the geometrical layout of the gob ventilation network, which undergoes continuous deformation (closure) processes taking place during extraction operations. The tracer-gas test may prove very useful in solving difficult and complex mine ventilation problems. In particular it may be applied to: – Measuring the velocity of gas flows, and of mine air flows, in mine workings, gobs, fissure leaks, ventilation holes, and methane-drainage boreholes. – Testing air-tightness of rock-mass and gob. – Testing air-tightness of ventilation seals. – Control and measuring of airflow changes in time. The tracer-gas technique is relatively simple from the technical point of view but in practice happens to be difficult in the sphere of interpreting results. Several cases of applying SF6 to tracer-gas tests in the conditions of mine ventilation systems are presented below. CASE NO. 1 In the Colliery R the longwall 619/620 g (figure 1) reached its mining limit and was in the process of being closed down. Because an increase in carbon monoxide concentration in the longwall’s gob


196

Figure 1. Spatial ventilation plan of a district in the safety pillar for the town of Bytom—level 660 m had been confirmed and it began to leak into the active workings, contaminating the circulating air current (at a rate of about 10 litres of CO per minute) it was decided to start preventive actions. They consisted in supplying ashes into the gob of the longwall 619/620 g. Moreover the longwall face was temporarily isolated by means of the seals TI-116 and TI-113. Additionally, nitrogen was forced into the gob behind the seal TI116. In order to determine how the nitrogen migrates within the area of the longwall face and to assess the nitrogen’s effectiveness the SF6 tracer gas was injected in a quantity of 100 dm3 into the pipeline supplying nitrogen into behind the seal TI-116 located in the incline Ia. Air samples were taken in the following places (figure 1): – From a 185-mm borehole, drilled from the packed-out road 1 into the road used to disassemble the longwall 619/620 g. – From behind TI-112—erected in the incline II. – From behind TI-113—erected in the packed-out road 2. – From behind TI-109—erected in the incline leading to settlers. – From the caving-gob part of the longwall 621/622g area, at the longwall’s tail side. – From behind TI-525—erected in the ventilation road in the seam 507. Samples were taken during the A shift at a frequency as follows: – Twice before injecting SF6 (background). – For the first hour—every 15 minutes. – For the second hour—every 30 minutes.


197

– For the remaining time—every 1 hour. MEASUREMENT RESULTS 1. The SF6 tracer gas was confirmed in the samples taken: a) From the 185 mm borehole, drilled from the packed out road 1 into the disassembly road of the longwall 619/620 g. The tracer gas reached the sampling point after 20 minutes from the beginning of injecting it and was present for the whole time of observation, i.e. till 1415 o’clock. The highest concentrations were noted from 915 o’clock to the end of the observation time (the top concentration took place at 1415 o’clock and amounted to 4590 ppb). b) From behind TI-112—erected in the incline II. The tracer gas reached the sampling point after 65 minutes from the beginning of injecting it and was present during the whole time of observation. The highest SF6 concentration occurred at 1315 o’clock and amounted to 1887 ppb. c) From behind TI-113—erected in the packed out road 2. The tracer gas reached the sampling point after 20 minutes from the beginning of injecting it and was present till the end of observation. The maximum concentration of SF6 was observed at 1415 o’clock (the last measurement), and amounted to 282 ppb. d) From the caving-gob of the longwall 621/622g area, at the longwall’s tail side. The tracer gas reached the sampling point after 95 minutes from the beginning of injecting it and was present till the end of the observation time, i.e. until 1415 o’clock. The highest SF6 concentration was measured at 1315 o’clock and amounted to 3, 2 ppb. e) From behind TI-525—erected in the ventilation road in the seam 507. The tracer gas reached the sampling point after 20 minutes from the beginning of injecting it and was present there all the observation time. The highest concentration took place at 1315 o’clock and amounted to 17 ppb. 2. No presence of the tracer gas was observed in samples taken: a) From behind TI-109 erected in the erected in the incline leading to settlers.

Table 1. Results of the analyses of SF6 content Sampling place Sampling Sampling date time 15th July 2003

Behind TI-113 The 185 mm borehole Behind TI-112 erected in the drilled from the packed-out erected in the packed-out road road 1 to the disassembly road incline II 2 of the longwall 619/620 g Concentration of SF6 (ppb)

Concentration of SF6 (ppb)


7:15

0,0

0,0

0,0

7:15

0,0

0,0

0,0

8:15

0,0

0,0

0,0

8:30

35

0,0

1,0


198

8:45

43

0,0

29

9:00

135

0,0

45

9:15

900

41

83

9:45

1900

219

81

10:15

3350

590

57

11:15

4560

1435

83

12:15

4360

1294

77

13:15

4510

1887

252

14:15

4590

1384

282

Table 2. Results of the analyses of SF6 content Behind TI-109 erected in the incline leading to settlers

The caving-gob of the longwall 621/622 g at the longwall’s tail end

Behind TI-525 erected in the ventilation road in the seam 507

Sampling time




7:15

0,0

0,0

0,0

7:15

0,0

0,0

0,0

8:15

0,0

0,0

0,0

8:30

0,0

0,0

7

8:45

0,0

0,0

11

9:00

0,0

0,0

13

9:15

0,0

0,0

16

9:45

0,0

0, 7

15

10:15

0,0

1, 3

16

11:15

0,0

1,3

15

12:15

0,0

2,6

16

13:15

0,0

3, 2

17

14:15

0,0

2,8

16

Sampling place

Sampling date

15th July 2003

CASE NO. 2


199

The objective of the exercise was to confirm or exclude the existence of a ventilation connection between adjacent collieries M and W (figure 2). On 30th October in the colliery M about 110 dm3 of the SF6 tracer gas were injected into a suction hole drilled into a gob part of the seam 418 on the 500-m level (the gate road adjacent to the gob of the longwall 304). The pertinent gob portion of the seam 418 (i.e. caving-gobs of the longwalls 304 and 305) in the colliery M is located right at the boundary between the lease areas of the collieries M and W. Mine air samples for the analyses of SF6 tracer-gas content were taken at the following sampling stations in the colliery W (figure 2): a) On the 665-m level: – Station No. 1 was located behind TI-44 in the road III East in the seam 510D. – Station No. 2 was located behind TI-41 in the road IVa East in the seam 510D. – Station No. 3 was located behind TI-42 in the road V East in the seam 510D. b) In the methane-drainage road I East in the seam 418: – Station No. 4 was located behind TI-424. – Station No. 5a was located in the reconsolidation hole to the seam 510 TM-7. – Station No. 5b was located in the reconsolidation hole to the seam 510 TM-8. – Station No. 6 was located in the methane-drainage pipeline. The air samples were taken at all the sampling stations in accordance with the following frequency pattern: – At 800 o’clock—before injecting the tracer gas, in order to determine the measurement background. – From 830 to 1030 o’clock—every 15 minutes. – From 1100 to 1300 o’clock—every 30 minutes. – From 1300 to 1700 o’clock—every 60 minutes. The analyses of the air samples were performed in the chemical laboratory of the Central Mine Rescue Station in Bytom (Poland) by means of a gas chromatograph with an electron capture detector.


200

Figure 2. Spatial ventilation plan of the seam 510D at the level 660 of the colliery W RESULTS OF THE TRACER-GAS TEST Station No. 1. The tracer gas arrived at the station at 900 o’clock. At that time the concentration of the gas amounted to 0,3 ppb. That concentration changed only very slightly for the next two hours. Then its concentration increased up to the value of 11,9 ppb, which was most probably caused by disturbances in the ventilation network resulting from a stop of the main fan at the southern ventilation shaft. The disturbances occurred between 1040 and 1050 o’clock. At the later sampling times tracer-gas concentrations were near zero. Station No. 2. The tracer gas concentration rose slightly above the background level (0,1 to 0,2 ppb). No distinct tracer gas presence was observed. Station No. 3. The tracer gas concentration rose slightly above the background level (0,1 to 0,2 ppb). No distinct tracer gas presence was observed. Station No. 4. After 1,5 hours from injecting the tracer gas a distinct increase of the gas content (up to 2,6 ppb) took place. In the later part of the observation the concentration amounted to zero or was slightly higher than zero (i.e. from 0,1 to 0, 3 ppb). Station No. 5a. The tracer gas concentration rose slightly above the background level (0,1 to 0,2 ppb). No distinct tracer gas presence was observed. Station No. 5b. A distinct content increase (up to 1,4 ppb) of the gas occurred after 1,5 hours from the injection time. In the later part of the observation the concentration amounted to zero or was slightly higher reaching 0,1 ppb.


201

Station No. 6. A small content of SF6 (i.e. 0,4 ppb) occurred after 3 hours from its injection. In the later part of the observation the content ranged between 0 and 0,1 ppb. Station Station Station No. 1 No. 2 No. 3 behind behind behind TISampling TI-44, TI-41, 42, place road III road IVa road V East East East seam seam seam 510D 510D 510D

Station No. 3 behind TI-424

Station No. 5a recons olidation hole to the seam 510 TM-7

Station No. 5b reconsol idation hole to the seam 510 TM-8

Station No. 6 methanedrainage pipeline

Sampling Concen time tration of SF6 (ppb)

Concent ration of SF6 (ppb)

Concen tration of SF6 (ppb)



8:00

0

8:30

0

8:45

0

9:00

0,3

9:15

0,1

9:30

0,2

9:45

0,2

10:00

0,3

10:15

0,3

10:30

0,0

11:00

11,9

11:30

0,1

12:00

0,0

12:30

0,0

13:00

0,0

14:00

0,1

15:00

0,0

16:00

0,0

17:00

0,0

Concen tration of SF6 (ppb)

Concentr ation of SF6 (ppb)


202

CONCLUSIONS The performed tracer-gas tests proved the existence of fissure connections between the gobs of the seam 418 at the 500-m level in the colliery M and the gobs of the seam 510D as well as the methane-drainage road 1 East in the seam 418 at the level of 665 m in the colliery W. The tests with the use of the SF6 tracer gas confirmed results of an earlier survey of ventilation potential of the portion of the ventilation sub-system of the southern ventilation shaft of the colliery M (the results and their analysis are in possession of the colliery). Identification of the airflow direction by means of the tracer gas clearly showed its migration from the seam 419 in the colliery M towards the seams 418 and 510 D located in the taking of the colliery W.

Risk of Coal Dust Explosion and its Elimination

K.Lebecki, K.Cybulski, Central Mining Institute. Katowice, Poland A.Szulik State Mining Authority. Katowice, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) ©2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: Coal dust explosion is the biggest danger for underground mines. Paper presents its causes in connection with extraction principles, exposing physical and chemical properties of coal, its flammability, amount of dust deposited in the workings. Authors conclude that existing safety rules, if met, are sufficient to ensure the high safety level. Accession of Poland to the European Union brings the necessity to adapt Polish law to meet the EU requirements and introduce such notions as risk assessment and explosive atmospheres. The basic requirements of the EN1127–2 standard, the basic document related to explosion hazards in underground mines, are presented. KEYWORDS: Explosion, coal dust, risk, EN 1127–2 standard

INTRODUCTION The phenomenon of a coal dust explosion has been known since more than 100 years. Almost 100 years ago the Courrieres disaster, costing 1099 lives took place. Other tragedies followed and continue to occur. We would like to be good prophets and tell that the coal dust explosion hazard is over, but experience and statistics tell us that explosions have occurred and will continue to occur. The question appears—is the lesson of many disasters not learned? Do they result because of lack of knowledge and protection means? What is the cause of the biggest mining hazard? Honest answer for these questions is a simple one—knowledge of coal dust explosion causes, how it occurs, and how to avoid it which knowledge—wise is almost complete. But there are troubles with use this knowledge in everyday mining practice. Investigations of separates cases show that


204

breaking of elementary safety rules always caused the explosion. This was proved in the recent catastrophes in Poland (Cybulski 2002) and abroad (McPherson 2001). The general conclusion of such investigation is need for simple and reliable protection systems. In this paper we want to consider what really should be done from the point of view of mining technology to avoid a coal dust explosion. The Courrieres disaster appeared to be a milestone because it showed the dominating role of coal dust in the mining explosion- the importance of the presence of the firedamp, even though as an explosion factor, it is not necessary for explosion to occur. To answer the above questions the simple, general model of an explosion of any flammable substance is very useful. This model relates to all explosion phenomena with combustion, not only in underground mining. Appearance of explosion requires the simultaneous appearance of five factors, forming the so-called explosion pentagon.

Figure 1. Explosion pentagon Lack of one of the elements of the pentagon or braking connection between them excludes the explosion’s appearance. In the case of the underground coal mine this model says that: – combustible (coal dust) exists by its nature and in the explosion the coal dust participates, – oxidizer is atmospheric air, – ignition sources results from mining technology, – underground mine workings are by their nature confined rooms where overpressure can grow, without dissipating as it otherwise would in open space; note that there is not an explosion hazard in open cast mining, – mixing of combustible and oxidizer or the forming of the dust cloud in air occurs in coincidence with ignition. in a normal state the dust cloud is not in explosion concentration range for the mine air; the presence of people in such a cloud of dust is not possible. Those are the main factors of explosion risk in mine working. They enable one to formulate the elementary principles, which upon meeting diminish the explosion risk.

Risk of coal dust explosion and its elimination

205

WHAT IS THE ORIGIN OF DUST EXPLOSION HAZARD? Among the specialists there is not an unanimous opinion to this question: is the coal dust explosion a natural hazard or technical one? Dust is created by the technology of extraction so it is reasonable to treat this hazard as a technical one. But in this case a common understanding of the technical hazard limit is needed by the contact man- this pertains to the machine and a moving object (typical example being falling rock). So, according to Polish regulations dust hazard is considered to be a natural one. As a consequence of such qualification there is the dividing of workings, seams or other part of mine into classes, either categories or other states of hazard. Protection means used in the given working depend on its class. Openly saying, the current Safety Regulations, if strictly met, reduce the risk of explosion. Looking at the problem of explosion risk, according to the Standard PN-N18000 one may say that the risk of explosion in the given working depends on: – the amount of dust accumulated, – size distribution of dust, – depositing of new dust layers on the working surface, – volatile content of coal, – incombustible content of layered dust. Dust accumulated on the floor, sidewalls, pipelines, machines and other devices is not strictly (exclusively) coal dust. It contains limestone dust used for inertization, natural rock dust coming from rock cutting, surface moisture, and other components different than coal. Such a mixture is called “mine dust” or float dust in the US. According to regulations (2002) “mine dust—created during mining and processing of minerals includes substances added to inertize the dust”. Nevertheless, in the workings where no inertizing by adding of limestone dust is pursued “mine dust” is very close in character to “coal dust”. In the National Regulations (2002) there are two distinctive notions used: safe dust and unsafe dust. “The safe dust” contains: – at least 70% of incombustible if deposited in non gassy fields, – at least 80% of incombustible if deposited in gassy fields, and – free water in quantity to not enable dust propagation. Coal dust is created during mining and transported by the ventilation stream. Workings are covered with dust of different grain size. Cybulski’s investigation (1973) proved that layered dust actually contains two important components described as being different in size: – dust “d85” or “fine dust” containing 85% of grain mass passing through a sieve of 75 µm (this dust accumulates at the upper parts of workings), – dust “d25” or “medium dust” containing 25% passing through the 75 µm sieve (accumulates in lower parts of workings). Such inhomogeneous distribution of dust causes the phenomenon of “separate dust clouds”—described by Cybulski (1973).


206

Exact data is lacking on the total amount of dust produced in Polish coal mines. The main sources of dust are: – shearers and roadheaders, – roof support movement, – transportation of minerals particularly at transfer points, and – processing (screening, crushing). It is admitted that 2–3% of the extracted rock is transformed into dust, which gives in Polish coal mining an amount of dust up to 3 millions tons a year. It does not mean that all this dust creates a danger of explosion. Although effectiveness of dust suppression systems (water spraying, dust collectors etc) is still considered to be insufficient these systems diminish the amount of dust deposited in the roadways. Dust deposition is dynamical; it is a continuous process that increases the amount of dust in the workings. It is characterized by intensity of deposition defined as a mass of dust deposited on the given surface at the determined time. To characterize the explosion hazard only flammable coal dust is taken into account. The measurement unit of the intensity of deposition is gram per square meter per day (g/m2/day). Amount of deposited dust at a given location depends on: – intensity of the source expressed in unit of mass of dust created in the unit of time, – size distribution of dust, – air flow velocity in the working, and – flow resistance per unit of working’s length. Typical values of intensity of deposition are in range from almost zero to 150 grams per square meter per day. The upper limit did not changed since introducing high productive longwalls, with the dusting being bound through the use of modern effective dust suppression means. Deposited dust is hazardous if raised into air forming a cloud with sufficient concentration to propagate explosion. The lower explosion limit is determined as 50 grams per cubic meter. Such a big concentration makes a man’s breathing impossible: such a high dust concentration (although being a minimum one) does not normally exist at working places. Being different than flammable gases, entrained dust in typical practice does not reach the upper explosion limit. In mining practice it is very difficult to create a homogenous dust cloud of a concentration more than one kilogram per cubic meter. Although Cybulski (1973) determined the upper explosion limit to be one kilogram per cubic meter, it means only that it was the upper estimable concentration formed in an underground gallery. It would be a tragically misunderstanding to consider “upper explosion limit” in the same way as for gases—in the case of mine gases, beyond the upper explosion limit the flame does not propagate. Amount of dust accumulated in working is determined as nominal concentration—that being measured as the total mass of dust (determined by a special sampling procedure) in given roadway section divided by its volume. The nominal concentrations are in range from grams per cubic meter to kilograms per cubic meter. DUST EXPLOSION HAZARD AND EXTRACTION TECHNOLOGY


207

Ability of coal dust to explode depends on its volatile content. If volatile content is lower than 10% dry ash free coal, such a dust is considered to be not hazardous. After the closing of Lower Silesia Coal Basin there are no longer the coal seams in Poland with such a low volatile content. Volatile content is defined as “volatile products of combustible nature formed from decomposition during heating under standardized conditions”. Volatiles are simply gases released during quick intensive heating of coal. The main components are according to Cybulski (1973)—carbon monoxide (73%) and hydrogen (18%). So, the main parameter needed to be known before the extraction begins is the volatile content in coal. It decides whether an explosion hazard exists. Nowadays all coals extracted in Poland are explosible. The examples of explosibility parameters are given in the table 1.

Table 1. Parameters of explosibility and flammability of coal dusts No.

Mine

Vdaf [%]

Aa W c [%] [%]

d60 [%]

pmax [bar]

KSt,max [m·bar/s]

42,4

1,1

100

7,1

62

590

>3,25

225

21,2

1,2

80,0

7,6

69

610

0,73

176

44,7

3,6

100,0

6,6

54

630

>3,25

240

Tcl T5 min T5 mm [°C] [mJ] [°C]

1

Anna 298/96

2

Chwałowice 307/97

3

Chwałowice 298/98

4

Marcel 299/96

38,4

28,9

2,7

72,2

7,7

86

590

1,9

210

5

Jankowice 298/96

40,4

22,7

1,5

100,0

8,1

140

580

0,12

180

6

Zofiówka

37, 5

33,1

1,5

96,2

7,5

62

560

42

–

7

ChwałowiceDębieńsko

37,7

17,6

2,6

72,7

7,9

79

630

1,1

194

8

Murcki

37,4

9,5

5,5

72,7

7,6

81

610

0,59

158

9

Dębieńsko

26,4

15,1

1,6

63,4

8,4

87

640

–

182

10

Zofiówka

20,7

15,4

1,1

93,0

9,0

162

620

–

188

11

Krupiński

23,0

22,9

1,8

93, 0

8,6

130

620

–

182

12

Pniówek

21,8

18,6

1,2

93,0

9,0

102

620

–

182

13

Jas-Mos Jastrzębie

20,4

19,0

0, 5

93,0

8,5

95

630

–

235

36,5

where Aa—ash content (%), Wc—humidity (%), d60—grain mass below 60 µm (%), pmax— maximum explosion pressure (bar), KSt,max—explosibility index, Tcl—minimum ignition temperature of cloud (°C), Wmin—minimumignition energy (mJ), T5 mm—minimum temperature of ignition dust layer (°C).


208

The KSt value is a synthetic index describing combustion dynamics; it depends on rate of pressure rise, This index is useful in ranking the explosibility of the dusts, According to this ranking all coal dust are weakly explosible. IGNITION SOURCES The problem of ignition is a separate one, Dusts are generally much less flammable than gases, Also, deposited dust, if not settled on the hot surface, is not dangerous until it is raised in the form of a cloud having a respectable concentration, In the modern mining practice there are two main processes (characterized according to ignition sources) that are able to raise the dust and ignite it, These are: – methane explosion called often the primary explosion, – shotfiring. There is the separate group of methane ignition sources, such as electrical equipment, electrostatics etc. The causes of coal dust explosion are clearly seen on the fault trees, generally considered to be one of the important techniques for risk assessment. However, in case of assessing the risk of a coal dust explosion, the methodology of risk assessment, using the fault tree approach, contributes not very much because the main risk factors had already been recognized a long time ago. Yet the fault tree approach is very illustrative and suitable for giving people a broader perspective in their overall training. The fault trees that are presented on the figures 2 and 3 summarize knowledge about coal dust explosions. It is assumed that the main ignition source is primary methane explosion. Logical symbols: as “and” gate and “D” as “or” gate are used.


Figure 2. Fault tree—coal dust explosion

209


210

Figure 3. Fault tree—ignition sources From these drawings the principal role of methane hazard suppression is seen, Diminishing of methane hazard is achievable by use of methane drainage and proper ventilation, However, it should be done very strictly because methane concentration even inside the allowed limits is a combustible in addition to that from coal, Commonly, one says that presence of methane in air diminishes the lower explosion limit of coal dust. APPROACH TO COAL DUST EXPLOSION RISK ACCORDING TO THE EN 1127–2 STANDARD The European Standardization Committee approved the European Standard EN 1127–2 “Explosive atmospheres: Explosion prevention and protection, Part 2: Basic concepts and methodology for mining”. It is a second part of the standard EN 1127–1 “Explosive


211

atmospheres—Explosion prevention and protection: Basic concepts and methodology”. Both standards are harmonized with following Directives 98/37/EC, “Machine Safety” and 94/9/EC ATEX Directive. Both standards have been implemented into Polish Law, European Standard EN 1127–2 introduces notions less used in Polish coal industry. Some definitions are: – explosive atmosphere—mixture with air, under atmospheric conditions, of flammable substances in the form of gases, vapours, mists or dust, in which, after ignition has occurred, combustion spreads to the entire unburned mixture, – hybrid mixture—mixture of flammable substances with air in often in a rather complex physical (multi-phase) state. Examples of hybrid mixture: methane—coal dust—air, petrol drops—petrol vapour—air, – potentially explosive atmosphere—atmosphere which could become explosive due to local and operational conditions, and – protective system—protective system means design units which are intended to halt incipient explosions immediately and/or to limit the effective range of explosion flames and explosion pressures. Protective systems may be integrated into equipment or separately placed on the market for use as autonomous systems. The notion of explosive atmosphere, being basic in European Standardization in the fields of explosion safety was not previously used in the Polish technical dictionary, Below, is presented the most important statement of the standard; the elements are different than those used in National Regulations: Hazardous conditions 1 (explosive atmospheres): underground parts of mine and associated surface installations of such mine endangered by firedamp and/or flammable dust. This includes underworkings where the concentration of firedamp is within the explosion range, e.g. as a result of malfunction (e.g. breakdown of fans), sudden release of large amounts of firedamp (gas outburst) or increased gas emission) due to decrease of air pressure or increase in coal winning). Hazardous conditions 2 (potentially explosive atmosphere): underground part of mines and associated surface installations of such mines likely to be endangered by firedamp and/or flammable dust. This includes underworkings where the concentration of firedamp in the ventilating current or in the firedamp drainage system is outside the explosive range. ATEX Directive specifies different means of protection, which reflect the requirements of the different atmospheric conditions. Criteria determining the classification into categories are the following. Category M 1 comprises equipment designed and, where necessary, fitted with additional special means of protection to be capable of functioning in conformity with the operational parameters established by the manufacturer and ensuring a very high level of protection. Equipment in this category is intended for use in underground parts of mines as well as those parts of surface installations of such mines endangered by firedamp and/or flammable dust. Equipment in this category shall remain functional even in the event of rare equipment faults in an explosive atmosphere and have explosion protection measures so that either:


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– in the event of failure of one means of protection, at least an independent second means provides the requisite level of protection, or – the requisite level of protection is assured in the event of two faults occurring independently of each other. Category M 2 comprises equipment designed to be capable of functioning in conformity with the operational parameters established by the manufacturer and ensuring a high level of protection. Equipment in this category is intended for use in underground parts of mines as well as those parts of surface installations of these mines likely to be endangered by firedamp and/or flammable dust. ELEMENTS OF RISK ASSESSMENT The risk assessment shall always be carried out for each individual situation in accordance with EN 1050 “Safety of machinery—Risk Assessment”. Risk assessment includes the following elements for which the standard gives guidance: a) Hazard identification. The safety data assists in the identification of hazards by demonstrating whether substances are flammable and indicating their ease of ignition. b) Determine whether an explosive atmosphere is likely to occur and the amount involved. c) Determine the presence of and likelihood of ignition sources that are capable of igniting the explosive atmosphere. d) Determine the possible effects of an explosion. e) Evaluate the risk; and f) Consider measures for the minimization of risk. A comprehensive approach shall be taken, especially for complicated equipment, protective systems and components, plants comprising individual units and, above all, for extended plants. This risk assessment shall take into account the ignition and explosion hazard from: – the equipment, protective systems and components themselves, – the interaction between the equipment, protective systems and components and the substances being handled, – the particular industrial processes performed in the equipment, protective systems and components, – interactions of individual processes in different parts of equipment, protective systems and components, and – the surroundings of the equipment, protective systems and components and possible interactions with neighbouring processes. These elements of risk assessment are actually present in the National Regulations, often not directly exposed. It is a decision of State Mining Authority, if admission of EN1127– 2 standard is made, it will radically change the Regulations. It is most certain that the mentality of authors of the standard (there was not a Polish representative on the team) is


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different than that of authors of our Polish regulations. But safety in regard to human life is unique. The future will show that safety regulations in EU countries will be close together but not identical. The Standard leaves to the National Authorities many decisions such as level of de-energizing of equipment. or considering that all mines as being gassy ones independently of the presence of firedamp. The basic difference between current National Regulations approach and that of the EN Standard is that the EN one takes a comprehensive approach to explosion hazard without formal division into methane and dust hazards while the Polish one includes the division. CONCLUSIONS 1. Conditions and technology of coal extraction influence the risk of coal dust explosion in a mine. Extraction technology should be properly chosen to minimize dust generation during cutting and transportation of coal and the probability of the occurrence ignition sources. All these requirements are expressed in the current Polish Safety Regulations. They are modern and in accordance with the current coal extraction technology. 2. The necessity to harmonize of national regulations with the EU requirements will appear in the near future. It will not be bound with the revolution in the safety rules and does not go towards diminishing the safety requirements. But it is expected that the use of such notions as “explosive atmosphere” and “risk assessment” introduced and made permanent in the regulations.

REFERENCES Cybulski K. 2002: Badania i ocena stanu zagrożenia pyłowego po wybuchu pyłu węglowego 6 lutego 2002 roku w KWK “Jas-Mos”. Bezpieczeństwo i Ochrona Środowiska w Górnictwie, nr 3 (103) 2002. Cybulski W. 1973: Wybuchy pyłu węglowego i ich zwalczanie. Wydawnictwo Śląsk, Katowice 1973. McPherson M. 2001: The Westray Mine Explosion. Proceedings of the 7th International Mine Ventilation Congress, Krakow 2001. Ed. RDC Emag. Rozporząjizenie MG 2002: Rozporządzenie Ministra Gospodarki z dnia 28 czerwca 2002 roku w sprawie bezpieczeństwa i higieny pracy, prowadzenia ruchu oraz specjalistycznego zabezpieczenia przeciwpożarowego w podziemnych zakładach górniczych. Dz. U. nr 139, poz. 1169. Rozporząjdzenie MSWiA 2002: Rozporządzenie Ministra Spraw Wewnętrznych i Administracji z dnia 14 czerwca 2002 roku w sprawie planów ruchu zakładów górniczych. Dz. U. nr 94, poz. 841.

Methane Control in Coal Mining Safety System

N.O.Kaledina Moscow State Mining University, Department of Mine Aerology and Safety. Moscow, Russia International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: Methane safety in coal mines may be provided by using complex engineering methods. It is necessary to take into account all relevant factors: technological, ventilation, and dynamics of gas and coal dust. Besides, this problem must be solved taking into account the environmental issues. The aim of this work is to provide help in increasing the safety of underground coal mining with simultaneous decrease of atmosphere methane pollution by means of methane extraction and the use of an optimal ventilation system. A criterion for methane control is offered, which may be used with any control method and also takes into account the development of new technologies of methane utilization.

INTRODUCTION The most widespread methane control methods are ventilation and drainage. The main purpose of these methods usually was getting rid of the gas. But methane has a value as a fuel and raw material for the chemical industry. So most methane should actually be drained with the aim of its utilization with the benefit both to mining safety and environment. With increasing depth the total methane emission from mines and contribution of gobs in their gas balance are growing. Increasing the intensity of ventilation in order to combat methane hazard increases methane emission into the atmosphere, if methane is not utilized. Methane utilization is feasible only if stable outputs of highly concentrated gasair mixture are maintained, which may be achieved only with the use of combined methods of control, including gob area drainage control. Efficiency of gob area degasification depends to a large extent on gob ventilation system: the increase of


216

methane volume in ventilation flow reduces gas volume and its concentration in drainage boreholes. Therefore it is necessary to take into account the impact of ventilation of gob areas when methane control by combined ventilation and degasification methods is applied. 1. PHYSICAL SIMULATION OF AEROGAS-DYNAMIC PROCESSES IN GOB AREAS The problem of airflow through gob is one of the most complex in mine ventilation. Gas emission from gob in a coal mine is an extremely complex process of active gas diffusion in air leaking through porous environment. Dynamic gas processes were investigated with the help of the physical models in the Moscow State Mining University. The processes were examined as stationary and nonstationary, for different types of ventilation layouts, taking into account drainage of methane from gob areas, as well as its absence. For the description of laws governing airflow in gob most expediently is to use the two-partial law of resistance, in which the first (linear) part of the right-hand side of this equation reflects influence of viscosity, the second (square-law)—of inertia: (1) where h—depression, Qga—gob area flows, L—length of filtering way, S—filtering square. Coefficients R’ and R” and their dependence on gob pore characteristics can be calculated with the use of turbulent filtering equations by E.Minsky (1958). In these equations µ/k and ρ/l are analogous to specific aerodynamic resistance r(x) that is equal to the gob area volume resistance of 1 m2 square over 1 m length. Variation of parameters k and 1 with the length of gob area x can be described by empirical relations, containing parameters that do not depend on ventilation regime but on the properties of collapsed strata in gob areas: (2) where b, d—coefficients, depending on the type of surrounding rock; c—coefficient, depending on the working face advance rate. The modeling results were comparable to the actual data, which confirmed correctness of the assumed border conditions. The received representations of laws governing ventilation flow were hereinafter used in mathematical modeling. 2. METHANE DYNAMICS IN COAL MINE GOB AREA Gas emission from gob in a coal mine is an extremely complex process of active gas diffusion in air leaking through porous environment. This process is defined by a set of diverse factors and only approximately can be described by a system of differential

Methane control in coal mining safety system

217

equations. Their solutions by both analytical and numerical methods do not give accuracy acceptable to practice because of complexity of the description of boundary conditions, distribution of intensity of gas emitting sources and structural characteristics of filtering environment. So it is expedient to proceed from complex models including multiple parameters to dynamic models of integrated character, using the most essential interrelations between major factors and parameters of described processes. It simplifies the description of the complex phenomena without loss of necessary accuracy, reduces volume of the analyzed information and facilitates adoption of appropriate control measures. Dynamic gas processes were investigated with the help of the physical models in the Moscow State Mining University. The processes were examined as stationary and nonstationary, for different types of ventilation layouts, taking into account drainage of methane from gob areas, as well as its absence. For the description of laws governing airflow in gob area it is most expedient to use the two-partial law of resistance. In such form the aerodynamic resistance coefficients are dependent on the gob area pore characteristics. Variation of the coefficients is described by empirical relations, containing parameters that do not depend on ventilation regime but on the properties of collapsed strata in gob areas (Puchkov 1993). On the base of the received descriptions of Reynolds number fields for main types of ventilation layouts an average-integrated Reynolds number (Re*) is accepted. It is a complex integrated parameter, taking into account all major factors, determining specific character of gob area ventilation properties: (4) where X, x0, L—geometrical dimensions of gob area, x, y—coordinates. Ventilation conditions, determining methane distribution in system “workings—gob area—gas drainage boreholes”, are characterized by integrated Reynolds criterion, which can be determined from the data gathered by mine atmosphere monitoring on the base of established laws of leakage-flow through a waste zone. It is established, that formation of a methane concentration zone is determined, mainly, by permeability, spatial structure of leakage-flow (i.e. ventilation layout) and intensity of gas-emitting source. For any type of ventilation layout a zone of high gas concentration is formed, which becomes a main source of gas emission into the workings. The high concentration zone (HCZ) localization depends on distribution of emitting sources over the length of gob area determined by character of rock-mass displacement and yield due to coal seam extraction, ventilation system and layout. Changes of maximum methane concentration with the integrated Reynolds number (Re*) is described by exponential functions. Thus ventilation system and layout are the main factors, influencing the efficiency of gas extraction from gob area. The relation between methane inflow and flow of ventilation through the gob area can be described by the following formula: (5) where A—coefficient, depending on ventilation layout and spatial sizes of gob.


218

Researches have shown, that for extractive unit there is some range of ventilation conditions that are most rational from the point of view of combined ventilation and gas extraction, ensuring methane safety in the workings and high efficiency of drainage. There will be an optimum; least intensive ventilation system and layout, providing the permissible methane concentration in return air (Puchkov, Kaledina 1995): 0,01Qu(Re*)=If+Iga−Id(Re*) (6) where If—volume of gas inflow from working faces, m3/min; Iga—volume of gas inflow from gob areas, m3/min; Id—volume of methane, extracted by drainage, m3/min; Qu— volume of intake airflow, m3/min. The values If and Iga change in accordance with advance rate and even during a shift with characteristic cycles. Because of this they have to be calculated from the data gathered by mine atmosphere monitoring with the use of established laws. So the algorithms of mine gas extraction systems, based on a principle of maximum total methane inflow by means of optimum distribution of vacuum along boreholes and pipelines, have to be added by a method of aerodynamic influence on gas emission from gob. 3. THE CRITERION OF AERODYNAMIC METHANE CONTROL The strategy of methane control by ventilation should take into account not only gas hazard, because the mine ventilation conditions are, as a whole, defined by a set of diverse factors, reflecting individual characteristics and influence of each piece of underground equipment, as well as technology and organization of the mining process. The optimum mine ventilation system and layout have to be chosen, as a whole, in view of all the adverse factors. The functioning of coal mine’s ventilation systems can be presented by a multi-level hierarchical model of air and gas flow aggregate. The model’s levels conform to the technological hierarchy levels. For every level the criterion of optimal gas control is to supply maximum output of suitable for utilisation (with high gas concentration) methane by drainage. Apart from that, the optimal solution should not only provide the necessity of mine methane usage, but also opportunity of development of new technologies of its utilisation. Such criterion can be presented as follows: (7) where Id, Iv—methane flow from drainage and ventilation system. The coal mining experience and numerous scientific researches prove the necessity to coordinate control of gas emission with the use of both air- and gas-dynamic methods. Besides, ventilation and gas extraction are necessary to be considered as a uniform system.


219

4. RESEARCHES OF METHANE FLOW FROM THE GOB AREAS IN COAL MINES The total volume of methane being drawn in a mine comes from several sources: from working faces, from extraction units and from gob areas. Volume and intensity of methane inflow from the working faces can vary in time rather considerably, because they depend on volume of production, longwall face advance rate and ventilation layout. Methane inflow from the gob areas into the workings is not subject to sharp fluctuations, because the processes of active collapse and compacting of rock-mass here are practically completed. As an unpredictable and significant methane source the gob areas most actively show themselves in gas balance of mines during sharp fluctuations of atmospheric pressure, when methane flow from them into the workings increases considerably. This results in exceeding the allowable methane concentration in the working places, with all following from here consequences (switching-off of the electric power, cessation of work, leaving of affected workings by the people). The research which have been carried out by the authors, and also work of other researchers show, that gob areas can contribute about 40%, and can reach 70% of the total volume of methane in the gas balance of mines. In table 1 the average data on gob area contribution into the total gas balance of mines of joint-stock company “Vorkutaugol” from 1992 to 1996 are given. In table 2 the average data on the basic methane sources of some mines of this company from 1992 to 1997 are given (Kaledina, Mescheryakov, Semenov 2000). In table 3 the data of gas balance of the mine “Raspadskaya” (Kuzbass) are given. Here Io—is the methane volume in fresh airflow. Methane flow from the gob area is represented by two components—the first one—in a ventilating current and the second one—in a drainage system. The range of change of values is specified in numerator (at minimal length of the gob area—200 m), average values for the considered period of improvement are given in a denominator.

Table 1. Methane flow of mines “Vorkutaugor

Name

Methane Degasification flow efficiency total total m3/min. % (m3/min.)

Gob areas methane flow m3/min.

Gob area issue in the total gas flow %

Degasification efficiency of gob areas % (m3/min.)

Centralnaya

122

32 (39)

61

50

7 (3, 5)

Komsomolskaya

163

37 (60)

60

37

0(0)

Severnaya

178

36 (64)

63

35

2 (0, 7)

Zapolyarnaya

142

34 (48)

19

13

0(0)

Yuznaya

98

40 (39)

49

50

0(0)

Ayach-Yaga

82

31 (25)

32

39

5 (2)


220

Table 2. Main coursers of methane flows into ventilation system Methane flow total Mine

Methane flow from gob areas

Methane flow from mining workings

m3/min.

%

m3/min.

%

m3/min.

%

Centralnaya

83

100

46, 5

56

36, 5

44

Komsomolskaya

103

100

60

58

43

42

Severnaya

114

100

62

54

52

46

Zapolyarnaya

94

100

19

20

75

80

Yuznaya

59

100

49

83

10

17

Ayach-Yaga

57

100

30

53

27

47

Table 3. Gas balance of mine “Raspadskaya” (%) Mining unit

Gas balance issues Io (inflow)

Icb (coal bed)

Ig-act (active gob area)

Ig-age (old gob area)

5a-6–16

0,0–10,8 2,2

41,4–6,8 27, 8

56, 0–88, 9 66, 7

5, 9–0, 1 3,4

5a-10–14

0,0–14,4 4,1

38, 3–8, 6 20,2

55, 8–91, 5 72, 5

19,4–0,8 3, 2

An opportunity of a rough estimation of volume of residual gas flow and the time it will be kept has shown itself in the case of “Yuznaya” mine, which was recognized as unpromising and therefore from 1996 the mining was completely stopped. Methane flow to this time was 89, 5 m3/min, including gas from the old gob areas of about 42 m3/min. The workers of the mine’s ventilation service carried out measurements of air-volume and methane concentration during all the time, right down to when the main fan was stopped. Within one month after the mining operations were discontinued the methane flow has decreased to 55 m3/min. It is obvious, that in this period the process of active methane allocation, caused by collapse and compacting of roof in was over. For the next five months methane inflow has decreased from 55 to 45 m3/min, i.e. has come nearer to the average gas inflow from old gob areas in the gas balance of the mine before mining was stopped. Today the methane extraction is not made from any closed mine in Russia. Only in some Feasibility Reports measurements of methane concentration at the most probable points of its exit to surface is stipulated as a safety measure. The advantage of this is doubtful, since in the Feasibility Report the measures allowing the control of gas flow are not stipulated. Methane control of a closed mine is understood as a system of organizational and technical measures allowing influencing methane outflow in order to exclude possibility of accidents.


221

CONCLUSIONS 1. The choice of an optimal mine ventilation system is considerably influenced by dynamic interactive processes taking place in gob area between air, gas and extraction units. 2. Ventilation conditions, determining methane distribution in system “workings—gob area—degassing holes”, are characterized by integrated Reynolds’s criterion, which can be determined from the data gathered by mine atmosphere monitoring on the base of established laws of leakage-flow through a waste zone. 3. Usage of established laws governing gob ventilation flow and its influence on area drainage parameters helps to increase gas extraction efficiency and to provide control of gas emission into the mine ventilation system. 4. For effective methane extraction from coal mine’s gob areas it is necessary to isolate them after mining is complete to get no aerodynamic interaction with surface or mine workings. 5. The functioning of coal mine’s ventilation systems can be presented by a multi-level hierarchical model conforming to the technological hierarchy of mine. For each level the criterion of optimal gas control has to provide maximum output of methane-air mixture, suitable for utilisation. 6. The criterion, offered for methane control, is universal, because it may be used with any control method and also takes into account the development of new technologies of methane utilisation. 7. Obtained results may be used to design mine ventilation and gas extraction systems. 8. The residual methane flow from the gob area of a mine, which stopped operation, is close to the average volume of methane outflow from old gob area of this mine during the last years of operation. Exact estimation of the duration of methane outflow requires careful monitoring of changes in methane volume during preparation of the mine to closure. 9. In order to adopt practical and efficient way of methane extraction from mines discontinuing operation, it is necessary to take into account geological and technological conditions and the tasks, which have to be undertaken. Any combination of measures allowing achieving steady and controlled flow of air-methane mixture to the required draw point is acceptable.

REFERENCES Kaledina N.O., Mescheryakov D.A., Semenov A.S. 2000: Determination of Methane-Flow from Old Gob Areas. Mining Informatics and Analytical Bulletin, № 7–2000, p. 77–79. Moscow State Mining University. Moscow, Russia. Puchkov L.A. & Kaledina N.O. 1995: Methane Dynamics in Coal Mine Gob Area. Moscow State Mining University. Moscow, Russia. Puchkov L.A. 1993: Underground Mining Gob Area Aerodynamics. Moscow State Mining University. Moscow, Russia.

Rock Bursts—Preventive Measures Undertaken in the Polish Mines

Józef Dubiński, Władysław Konopko Central Mining Institute. Katowice, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 1. INTRODUCTION In the Polish underground (deep) mining bump hazard occurs in about 70 per cent of collieries and in all copper mines. The term of bump hazard means a possibility of a bump occurrence as a result of unfavourable geologic and mining-technological conditions in an underground working or its vicinity. A bump itself is a dynamic phenomenon causes by a rock-mass tremor resulting in a violent damage or destruction of a working or a part of it, which caused a complete or partial loss of its functionality or safety of using it (Kidybiński 2003). Thus a bump is a particular kind of rock-mass tremor. In hard coal mining and copper ore mining there are rock-mass tremors of seismic energy of up to 1010 J (up to 4, 5 magnitude of Richter scale). The lowest energy values at which bumps were observed in collieries amounted to 5·103 J, in case of copper ore mines they were one order of energy values higher and amounted to 5·104 J. The higher the seismic energy of a rock-mass tremor the higher is the probability of bump occurrence. While at a rock-mass tremor with the seismic energy of the order of 104 J that probability amounts to circa 10−4, at the tremor energy of the order of 1010 J it approaches the value of 1,0. Rock-mass tremors, in particular the high-energy ones, pose a specific nuisance to the inhabitants of mining regions both in the sphere of damage caused to surface objects and by arousing some psychical discomfort. In order to restrict the number and effects of rock-mass tremors and rock bumps in the Polish mining a wide research on the causes and circumstances of their occurrence is done, and judicial, technical and organisational measures taken in this sphere bring about expected results. While in 1960–1965 above 50 bumps per year were noted in the collieries, currently—at a comparable production volume—the number of bumps occurrences decreased to several ones yearly. Preventive measures applied in the Polish collieries aimed at restricting the number and consequences of those hazardous


224

phenomena are presented below—as briefly as the requirements towards the article demanded. 2. THE PHENOMENON OF A BUMP A rock-mass tremor manifests itself at a mine working by ground motions and strong acoustic effects. High-energy tremors, in particular the ones the foci of which are situated close to a working, can cause measurable effects in the working. For practical reasons we divide the effects into stress reliefs and bumps. Stress reliefs—apart from the symptoms characteristic of a rock-mass tremor—also cause cracks and fissures in rocks surrounding the working (rock-mass fracturing) and as a result only a slight damage, causing no worsening of the working’s functionality and safety of operations performed therein.

Figure 1. Symptoms of dynamic phenomena in a rock-mass and causative relationships between them In opposition to that, a bump causes a series of negative effects up to total destruction of a mine working or some part of it. They are presented schematically in figure 1.

Rock bursts—Preventive measures undertaken in the Polish mines

225

3. CAUSES OF THE BUMP HAZARD IN THE POLISH UNDERGROUND MINES A bump hazard results from many factors of geological, mining-technological and technologic-organizational nature (Dubiński, Konopko 2000). They may be presented systematically as it follows in figure 2. The Upper Silesian hard coal deposit has originated in the Upper Carboniferous period. It is a multi-seam deposit. The thickness of the seams is diverse, the most coalbearing seam reaches 20–24 m in the areas of its highest thickness. The total thickness of all seams in some areas of the coal basin reaches 60–70 m. The formation incorporates power coal, coking coal and orthocoking coal. While the coking and orthocoking coals are characterised by low strength (up to 15 MPa), the uniaxial compression strength of the power coals ranges between 20–35 MPa. The hard coal seams occur at various distances apart. Sometimes they form groups of 3 to 4 seams separated by waste-rock strata of several-metre thickness and mainly of low strength. There are also seams considerably distant from each other. In those cases 40– 80-metre thick and 60–100 MPa strong monolithic sandstone strata or mudstone strata of similar strength separate the coal seams.

Figure 2. Causes of bump occurrence The deposit is criss-crossed by numerous faults mainly of the parallel-of-latitude and meridional trends. The largest faults have throws ranging between 150–300 m. The coal deposit, in the areas of its larger resources, is fully developed in terms of mining and the day surface above is extensively urbanised. The depth of mining reaches up to 1200 m, only longwall mining systems are in use, mostly with caving, not so often with hydraulic filling. Dog headings or passageways are supported with V-section steel arches whereas longwall faces with powered shield supports. In the backfilling increment area the roof is temporarily secured with timbering (maximum exposure of unsupported roof—10 m). The collieries are old, i.e. aged between 50 and 150 years.


226

The aforementioned geological and mining-technical conditions determine the bump hazard in the existing collieries. In a multi-seam deposit extracted for many years, in the gobs there are numerous unmined fragments of coal seams left over which acting as abutments concentrate stresses and transfer them to the neighbouring seams. As the stresses transferred by many unmined seam fragments and edges existing in many seams superimpose and are summated, it comes to a situation that in a currently extracted seam being under such influence, the local state of stress considerably exceeds the gravity stresses and the gradient of stress increase is significant. The undermined thick strata of high strength become deformed and break while generating high-energy tremors in the rock-mass (such strata are called seismogenic). In the vicinity of faults there occur residual stresses diversely oriented. In many formerly extracted seams various fragments of coal body or remnants have been often left over close to faults. That additionally increases the already existing high level of stress and bump hazard accompanying mining operations in areas not far away from faults. Considerable thickness of seams and extensive development of the day surface above impose application of mining methods not always the best from the point of view of the bump hazard. All in all it is legitimate to state that the bump hazard in the Upper Silesian collieries results from the (Konopko 1994a): – considerable depth of mining and the relating state of high stress, – concentration of stresses resulting from the occurrence of remnants and edges of neighbouring seams, – significant faulting of the deposit, – occurrence of monolithic and thick rock strata of high strength in the roof of coal seams, – considerable thickness of seams of high strength and elasticity, – great mining depth, approaching 1200 m, – significant depletion of deposit reserves, imposing extraction of residual seam parts, – requirement to protect the extensively urbanised surface that not always permits using of mining methods most suitable from the point of view of bump hazard. In these conditions bumps may occur as a result of a state of high stress in a seam (the socalled seam bumps) as well as the more difficult to control ones, connected with breaking of the undermined seismogenic strata located in the roof of a coal seam which cause a dynamic, impact loading of the coal seam (the so-called roof bumps). Of course at some very high stress state in a seam even a relatively slight impact of roof strata may cause a bump. Predominantly, workings driven in seams are most exposed to the destructive influence of bumps. Only sporadically bumps occur in headings driven in waste rocks at a small distance from a seam. The Polish copper ore deposit has the form of a seam the thickness of which locally reaches up to 20 metres. It is located in the border area between the Rotliegendes sandstones and Jurassic dolomites. The roof part of the Rotliegendes sandstones is most often the ore-bearing rock. Clayey shale of several tens of centimetres in thickness rests on it and the floor part of Jurassic dolomites lies above them. The clayey shale may not occur, but only the sandstone formations or sometimes the dolomite are ore bearing.


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The characteristics in this formation are the proportions between the strength parameters of the floor rocks, deposit, and the roof rocks (Butra 2002). The roof part of the Redliegendes sandstones with a clay binder is characterised by slight uniaxial compression strength, i.e. 20–40 MPa. Sporadically sandstone with an anhydrite binder occurs and then the strength may increase even up to 50–60 MPa. The average strength of the deposit’s rock being under extraction ranges between 40–100 MPa, whereas the strength of the roof dolomite between 50–110 MPa. Above the dolomites there is a very thick anhydrite stratum the thickness of which reaches even 150–250 m, and uniaxial compression strength of its rock—150–250 MPa. In some parts of the formation, not far away from the deposit a rock-salt seam occurs that can locally reach the thickness of 200 m. The copper ore deposit is criss-crossed with numerous faults, many of them having throws between 50–150 m. The latter are seismically active, and conducting mining operations in their vicinity is dangerous because of the bump hazard existing there. Currently the copper ore deposit is mined at a depth approaching 1250 m by means of room-and-pillar methods with roof bending, caving, stowing (i.e. stowing waste rock in gobs) as well as hydraulic filling. The roof and sides of workings are secured almost exclusively by means of roof bolting with expanding bolts, grouted bolts and in case of large heading crossings with up to 6,0 metre long rope bolts. The bump hazard in the copper ore mines results first and foremost from breaking of the anhydrite stratum and the impact load of the deposit following from that or it is connected with the seismically active faults having large throws. Some increased bump hazard also occurs in the areas where sandstone deposits with an anhydrite binder lie in the floor or during multi-heading (mostly 3) opening-out and development operations conducted close to and/or in a deposit. Similarly to what happens in the collieries, roof (impact) bumps decisively predominate. It is not possible to exclude seam (in deposit) bumps, in particular when a sandstone stratum with an anhydrite binder occurs in the deposit’s floor. 4. METHODS OF BUMP PREVENTION Appropriately performed bump prevention should assume: – assessment of the bump hazard state, – determination of causes of the hazard, – selection of appropriate methods to limit the hazard level, – effectiveness inspection of applied methods for reducing the hazard. Because of the multi-cause origin of bump hazard any bump prevention should also be accomplished multi-directionally. In particular, that concerns the assessment of the bump hazard state and determination of its sources. Correctness of findings in this sphere has a fundamental and decisive impact on appropriate selection of bump prevention methods and their effectiveness in given geological and mining-technical conditions. The so-called “past mining events” should be analysed with particular care. Under this term we understand a detailed identification, in time and space, of all areas and methods of mining done in the analysed rock-mass part, in particular, localisation of pillars, edges, faults that


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may have negative influence on specific areas of the seam with planned mining operations. Knowledge of the “past mining events” and geological profile of the rockmass as well as strength-and-deformation properties of specific rock strata enables finding these areas of the seam in which stress anomalies occur. Analyses of these data may be performed by means of appropriate, commonly available (commercially distributed) or developed by oneself computer programmes, proved in conditions of the Polish collieries or Polish copper ore mines. This is a right thing to confirm the analytical findings by geophysical methods. The methods often allow us to discover rock-mass and/or seam areas with stress anomalies, which for diverse reasons could not have been identified analytically.

Figure 3. Bump control methods Bump control methods (reducing bump hazard state) are based on removing elasticity features from rock-masses, in particular, from seismogenic strata, seams, and rocks in immediate vicinity of mine workings. The aforementioned may be accomplished by means of long-term methods through appropriate sequence of seam extraction or by active means causing destruction of structure of seismogenic strata, and seam and rocks surrounding the mine workings. It is important to remember that fully effective bump control measures may exclusively be those that affect the hazard’s source, i.e. the seismogenic strata in case of roof bump hazard or a seam (deposit) in case of seam bump hazard. Of course these are extreme types of bumps. Most often bumps occur in conditions characterised by high stress state in a seam (deposit) associated with the occurrence of the seismogenic strata. So bump control methods and places of their application should be accordingly selected. A set of bump prevention methods applied in the Polish mining industry is presented in figure 3. 5. IDENTIFICATION AND MONITORING OF BUMP HAZARD Premonitory recognition and assessment of bump hazard state are performed by means of the aforementioned analytical methods and by mining reconnaissance. The latter one is


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an expert method consisting of a detailed analysis of geological and mining conditions in a given deposit area characterised by 20 different factors to which are assigned specific weights (points). Their sum makes up an assessment of bump hazard state. Depending on that sum a working under consideration may be classified into one of the four following bump hazard states (Principles and Range of Applying the Integrated Method for Assessing the Bump Hazard State in Collieries 1996): a—no hazard, b—slight hazard, c—medium hazard, d—high hazard. Depending on a hazard state—there are specific requirements with regard to planning and performing mining operations. Of course hazard state is different in various parts of a hazardous deposit area depending on geologic and mining conditions including the socalled “past mining events”. In the course of mining operations a bump hazard state is assessed by means of the integrated approach that encompasses—apart from the method of mining reconnaissance—a combination of the following methods (Dubiński 2000), (Dubiński, Konopko 2000), (Dubiński, Mutke 2002): – method of probing drilling, – seismologic method, – seismoacoustic method. The commonly known method of probing drilling relies on boring holes in a seam by means of a standard drill bit of 42 mm in diameter and on assessment of a hazard state based on the output of drillings from particular boreholes and the bore hole depth at which the bore dust output increases. Using this method, mining reconnaissance method and the other methods of the aforementioned integrated method enables determination which of the four bump hazard states mentioned above (i.e. “a”, “b”, “c” or “d”) exists in a working under consideration. The aforementioned seismologic method uses a set of sensors installed in a rock-mass and its configuration depending on the form of the area examined (installation of seismometers) in order to localise tremor foci and to assess seismic energy of tremors. Lately foci of tremors have been localised although the accuracy of depth determination is still doubtful. Based on the model of a tremor focus it is possible to: – determine the number and seismic energy of tremors as well as dynamics of their changes testifying about changes in a bump hazard, – identify the system of causative forces responsible for the occurrence of a given tremor, – determine the parameters describing the spatial position of a focal plane (in case of occurrence of the shear mechanism and the forces causing that process), – determine the seismologic model of a focus. Information obtained allows assessing anomalous states of rock-mass instability in a wide area and determining changes in bump hazard state of mine working. The seismoacoustic method allows investigating acoustic emission of higher frequency than the seismologic one, i.e. of frequency not lesser than 103 Hz. This kind of


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survey is performed in deposits at sites located close to headings and working faces. Deviations of acoustic emission and gradient of their changes in time allow assessment of the bump hazard state in the given part of a working, because due to attenuation of highfrequency waves observation distance is limited. Combined interpretation of results obtained from the mentioned methods gives a chance of making a reliable assessment of a bump hazard state, because: – the method of mining reconnaissance and the seismologic method enable an assessment of rock-mass state in a greater area, – the method of probing drilling and the seismoacoustic method enable an assessment of readiness of a seam to bump at a specific place. In the copper ore mining the method of probing drilling is not in use. Also trials of core drilling and determining “core disking” have not given premises that would justify its application. In both the groups of mines—i.e. collieries and copper ore mines—a lot of information on bump hazard may be obtained through simplest observations that should not escape the notice of mines’ supervisory staff and even miners. These are: – “rock spalling”, which means dynamic spalling of coal (ore) faces and/or side walls, – cracks in rock-mass, – high sonority of rock-mass (in the copper ore mines), – intensive closure of workings, damaging supports, – change in grain size composition of broken rock, – “overblasts” (breaking down a larger volume of body of rock (coal) than normally at a given quantity of fired explosive charge. In justified cases, in particular while affirming build-up of bump hazard, also other geophysical methods are in use: – the seismic method in its profiling and tomographic variants, enabling determination of places with anomalous velocities of the seismic wave propagation and the resulting anomalous stress fields (Dubiński 1989), – the gravimetric method based on measuring changes in the terrestrial gravity force, enabling assessment of volumetric strains of rock-mass in the vicinity of a working and prediction of seismic behaviour of a rock-mass, – the geoelectric method based on investigation of electric properties of rocks in relation to changes of stress conditions in a rock-mass, – the method of induced seismoacoustic activity (ISA) based onn inducing vibrations in a medium through firing a small explosive charge and measuring seismoacoustic activity at precisely defined distances form the blast hole. Its value and gradient of decay in time is a measure of bump hazard. The method of passive tomography has intensively been developed lately. It uses records of natural rock-mass tremors, in particular the ones occurring in areas being under examination and its vicinity (Dubiński, Lurka, Mutke 2000). In the process of calculations an image of velocity field of seismic waves is obtained. In order to find a minimum of the functional the evolutionary algorithms are used due to their considerable stability. In the applied methodology of calculation tremor foci additionally undergo


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delocalisation. Moreover an option of taking into account varying dimensions and geometry of the network is introduced in the velocity model of a medium. Current experiences gained during applying the passive tomography show that foci of strong rock-mass tremors are mostly connected with zones of anomalous great velocities of P-wave propagation in rock medium and slopes of the zones on which high gradients of changes in the velocity field occur. In other words the strongest tremors of mining origin should be situated in zones with stress concentrations in which strong and rigid strata occur which are capable of building up a great strain energy (larger velocities of propagation of the longitudinal P-wave). Zones with low values of velocity field correlate in principle with zones of low seismicity. They may also arise as a result of strong bumps and tremors having occurred in the past that caused local destruction of structure of rocks. In order to obtain good results of distribution of velocity field by means of the method of passive tomography it is indispensable to assure an appropriately dense and regular coverage of the area under examination with seismic ray paths. Therefore calculating images of the velocity field is performed for a whole group of tremors simultaneously (tremors selected from some time window). Because of the changeability in time of the strain-and-stress field also the velocity field is subject to changes indicating in this way areas endangered at a given time and stage of extraction work. Therefore images of the velocity field should be drawn up cyclically so as to enable tracing changes in localisation of endangered areas. Results of calculations of the method of passive tomography may be useful in determining and tracing changes in more extensive zones of seismic hazard. Thus, such information enables better planning of development and extraction operations as well as indicates areas in which it should be supplemented (e.g. with the seismic method) or the methods of active bump control should be applied. The passive tomography is currently the only one method enabling non-invasive, in situ, comprehensive tracing of rock-mass conditions in areas distant from a seam (inaccessible to other methods). A map of the velocity field calculated with the method of passive tomography for the area of the longwall face 772 in “Bielszowice” Colliery is presented in figure 4. The image refers to the state at the end of 2001. Foci of tremors are located in the zone with highest values of velocities of seismic wave and slopes constituting high gradients. Because, as extraction is advanced, stress states change (the zones of stress concentration move), images of the velocity field should be calculated cyclically. The frequency of drawing up tomographic images of the velocity field depends on seismic activity of an area under examination and on development of the mining situation in the extraction area under examination.


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Figure 4. A map of the seismic wave velocity field calculated using the method of passive tomography for the area of the longwall face 772 in “Bielszowice” Colliery The method of passive tomography also allows calculating directional velocities applied to localise foci of tremors. Because of the complex calculation procedure and need of using high performance computer equipment the method of passive tomography should be treated as a supplementary method (used in case of need). The comprehensive method of assessing bump hazard state is used obligatory in all mines operating in hazardous conditions and the mines have special departments for bump issues and are equipped with appropriate measuring instruments. Other methods of assessing bump hazard state are applied if necessary based on ones own measuring means or on specialised firms (institutions). 6. BASIC METHODS AND TECHNOLOGIES OF BUMP CONTROL 6.1. Planning mining operations in bump hazard states An appropriate long-term mining plan worked out with the use of findings of the analytical method and the method of mining reconnaissance is the most inexpensive and simultaneously the most effective method of controlling bump hazard. Of course, not in all conditions it may be a sufficient method. Appropriate planning does not exclude a need of applying other methods of reducing the hazard. Some canons of planning mining operations in bump hazard states may be set out. These are as follows (Konopko 1994): – Coal seams should be opened out by means of the in-stone development skeleton, i.e. with opening-out headings being driven in waste rocks; such a solution, although


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expensive, in a longer period of time may be useful also from the economic point of view. – It is necessary to avoid driving passageways ahead of an extraction front and number of passageways in a seam should be limited to an indispensable minimum. Troubles connected with advancing an extraction front across each one heading existing in a seam or a bundle of development workings in copper ore mines are very persuasive in that matter. – In a multi-seam deposit in particular while working a group of seams, as the first should be extracted the so-called distressing seam. It should meet the following conditions: – a distressing seam should be the least bump-hazardous seam in the group to be extracted, – if all seams of a group are equally hazardous—then the thickness and regular lying of the distressing seam should warrant its clear extraction, without leaving over any remnants, – the thickness and distance of a distressing seam from other seams of a group should assure an as effective as possible stress relief in at least one seam of the group of seams, – the other seams in the group when distressed should be extracted in an area laid out by the contour of gobs of the distressing seam but reduced by the range of influence of its edges/abutments (the range may be examined by means of, for example, the method of seismic profiling). – A thick seam may be effectively distressed through extracting its one slice. – Effectiveness of distressing in terms of time and space is greater in case of undermining than in case of overmining of the distressing seam; similarly effectiveness of distressing is greater when the distressing seam is extracted with caving than in case of filling (in particular hydraulic filling). – While mining seams more distant from each other—the gained distressing results have practically non-measurable values. In such conditions, depending on the place of occurrence of a seismogenic stratum it may appear necessary to depart from mining seams in the top-to-bottom sequence, and to undermine the seismogenic stratum by a seam lying far beneath the stratum. Mining next seams in the bottom-to-top sequence may efficiently disintegrate the seismogenic stratum and hence deprive it its elastic qualities, in-rock-mass strength, and through that its ability of generate high-energy rock-mass tremors. – It is important to avoid situation in which an extraction front approaches fault(s) or gob. An extraction front should move along gobs or faults or move away from them. – The roadways at gobs should be driven in the direct vicinity of the gobs or not far from them. – Faults influence changes of stress state in a rock-mass individually. Hence localisation of roadways in relation to faults should be determined based on results of examinations of influence ranges of faults. – In a deposit with a differentiated stress state roadways should be situated outside stress concentration zones, and in case there is a need of drivage across such a zone it is necessary to plan specially reinforced roadway supports and appropriate current (active) preventive measures.


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– The first longwall face in a panel (the so-called opening longwall face) should be situated in a zone of the most unfavourable stress state. – The next longwall face in a panel, which causes that total gob width reaches 40 to 60 per cent of the mining depth as a rule, induces the most intensive seismic activity and thereby it is most bumps hazardous. Therefore a mining plan should be worked out in such a way that the longwall face producing that width of a gob is located in the most favourable part of its panel, from the point of view of geologic and miningtechnological conditions. – No relationship has been found between the width of an extraction front and a bump hazard. That does not concern the last longwall face in a panel when the face is situated between gobs or between a gob and a fault. The face’s width should be defined individually depending on the existing geologic and mining-technological conditions. – Bump hazard in given geologic and mining-technological conditions increases with the height of a mine opening. It is recommended to avoid mining too thick slices in very thick seams. The most appropriate thickness of a distressing seam or slice amounts to 2, 0–3, 0 m. – In case of seams lying close to each other it is necessary to avoid crossing workings driven in different seams (in fact it concerns crossings of their projections onto a horizontal plane, the so-called apparent crossings). It is also necessary to avoid driving parallel roadways in different seams too slightly shifted horizontally in relation to each other (the so-called apparent pillars). It is also purposeful to keep one direction of extracting a deposit (i.e. one direction of movement of extraction fronts). – In the vicinity of faults it is essential to extract the deposit first in the downthrown block and next in the up thrown block. – In a bump-hazardous deposit supports of workings should be appropriately reinforced. – In case of various co-existing hazards, in particular bump, methane and fire hazard— electrical devices and cables should meet given safety requirements. – It is necessary to use experiences from operations in a given mining section or panel or similar geologic and mining-technological conditions. Causes of some bump occurrences are not to be explained even post factum. In the neighbourhood of such zones it necessary to take special precautions and take possibly comprehensive preventive measures. 6.2. Technologies of the active bump control Bump hazard occurs as a result of mining operations done in rocks of high strength, capable of building up elastic energy and violently releasing it in some strain and stress states. In rocks of low strength and high plasticity parameters it is difficult to maintain mine workings in good repair because of their pseudo-continuous closure. But there are no violent rock displacements, characteristic of bumps. Hence the active bump control is aimed at changing the properties of deposit and/or surrounding rocks, depriving them of the bursting capability or at least diminishing their strength and elasticity. Mostly active bump control activities are the blasting operations. The objective here is to produce a zone of fractures around mine workings or to destroy deposit rock structure, in particular of the seismogenic strata. The energy of vibrations caused by rock-mass


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tremors can be attenuated in that zone and impact load is not transferred onto supports of mine workings or at least considerably reduced. In practice, concussion blasting or concussion-and-winning blasting is performed. The former is aimed at inducing rock-mass tremors and causing rock fracturing. In general, boreholes for concussion blasting are drilled in a deposit and have a depth 3 to 6 times longer than the standard depth of webin given conditions. Mostly the explosive charge takes 0,5 to 0,7 of the depth of a blast hole, so it is localised outside the web zone. Firing the charge of explosive is not expected to win any part the body of rock (coal) but only to produce fractures therein. After a heading or working face had advanced through a distance not longer than a half of the length of the blast holes the operation of concussion blasting was repeated. In the copper ore mining where winning operations are performed exclusively by means of blasting—blasting operations simultaneously serve as ore winning blasting and concussion blasting. It is done by concurrent blasting a number of faces; sometimes in case of long extraction fronts, up to 30–40 faces are blasted at the same time. When the fronts are shorter the number of blasted faces amounts to n-2, where “n” is a total number of faces. Such blasting in specific hazard states induces about 50–80% of all high-energy tremors. Sometimes stress reliefs and even bumps are trigged off. In the copper ore mining induced bumps in general do not cause essential damage to property as the mobile equipment is withdrawn from development drivages, where bumps almost exclusively occur, to a safe place for the time of blasting. Such a blasting employs as large explosive charges as possible (up to 2000–3000 kg). It is often that single, large-diameter boreholes, 2 to 3 times longer than the ones for winning purposes, are drilled into the room faces and then fully filled with an explosive charge in order to augment the inducing effect of blasting. In collieries, in case of occurrence of a seismogenic stratum at not a large distance from the roof of a seam to be extracted—explosive charges are fired in such a stratum for the purpose of weakening it, and facilitating its transition to technological caving and to some inducing effect, in particular, to stress relief occurrence. This is because induced bumps cause considerable property losses by damaging expensive supports and machinery. The arrangement, dimensions and number of blast holes are adjusted to local conditions. In collieries, an explosive charge fired in one-shot amounts to e.g. 70–120 kg in case of concussion blasting in a 250-metre wide longwall face, 100–150 kg in case of disintegration or torpedo blasting in the roof strata, 20–30 kg in drivages. In collieries in zones with a very high level of bump hazard coal winning by means of machines can sometimes be substituted by blasting. Such cases more often take place in drivages than in longwall faces. Then shearer-loaders are used only as longwall face loaders. Numerous stress reliefs occurring after winning blasting confirm the legitimacy of such a solution. Large-diameter distressing drilling is not in a wide use in the Polish hard coal mining. High-pressure water injection into coal seams is widely used, not so often into rocks surrounding seams. Scientific research has shown that in case of an appropriate watering of a coal seam its uniaxial compression strength decreases by 20–40%, and in case of a sandstone with a clay binder even by 60% and thereby the coal’s and rock’s liability to dynamic disintegration can be reduced (Kabiesz, Konopko 1996).


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Water infusion of a seam may be carried out by means of short holes drilled from a longwall face or drivage as well as by means of long holes bored from gate roadways. Instructions to the water infusion process have been worked out which comprehensively set out needed tests and examinations, technologies as well as measures of controlling the process’s effectiveness. For about 10 years the so-called directed fracturing of rocks accomplished by the hydraulic technique (DHF) or blasting (DFB) has been implemented in the Polish collieries. It established a new quality in the sphere of technology-bound bump hazard control (Konopko et. al. 1997). The directed hydraulic fracturing of rocks (DHF) includes: – drilling a bore-hole into the rock strata in which a fracture (strata separation) should be produced, – producing in the hole a nucleation fracture (figure 5), – sealing the hole, – injecting water under an appropriate pressure and at an adjusted flow rate into the area of the nucleation fracture. Water pressure concentrated in the sharp end of the nucleation fracture causes propagation of the fracture along its plane. When the pressure and the flow rate are high enough the range radius of the hydraulic fracturing may in practice reach a value ranging from 15 to 25 m, and even more. The directed hydraulic fracturing of rocks enables a simple, effective, and inexpensive changing of technological properties of rocks, and so enables improvement of mining conditions, for example it is used: – In order to reduce bump hazard and/or to improve roof cavability—monolithic strata of solid rocks may be divided into a series of thin slabs and/or blocks. That results in diminishing the roof-caving step and the seismic activity generated by straining the rock strata. – In order to improve maintenance of gate roadways or longwall faces—using DHF makes it possible to induce caving of a hanging roof (instead of blasting). – In order to improve workability of a seam and to reduce dustness dust suppression, and in case of high longwall faces—to restrict coal face slabing. From research it follows that fracturing takes place at a fluid pressure of:

where —rock pressure (resultant) formed as a superimposition of the primary rock pressure following from the depth, past mining events in a seam and neighbouring seams, geologic troubles, etc., R r—immediate tensile strength of the rocks in the plane of the generated fracture. In the hole there may be produced a single radial nucleation fracture near the hole’s bottom as well as several nucleation fractures. The depth of the holes, their configuration, as well as the number and location of the nucleation fractures in individual holes depend on the structure and properties of the


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rocks to be fractured and on the goal of fracturing. The aforementioned parameters should be defined in a plan of the directed fracturing of rocks. The directed fracturing of rocks through blasting (DFB) is derivative to the directed hydraulic fracturing (DHF). Blasting fumes produce the pressure in a hole in the area of the nucleation fractures.

Figure 5. Diagram of the blast hole’s construction. a—with one nucleation fracture, b—with several (three) nucleation fractures, c—with one longitudinal nucleation fracture 6.3. Organisational measures Bump hazard is a multi-cause phenomenon. It is more difficult to recognise than any other mine hazard. Therefore there is a need to engage a bump service with a multi-level organisational structure. In the Polish mining at the level of a mine there is a Mine Team for Bumps. Its task is to work out plans for bump-hazard prevention, continuous monitoring of bump hazard and implementation of the active bump control measures (in co-operation with the staff of mining sections of a mine). At the level of the hard coal mining and the copper ore mining there are Commissions for Bumps brought into


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existence by the President of the Superior Mining Office. Their task is to initiate new research directions in the field of rock bumps. Their core activity is focused on performing analyses and giving opinions on long-term plans for extraction of bumphazardous seams in individual mines as well as possible modifications in those plans. The Commissions for Bumps are composed of practising engineers as well as researchers in the fields of geotechnics, mining and geophysics. In cases of an extremely high bump hazard the commissions give opinions, and often also co-operate in working out detailed solutions within technologies of specific mining operations and within bump control methods. When a bump hazard condition is intolerably high they propose to desist from extraction of specific deposit parts. Bump service staff of a mine conducts continuous analyses of the current bump hazard state in each mine working. Depending on their results they determine what functions some workings may fulfil, and restrictions regarding the stay of mining teams there and even their travelling along them. In extreme cases a complete prohibition of miners’ stay in such workings is introduced, and the workings are inspected based on special guidelines and instructions approved by the General Manager of the mine. In areas with increased bump hazard state the so-called zones of extreme bump hazard are determined. Mostly such a zone is instituted in gate roadways and embraces 50–200 m ahead of a longwall face. Appropriate signs mark the zones. Restrictions are laid on the crew’s stay in direct vicinity of supports, supports’ reinforcement elements, prohibition of storing supply materials, etc. During extraction shifts, in course of winning, there is prohibition for the crew to stay in such zones. The zones of extreme bump hazard may be moving, e.g. a section of a gate roadway ahead of an extraction front will move as the front advances. They may also be stationary, e.g. in the reach of influence of remnants, abutment edges, pillars, faults or other factors intensifying bump hazard. In workings with a high bump hazard that fulfil given functions in technological processes, e.g. haulage of mined coal, stationary work places, e.g. conveyor operators, are eliminated. Workings, hoppers and conveyor transfer points are monitored by means of an industrial television system. When bump hazard is high one of precautions being in use is the so-called waiting time which means a period of time that have to elapse from a blasting operation to the return of a crew to a face or other bump-hazardous working or its part. The waiting time is usually determined to last from 30 minutes to 8 hours (to an extraction shift). In particular a long waiting time is determined after concussion blasting and blasting, and in copper ore mines—after blasting groups of faces. 6.4. Individual protection means A rock bump is a violent phenomenon and its effects in the form of destruction of long sections of headings equipped with modern arch supports as well as of working faces with powered shield supports do not give much opportunity for means of individual protection to effectively protect the crew. Nevertheless it seems to be legitimate to recommend using helmets with a larger than hitherto impact-resistance, among others, to sideways impacts, provided with a military-like chin strip. Boots with protection


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reinforcements for toes and heel may restrict the noxious injuries caused by rock lumps falling out of walls of workings. In collieries often other hazards like methane hazard or fire hazard co-occur with bump hazard. In particular as a result of crushing a coal body and bursting coal out of a seam and fracturing surrounding rocks and thereby “making the rock-mass pervious”— surge inflows of methane into the working, making the atmosphere irrespirable, may occur. Therefore it is legitimate to equip the crew employed in such conditions with selfcontained self-rescuers and oxygen respirators. Yet, in general it is feasible to safeguard any working crew by effective bump hazard prevention, i.e. by comprehensive applying the aforementioned preventive measures— beginning with identification of the hazard’s sources and determining its state and further by organisational activities where the particular focus should be put on training the supervisory staff and working crew. 7. THE DIRECTIONS OF DEVELOPING BUMP PREVENTION The comprehensiveness of means applied in recognising the state and sources of bump hazard as well as the wide spectrum of methods used within the long-term (i.e. in the sphere of mining techniques) and active (immediate) methods of controlling this hazard and the gained results indicate that the assumed directions of action are right. Yet, each bump—is the one that should never have occurred, something that causes casualties and damage to property. Therefore the methods being in use should be improved and new, more effective, satisfying requirements of modern mining ones should be worked out. It is to be expected: – Elaborating of the method of passive tomography as the one that gives a chance for recognising the state of sources and continuous monitoring of bump hazard, therein measuring bump hazard state in the scale from “a” to “d” in compatibility with other methods. – Widening of the measurement range of seismologic surveys, recording and localising tremor focuses with Eo≥102J and finally working out and implementation of measuring instrumentation with greater dynamics. The above will simplify and make more accurate the seismologic method. – Perfecting analytical methods aimed at improving their preciseness in localising stress concentration zones. That will allow reducing the number and extension of places of application of the active methods to those ones, which are really bump hazardous and thereby cut costs of that expensive technology. – Working out a possibly low labour-consuming method for current investigation of a bump hazard state, e.g. a radar method. – Working out an active method of bump hazard control that would reduce the hazard in a whole mining section or mining area. There is a chance that DHF and/or DFB would become such methods after appropriate improvements in the equipment for producing the nucleation fractures and implementation of an appropriately performing drilling equipment were made. That would create an opportunity to increase concentration of


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working faces and thereby to fulfil the criteria of economical mining of bumphazardous seams. All in all, it is legitimate to state that bump hazard may be effectively controlled through a properly planned and implemented measures that include the hazard’s state and sources, an appropriate plan for operations, current monitoring of the hazard’s state, purposeful selection of methods and means and implementation of the active bump control. The effectiveness of the whole process of the above activities depends on the reliability of implementing individual measures, i.e. on the knowledge and commitment of all services responsible for safety at work. Mining operations may be safely carried out even in high states of natural hazards, the bump hazard inclusive. CONCLUSIONS The Polish underground mining, mainly hard coal and copper ores mining is characterised by a very high potential of seismic hazard as well as connected with it— rockburst hazard. Hence, preventive rockburst prophylaxis is an extremely important element of exploitation technique, which has a significant impact both for a work safety level as well as for the efficiency of production processes. In the paper, the present state of solutions in the field of widely conceived preventive activities is presented. For many years, these activities have been widely applied in Polish mines where the rockburst hazards occur. The results of many years investigations and practical experiences indicate that their reasonable application requires not only getting acquainted with symptoms of the rockburst in underground work environment but also the connection of this hazard with other mining hazards which may occur at the same time. They point out that necessary condition for improvement of preventive actions as well as their efficiency is evaluation of the local and regional reasons of a rockburst hazard. The preventive rockburst activities presented in the paper are divided into two groups: methods for evaluation of the state of rockburst hazard and methods for its combating. The essential role of geophysical methods in recent years in the process of evaluation of the hazard as well as significance of designing the mining works as an important element of long-term prevention in rockburst hazard have been emphasised. In a more detailed way there are presented new solutions in the field of an active prophylaxis of rockburst hazard, such as directional hydraulic rock fracturing and directional rock fracturing by blasting methods. The attention is also drawn to the significance of such elements of rock-burst prevention as work organisation in the conditions of rockburst hazard as well as personal protection. The recapitulation of the paper consists determination of the trends of further activities in the field of rockburst prevention.


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REFERENCES Butra J. 2002: The Development of Methods of Bump Prevention in Copper Ore Mining. International Scientific—Technical Symposium on Rockbursts 2002—The State of Research and Prevention. The Publishing House of the Central Mining Institute, Katowice. Dubiński J. 1989: The Method of Premonitory Assessment of the Bump Hazard in Collieries. Works of the Central Mining Institute, Katowice. Dubiński J. 2000: Causative Relationships of Tremors and Bumps. Mining Review, No. 2, Katowice. Dubiński J., Konopko W. 2000: Rockbursts—Assessment—Prediction—Control. The Publishing House of the Central Mining Institute, Katowice. Dubiński J., Lurka A., Mutke G. 2000: Monitoring the Seismic Hazard of the Longwall Face N-303 in “Bielszowice” Colliery by Means of the Method of Passive Tomography. Proceedings of the Winter School of Rock Mechanics. Ed. University of Mining and Metallurgy, Cracow. Dubiński J., Mutke G. 2002: Application and Development Directions of Geophysical Methods for Assessing Bump Hazard State. International Scientific-Technical Symposium on Rockbursts. The Publishing House of the Central Mining Institute, Katowice. Kabiesz J., Konopko W. 1996: Instructions for Applying the Method of Bump Control by Water Infusion of Coal Seams. Ed. Central Mining Institute. Series: Instructions No. 2, Katowice. Kidybiński A. 2003: The Rock Bump Hazard in the World Mining—Recognising and Control. Research Works of the Central Mining Institute—Mining and Environment, No. 1, 2003. Central Mining Institute, Katowice, 2003. Konopko W. 1994: Comments on Planning Extraction of Bump-Hazardous Seams. Mining Review No. 2, 1994. Konopko W. 1994a: Experimental Fundamentals of Classifying Mine Workings in Collieries Among the Degrees of Bump Hazard. Research Works of the Central Mining Institute, No. 795. Katowice, 1994. Konopko W. et.al. 1997: The Directed Hydraulic Fracturing of Rocks and Possibilities of using it. Research Works of the Central Mining Institute, No. 824, Katowice. Principles and Range of Applying the Integrated Method for Assessing the Bump Hazard State in Collieries, 1996. Ed. Central Mining Institute. Series: Instructions, No. 1, Katowice.

Technical and Economic Possibilities for Permanent Limitation of Water Inflow to Mine Workings

Roman Kuś G.Janik & R.Kuś PRGW Sp. jawna. Sławków, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: The subject of the paper is to analyse the possibilities of permanent elimination of water inflow to underground and opencast mine workings on the basis of the currently available technologies. There is a presentation of the up-to-date experience in this area on the basis of the works carried out in Poland and worldwide. Ecological and financial results of water inflow limitation to the mining workings will be demonstrated. KEYWORDS: Hydro insulating barrier, hydro insulating solution, deposit sealing, economic calculation

1. INTRODUCTION In the mining industry, hydro-insulating barriers have been made, with few exceptions, for the purpose of water hazards elimination or limitation of the harmful influence of water pumping on the environment. Rarely have such activities been undertaken for improvement of the financial results of mining, however it seems that there are considerable possibilities for improvement in this area, considering in particular the fact that mining activities are usually carried out for many decades and the drainage costs can only be expected to rise with the mining depth, increased energy prices and ecological fees. Water inflow to the closed underground coal mines in Poland reaches the yearly volume of about 90 million m3 (Dzbik et al. 2003). Annual water inflow to the “Bełchatów” brown coal opencast mine is 180 mln m3. Limitation of the inflows means multi-million PLN savings per year.

Technical and economic possibilities for permanent limitation of water inflow

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2. CREATING HYDRO-INSULATING BARRIERS IN THE MINING INDUSTRY Various technologies depending on hydro-geological conditions are used for limiting and stopping water inflow to underground workings. Such technologies as rock-mass freezing, sealing by cementation, technologies using chemicals, sodium silicate injections have been widely known and applied in the mining industry for several decades. Each of the above mentioned methods could be effective only for solving a specific hydrogeological problem. Due to high maintenance costs, the application of the universal method—rock-mass freezinghas been limited mainly to the works connected with shaft sinking. An exception is the application of this technology for creating a counter-filtering barrier in one of the opencast gold mines in Canada, where it was assumed that paying the costs of maintaining a counter-filtering barrier was more economical, cheaper and more environment friendly solution than draining the rock-mass around the mine. Sealing works in the areas of increased water inflow to mine workings, carried out from the surface or from the level of the workings, are aimed at limiting or stopping of water inflow and utilize cementation technology as well as a wide range of chemicals and mineral agents. However, the cementation is the most commonly used technology. In Germany, South Africa, Canada and Great Britain cementation is the main method used for limiting or removal of water inflow to mine workings. In Great Britain and Germany cementation is used in 80% of shaft sinking and in South Africa in 100% of shaft sinking. Creating hydro-insulating barriers by injection of sealing agents through injection holes to the rock-mass allows limiting water inflow to a considerable degree. The best results are achieved by building counter-filtering barriers in the rocks of the following types: sandstone, limestone, granite etc. The following features are of particular importance for the proper designing of hydro-insulating barriers within the water-bearing layer: – extent of fissure opening, – filtering properties of solid rock, – water chemistry and pressure. Various hydro-insulating solutions are applied depending on the fissure-spread rate, karst phenomena and thickness of the water-bearing layer. Sealing thick and fractured waterbearing layers, characterized by high water-bearing capacity, requires applying special fast-stabilizing solutions. Applying different grades and brands of cement with different fillers does not change the fact that these are still unstable mixtures characterized consequently by giving off water from cement stone. The technology of building the hydro insulating barriers in solid rocks has been described in detail by A.C.Houlsby (http://www.users.bigpond.com//houlsby) (Houlsby 1990). As a result, in case of sealing the fissure rock-mass with cement solution, significant portion of unconsolidated water is the reason for some water carrying channels to remain open in the sealed rock-mass. Giving off water is the reason why some portion of cement does not react, which consequently leads to a situation in which an unconsolidated portion of cement is carried outside the fissure system, additionally reducing the hydro-insulating quality of a barrier


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in this way. In order to extend the application scope of cement mixtures, increase cement permeability, achieve better stabilization, limit giving off water, reduce wash-out and assure resistance to aggressive water, the following additives are used: bentonite, various chemicals, polymers and stabilizers. Additives of the above-mentioned plasticizers allow the reduction of the volume of water given off from cement mixtures. Increased stability of cement mixtures is also achieved by intensive mixing in highspeed mixers and by using cements of specific surface of 5000 cm2/g and bigger (Kuś, Popov 2000). In order to reduce the costs of cement mixtures, various additives improving cement-setting properties are applied such as fine ground slag, limestone sludge, chemical processed waste and the post-flotation process. Additives of cement substitutes having necessary setting properties have been widely used in South Africa, USA and recently in Poland. In addition to the unsatisfactory effectiveness of cement sealing, applying this type of sealing in the areas affected by mining works leads to the situation in which, as a result of rock-mass stress, fissures filled with cement stone crack, and eventually the hydro insulating barrier starts leaking. Hydro insulating barriers made of cement can be used for limiting underground water inflow in the areas not endangered by the rock-mass deformation caused by mining works. It has been recently attempted to use various types of solutions based on synthetic resins for hydro-insulating works. However, such methods cannot be widely applied due to the high costs and complicated technology connected with the preparation and execution of sealing the works. The common features of all the sealing works carried out with the application of hardening sealing solutions are the following: – high material costs, – stiffness and consequently low plasticity of hardening materials, – danger of the material setting in pipes due to non-adherence to injection standards or as a result of equipment failure, – considerable loss of time resulting from the necessity to drill out cement plugs in the injection hole, – possible deflection of injection holes during drilling through hardened sealing material of excessive strength, – necessity to carry out the sealing works in many cycles, with the breaks necessary for sealing material to harden, – low sealing effectiveness in the case of thick and cracked water bearing layers. Choice of the technological process of carrying out the sealing works from the surface or from underground mining workings depends mainly on the equipment applied, the scope of work and different practices applied in different countries. For example, in Great Britain cementation from the surface is carried out to the depth of 200–300 m, in Canada, Germany and Belgium—to 700–900 m, and in South Africa to the depth of 1500 m. Considering non-uniform fracturing of water-bearing layers, effectiveness of hydroinsulating works can be only evaluated in a subjective way because a single test hole does not provide enough characteristics concerning insulating properties of a hydro-insulating barrier. The maximum injection pressure applied in case of cementation from the surface depends mainly on the equipment used and can range from 5 to 50 MPa. Choice of


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injection pressure is chiefly conditioned by practical experience. For example in Great Britain, the value of injection pressure is usually assumed as 1, 5–2 times higher than the hydrostatic pressure in the layer being sealed. In South Africa the value of injection pressure of hydro-insulating solution is assumed on the basis of the rate—0, 023 MPa per 1 m of the depth from the surface to the layer being sealed, not less than 3, 5 MPa. Analysis of the factors influencing the development of the techniques for making hydro-insulating barriers by cementation allows drawing the following conclusions: – sizes of hydro-insulating barriers designed for limiting water inflow to underground workings need not be big. Cement consumption is unreasonably high in the case of making the hydro-insulating barriers by cementation, due to washing out of the cement mixture by underground water during hardening process (stabilization) and large propagation range of cement mixture, – cementation does not produce expected results in the areas of wide-spread fissures and turbulent underground water flow, due to washing out of cement stone, – properties of cementation mixtures, in particular their sedimentation during hardening process and their tendency to crack, indicate low effectiveness of the cementation mixtures application for creating hydro-insulating barriers in complicated mining and geological conditions. Several recent years have seen the implementation to the mining industry practice the technique of creating hydro-insulating barriers with the application of poly-mineral clays as the main component of hydro-insulating solutions (Kipko et al. 1993). The method of making the barriers with the application of the hydro-insulating solution based on polymineral clays differs from the traditional cementation methods in the following elements: 1. the method is based on the scientifically analysed process of creating a hydroinsulating barrier in the area of a water-bearing layer, including the calculations concerning the barrier dimensions in the area of a water-bearing layer, number of injection holes, requirements concerning applied pressure as well as quality control of the sealing works, 2. the scope of engineering survey includes objective data concerning fissure characteristics and filtering properties of rock-mass, obtained through direct measurements and hydrodynamic tests carried out in holes as well as by means of the analysis of the historical geological data, 3. sealing works are carried out with the application of the hydro-insulating solution based on poly-mineral clays. Main rules concerning the sealing works carried out with the application of the hydroinsulating solution based on poly-mineral clays are presented below. The rules can be grouped as follows. – Survey of the test holes is carried out by specialized testing equipment. Thanks to this fact it is possible to obtain detailed data concerning hydrodynamic properties and fissure characteristics of an examined water-bearing layer. Obtaining this data is possible thanks to application of the hole flow indicators having the holes of the diameters 44, 57 and 76 mm. Application of theflow indicators enables at the same time to specify:


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– number of water-bearing horizons, – depth and thickness of each horizon, – hydrodynamic properties of each horizon. Additionally, other characteristics of each water-bearing horizon are evaluated, in particular: underground water pressure, filtration and permeability rates, open fissure ratio and average fissure spread in the water-bearing layer. Methods utilizing measurements of hydrodynamic phenomena by packer testing sets, allow estimating the silting rate of the near-hole area, layer hydrostatic pressure, permeability ratio, piezo-conductivity. Observation scope in the holes allows estimating the anisotropy rate of examined layers. Coordinating the hydrodynamic and geophysical tests with the results of the geological analysis assures providing thorough data concerning the water-bearing horizons. Such data constitutes the basis for designing hydro-insulating barriers. – On the basis of the main data and the results of the surveys the sealing works are designed. Applied hydro-insulating solutions based on poly-mineral clays have high consistency. Practically, they are not washed out by underground waters, do not stabilize during injection in holes and fissures and very quickly achieve plasticity strength when their injection process is stopped. As a result of rock-mass stresses and shocks the solutions gain their plasticity. Strength and geological properties of the hydro-insulating solution are to a wide extent conditioned by the properties of a base solution made on the basis of poly-mineral clays. This solution should have minimum plasticity strength in the range 0, 5–0, 8 MPa. By experiments the strength and geological characteristics of the hydro-insulating solution have been specified: – strength—on the basis of the calculations—size of the hydro-insulating barrier, – potential of utilized injection equipment—on the basis of technical analysis— possibleoutlines of solution expansion from one injection hole. Required number of injection holes, their location, volume of hydro-insulating solution for the needs of designed barrier are specified by calculations. – The equipment enabling vertical hole drilling as well as directional hole drilling is used for drilling of the injection holes. In case of drilling from underground workings, the holes are equipped with the devices protecting from inflow of the solution to the working (dampers, packers, packing gland). – In all the drilled injection holes comprehensive set of tests by flow indicators and test results analysis are carried out. On the basis of the data obtained by hole testing, the corrections of design calculations are made, with reference to the size of designed barrier and volume of hydro-insulating solution which assure sealing of water horizon. – Injection of hydro-insulating solution is carried out by pump units, which assure the pumping capacity of 20–600 dm3/min. The pumping capacity is smoothly regulated in relation to the rock-mass absorption capability. The pump units allow injecting at a pressure of up to 40 MPa. Injecting is carried out through an injection hole to the water-bearing layer or to separate fracture zones separately, according to a previously defined technological plan. It is possible to carry out the works according to such a plan thanks to the application of packers. In the case of sealing works carried out from


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underground workings, the first water-bearing is sealed through the conductor pipe and next by the packer. – In order to prepare the clay base solution it is necessary to use a specialized set of clay mixers, vibrating meshes, pumps, tanks. A properly assembled set of equipment can produce about 300 m3 of solution daily. – Monitoring the effectiveness of the works plays an important role during the hydroinsulating works. The works effectiveness can be measured, for example, by water inflow to the sealed zone. Examples of effective hydro-insulating barriers have been described (Kuś, Polozov 2000). Analysis of the constructed sealing barriers allow to draw the following conclusions: 1. Creating hydro-insulating barriers is an effective method of isolating the mining workings from water inflow in jointed rock-mass. 2. Sealing solutions of the type conditioned by chemical composition and temperature of underground water are used for creating hydro-insulating barriers. 3. The most promising for constructing hydro-insulating barriers are the solutions based on poly-mineral clays (kaolinite, hydromica) which are characterized by durability, resistance to corrosion as well as they are cheap and non-harmful to the environment. 4. Classification and properties of clays allows choosing and preparing in a short time the optimal recipe for a sealing solution having required technological parameters. Sealing solutions based on poly-mineral clays have so far proved their high effectiveness and usefulness for constructing all kinds of hydro-insulating barriers in the mining industry and in hydro-technical facilities in complex geological and hydro-chemical conditions. 3. PROPOSAL CONCERNING LIMITATION OF WATER INFLOW TO THE “POLKOWICE-SIEROSZOWICE” MINE The main source of water inflow in the “Polkowice-Sieroszowice” mine is the waters located in base limestone. Multi-year drainage processes in mines cause very significant hydrodynamic changes directly in the drained water-bearing layer of base limestone, as well as in the layers of overburden, mainly occurring under coal Tertiary sediments, sandstone of the Bunter, and in between-coal Tertiary sediments. Drainage of overburden layers is carried out by hydraulic-sedimentation and tectonic interactions and by deep percolation. As a result of the long-lasting drainage of the rock-mass the water hazard from the under-Tertiary-period outcrops of the deposit layers has been considerably limited. There has been a water pressure drop in the Tertiary under-coal water-bearing layer to the roof level of this layer, which is expressed by the water pressure in the range of several atmospheres in the drainage holes drilled from the mining workings. There has also been a noticeable stabilization of underground water inflows to the mines from the side of the under-Tertiary outcrops, which suggests considerable depletion of static resources of underground waters in the area of Polkowice.


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Currently drainage of the mine is carried out by means of drains (directional holes). Water coming to the works through the directional holes located ahead of the headings, runs with the dip along the floor of the working and is collected in the pumping stations from where it is pumped out to the surface. An alternative solution to the passive drainage is to utilize favourable hydro-geological conditions for sealing the rock-mass and limiting water inflow. In connection with the above it is proposed to make a sealing screen designed for limiting water inflow to the under-Tertiary outcrops. In order to limit underground water inflow from W-1 dolomite it is proposed to create a hydro-insulating barrier made of a stable hydro-insulating solution based on polymineral clays. Two variants of barrier making have been considered: from the surface and from the underground workings. According to the preliminary calculations creating the hydro-insulating barrier at a distance of 2500 m requires injecting of about 90,000 m3 of hydro-insulating solution. Creating the database was the first stage of the analysis concerning the possibility of limiting water inflow. The database has been prepared on the basis of the information from the surface hole charts (27 holes) and the directional hole charts, testing-drainage holes (52 holes). The data concerning lithology and stratigraphy of layers, hole depth and location has been input. Additionally, the directional hole charts included length, drilling direction and inclination angle of specific holes, drilling dates, purpose of drilling (hydrogeological, exploration). The drainage holes have also the data concerning the inflows during drilling and the information specifying the volume of obtained water and the depth from which water was obtained in dm3/min. (capacity changes of the holes). On the basis of the data a 3-D model of the layers in this area was prepared. Models of the following layers were prepared: main layers, the Tertiary period roof, roof and floor of the limestone-dolomite series (W-1) as well as the simplified model of the surface. The 3-D model of the deposit was prepared by means of the MODELLER program from the I/Mine package in the MicroStation graphic environment. The boundaries of the area for which the model was prepared were limited to the most interesting area of the “Polkowice-Sieroszowice” deposit, touching the outcrops of the under-Tertiary sediments and covering the area of the hydro-geological window. Consequently, the southern and southwest boundary is based on a structure elevating the Zechstein sediments. The northeast boundary is closed on the line separating the “S” part of the deposit from “N” part, i.e. the deposit part which is dry and where inflows are not noticeable or are very small. In the northwest the area borders on OG “Sieroszowice”, in the east the boundary runs along the line between the holes S-98 and S-146. The area plan was digitised with a rectangular grid in which all the blocks were initially square. In this way 6439 calculating cells were created. The model design was limited to one layer—the layer of limestone-dolomite series. In previous modelling the roof and the floor of this layer were created on the basis of the hole charts. While selecting the value of the filtration coefficient it was assumed that the modelled hydro-geological layer is uniform and anisotropic. Due to the very high changeability of this coefficient in dolomite and limestone sediments the preliminary average value of this parameter was assumed k=1e−5. The final value of the filtration coefficient was obtained in the process of model calibration.


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In the model the boundary condition of the first type H=const. was applied which simulated constant water outflow at the specified height of water level/piezometric pressure. No other boundary condition was applied on the remaining model boundaries. In such a case MODFLOW automatically assumes the condition of the II type Q=0, i.e. along the current line. Inflow in the model was assumed only by means of the hydrogeological window presented as the condition “GHB”. The window in the model was presented as the area of uniform permeability, which was selected on the basis of tarring. It was assumed that the water coming through the window is located at the height of 136 m. This is the maximum height of water level measured by in the piezometer measuring the Miocene level. Assuming the maximum height of water level resulted from the wish to monitor the model in the least favourable conditions, i.e. in the moment when the maximum water volume comes to the model area. The height of the Miocene water level was assumed, which together with the Oligocene level is subjected to indirect drainage. The Quaternary and Pliocene (in the overburden) are isolated from the lower levels and are not drained by the mine. Water collection in the model is carried out mainly through the workings, which are presented as the drains. It was assumed that the drains relating to the headings most protruding in the direction of SW and W have the highest water permeability. Since there was no data concerning the value of this parameter, it was defined in the model verification process that this value ranges between 0,1 and 0,8 m2/d. The drains are “sunk” i.e. water level is above the drains. Some portion of water comes also to the southern workings of the “Polkowice-Sieroszowice” mine and the “Lubin” mine. In order to present this fact in the model, the boundary condition of the first type (H=constant) was assumed, which allows for assuming constant water outflow. The volume of water running through the eastern boundary is small and the flow takes place at the ordinate of water level H=499 m below sea level. The model takes faults into account. They are presented as narrow zones having a different filtration coefficient. The fault located in the southeast part of the model is an impermeable structure, which was proved by the observations carried out in the mine. The fault is also the model’s boundary. The remaining faults are presented as narrow zones of k=1e−6. Calibration of the model was carried out using the method of successive approximations. After the model verification, the difference between the location of underground water level resulting from the piezometer measurements and the values obtained from the model calculations was up to 0, 05 m. Water balance error, i.e. the difference between the total outflow and inflow to the model, was 0,001. Simulations of the water level behaviour were carried out on the model presenting the hydrogeological conditions in the area of interests. The sealing barrier is located to the west from The barrier runs from the insulating zone on SW part of the model to the line separating the “S” the incline G–12, G–12a, G–13 and is parallel to the incline. The assumed length is about 2, 5–3 km. zone from the “N” zone of the deposit and limiting the area without inflows. “The Wall” function in the MODFLOW program was used to present the barrier. Inflow is stopped by applying 15 m wide hydro-insulating barrier having the filtration coefficient k=1e−9. Considerable limitations are noticeable when the barriers of k= 1e−8 are applied.


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4. ECONOMIC CALCULATION OF THE PROJECT Profitability evaluation concerning the sealing barrier making was based on the price calculation for the designed scope of work and the cost calculation of the main pump system including the costs of the mine water discharge to the Odra River. Drainage costs calculation: – Drainage cost for the area of the section G-33:

K=1,35 PLN/m3 – Current inflow in the area of the section G-33:

QTOTAL=19200 dm3/min=19,2 m3/min Consequently: – Yearly drainage cost:

KYEAR=K×QYEAR KYEAR=1,35 PLN/m3×10 091 520 m3/24 hours=13 623 552,00 PLN Sealing costs calculation: – Expected cost of creating the sealing barrier from the surface amounts to: 58 379 804,00 PLN/net/. – Expected cost of creating the sealing barrier from the mine workings amounts to: 33 723 420,00 PLN /net/ including the costs of making: – driveage work: about 2×1085 m, – 12 drilling cabbies, – 8 technological cabbies. 25 185 420,00 PLN /net/ assuming that the costs of the mining works will be included by O/ZG “Polkowice-Sieroszowice” in the working costs. Summary of the costs of creating the sealing barrier and drainage costs for the area of the section G-33 Cost of creating the sealing barrier from the surface amounts to: 58 379 804,00 PLN /net/ Cost of creating the sealing barrier from the mine workings amounts to:

24 hour and yearly drainage cost for the area of the section G-33 at current inflow Q=19,2 m3/min


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33 723 420,00 PLN /net/ including the costs of making: – driveage work: about 2×1085 m – 12 drilling cabbies – 8 technological cabbies Cost of creating the sealing barrier from the mine workings amounts to: 25 185 420,00 PLN/net/

24 hour drainage cost K24 HOUR =37 324,80 PLN

assuming that the costs of the mining works will be included by O/ZG Yearly drainage cost “Polkowice-Sieroszowice” in the working costs KYEAR=13 623 552,00 PLN

Calculation factors for the investment economic viability analysis: 1. Time required for creation of the sealing barrier from the surface—24 months. 2. Time required for creation of the sealing barrier from the mine workings—20 months. Remark: specified times concern drilling and injection but do not include the time required for preparation works. 3. Drainage cost at current inflow rate and considering valid tariffs for pumping and underground water discharge:

K1 year=13 623 552,00 PLN. 4. Cost of creation of the sealing barrier from the surface: 58 379 804,00 PLN. 5. Cost of creation of the sealing barrier from the mine workings: 33 723 420,00 PLN. 6. Pumping cost in the period of creating the sealing barrier from the surface (24 months) at current inflow rate and considering valid tariffs for pumping and underground water discharge: 27 247 104,00 PLN. 7. Pumping cost in the period of creating the sealing barrier from the mine workings (20 months) at current inflow rate and considering valid tariffs for pumping and underground water discharge: 22 705 920,00 PLN. Economic viability evaluation of creating the sealing barrier from the surface Inflow volume reduction by 50%: 1. Pumping costs reduction for 1 year: 6811 776, 00 PLN. 2. Pumping costs for 1 year after sealing: 6 811 776, 00 PLN. 3. Investment return period: 58 379 804,00/ 6 811 776,00≈8,5 years. 4. Assumed, proportional inflow drop per month during sealing works (period of 24 months) “-”QMONTH=0,4 m3/min (from QINITIAL 19,2 m3/min to QFINAL 9,6 m3/min). 5. Proportional, monthly pumping cost drop during sealing works assumed for the calculations (from QINITIAL 19,2 m3/min to QFINAL 9,6 m3/min), resulting from inflow reduction by “-”QMONTH=0,4m3/min. /see the table below/


No. 1 month 2 month 3 month 4 month

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Calculation of monthly pumping cost during sealing works, considering the assumed monthly reduction of inflow volume “-”QMONTH=0,4 m3/min—QINITIAL=19,2 m3/min 18,8 m3/min×60×24×30 days×1, 35 PLN/m3 3

3

18,4 m /min×60×24×30 days×1, 35 PLN/m

Amount [PLN] 1 096416,00 1 073 088,00

3

3

1 049 760,00

3

3

1 026 432,00

18,0 m /min×60×24×30 days×1,35 PLN/m 17,6 m /min×60×24×30 days×1,35 PLN/m

.. .. 22 month 10,4 m3/min×60×24×30 days×1, 35 PLN/m3 3

3

23 month 10,0 m /min×60×24×30 days×1, 35 PLN/m 3

3

24 month 9,6 m /min×60×24×30 days×1,35 PLN/m

606 528,00 583 200,00 559 872,00

The sum of pumping costs during the sealing works

19 875 456,00

3

Pumping costs at the inflow Q=19,2 m /min K2 Years=13 623 552,00 PLN×2 years Pumping cost reduction due to inflow reduction during sealing works

27 247 104,00 7 371 648

Economic viability evaluation of creating the sealing barrier from the surface Inflow volume reduction by 75%: 1. Pumping costs reduction for 1 year: 10 217 664,00 PLN. 2. Pumping costs for 1 year after sealing: 3 405 888,00 PLN. 3. Investment return period: 58 379 804.00/10 217 664,00≈5,7 years. 4. Assumed, proportional inflow drop per month during sealing works (period of 24 months) “-”QMONTH=0,6 m3/min (from QINITIAL 19,2 m3/min to QFINAL 4,8 m3/min). 5. Proportional, monthly pumping cost drop during sealing assumed for the calculations (from QINITIAL 19,2 m3/min to QFINAL 4,8 m3/min) resulting from inflow reduction by “-”QMONTH= 0,6 m3/min. /see the table below/ No. 1 month 2 month 3 month 4 month .. ..

Calculation of monthly pumping cost during sealing works, considering the assumed monthly reduction of inflow volume “-”QMONTH=0,6 m3/min—QINITIAL=19,2 m3/min 18,6 m3/min×60×24×30 days×1,35 PLN/m3

1 084 752,00

3

3

1 049 760, 00

3

1 014 768, 00

18,0 m /min×60×24×30 days×1, 35 PLNł/m 3

17, 4 m /min×60×24×30 days×1, 35 PLN/m 3

3

16,8 m /min×60×24×30 days×1,35 PLN/m

Amount [PLN]

979 776, 00

Technical and economic possibilities for permanent limitation of water inflow 23 month 5, 4 m3/min×60×24×30 days×1, 35 PLN/m3 3

3

24 month 4,8 m /min×60×24×30 days×1,35 PLN/m The sum of pumping costs during the sealing works 3

253

314 928, 00 279 936,00 16 376 256,00

Pumping costs at the inflow Q=19,2 m /min K2 years=13 623 552,00 PLN×2 years

27 247 104,00

Pumping cost reduction due to inflow reduction during sealing works

10 870 848,00

Economic viability evaluation of creating the sealing barrier from the mine workings Inflow volume reduction by 50%. 1. Pumping costs reduction for 1 year /12 months/: 6 811 776,00 PLN. 2. Pumping costs for 1 year after sealing /12 months/: 6 811 776,00 PLN. 3. Investment return period: 33 723 420,00 PLN/6 811 776, 00≈5 years. 4. Assumed, proportional inflow drop per month during sealing works (period of 20 months) “-”QMONTH=0,48 m3/min (from QINITIAL 19,2 m3/min to QFINAL 9,6 m3/min). 5. Proportional, monthly pumping cost drop during sealing assumed for the calculations (from QINITIAL 19,2 m3/min to QFINAL 9,6 m3/min) resulting from inflow reduction by “-”QMONTH= 0,48 m3/min. /see the table below/ No.

Calculation of monthly pumping cost during sealing works, considering the assumed monthly reduction of inflow volume “”QMONTH=0,48 m3/min—QINITIAL=9,2 m3/min

Amount [PLN]

1 month

18,72 m3/min×60×24×30 days×1,35 PLN/m3

1 091 750,40

2 month

18,24 m3/min×60×24×30 days×1,35 PLN/m3

1 063 756,80

3 month

17,76 m3/min× 60×24×30 days×1,35 PLN/m3

1 035 763,20

4 month

17,28 m3/min×60×24×30 days×1,35 PLN/m3

1 007 769,60

19 month

10,08 m3/min×60×24×30 days×1,35 PLN/m3

587 865,60

20 month

9,60 m3/min×60×24×30 days×1,35 PLN/m3

559 872,00

The sum of pumping costs during the sealing works:

16 516 224,00

Pumping costs at the inflow Q=19,2 m3/min K20 months=(13 623 552,00/12)×20 months:

22 705 920,00

Pumping cost reduction due to inflow reduction during sealing works:

6 189696,00


254

Economic viability evaluation of creating the sealing barrier from the mine workings Inflow volume reduction by 75%: 1. Pumping costs reduction for 1 year /12 months/: 10 217 664,00 PLN. 2. Pumping costs for 1 year after sealing /12 months/: 3 405 888,00 PLN. 3. Investment return period: 33 723 420,00 PLN/10 217 664.00 ≈ 3, 3 years. 4. Assumed, proportional inflow drop per month during sealing works (period of 20 months) “-”QMONTH=0,72 m3/min (from QINITIAL 19,2 m3/min to QFINAL 4,8 m3/min). 5. Proportional, monthly pumping cost drop during sealing assumed for the calculations (from QINITIAL 19,2 m3/min to QFINAL 4,8 m3/min) resulting from inflow reduction by “-”QMSC=0,72 m3/min. /see the table below/ No.

1 month 2 month 3 month 4 month

Calculation of monthly pumping cost during sealing works, considering the assumed monthly reduction of inflow volume “-”QMONTH=0,72 m3/min—QINITIAL=19,2 m3/min 1 8,48 m3/min×60×24×30 days×1,35 PLN/m3

Amount [PLN] 1 077 753,60

3

3

1 035 763,20

3

3

993 772,80

3

3

951 782,40

17,76 m /min×60×24×30 days×1,35 PLN/m 17,04 m /min×60×24×30 days×1,35 PLN/m 16,32 m /min×60×24×30 days×1,35 PLN/m

18 month

3

6,24 m /min×60×24×30 days×1,35 PLN/m

363 916,80

19 month

5,52 m3/min×60×24×30 days×1,35 PLN/m3

321 926,40

20 month

3

3

3

4,80 m /min×60×24×30 days×1,35 PLN/m

279 936,00

The sum of pumping costs during the sealing works:

13 576 896,00

Pumping costs at the inflow Q=19,2 m3/min K20 months=(13 623 552,00/12)×20 months:

22 705 920,00

Pumping cost reduction due to inflow reduction during sealing works:

9 129 024,00

Investment return analysis has been carried out also by: Joanna Kulczycka and Karol Konieczny from Pracownia Badań Strategicznych Instytutu Gospodarki Surowcami Mineralnymi i Energią PAN w Krakowie (Strategic Research Laboratory of the Institute of Mineral Raw Material and Energy Management of POLISH ACADEMY OF SCIENCE in Cracow) and by Michał Gientka from Zakład Geologii Gospodarczej PIG (PIG Economic Geology Department). The results of this analysis are shown in the below table. The summary of expenditures, operating costs and their reduction (potential income) and investment return period; period of 15 years (including the investment period). Version: Effects

Current

Sealing from Sealing from Sealing from Sealing from the mine the surface, the surface, the mine effectiveness effectiveness workings, workings, (75%) (50%) effectiveness effectiveness


Investment expenditures Operating costs (15 years)

255

(50%)

(75%)

33 723 420

33 723 420

−204 353 −167 084 396 −119 260 052 −141 292 716 280

−92 900 724

–

58 379 804

58 379 804

Cost reduction (vs. current version)

–

37 268 884

85 093 228

63 060 564

111 452 556

Investment return period (without the investment period)

–

7 years 7 months

4 years 9 months

4 years 1 month

2 years 6 months

Investment return period (including the investment period)

–

9 years 7 months

6 years 9 months

5 years 9 months

4 years 2 months

REMARKS – Costs of experimental part of the project have not been taken into account. – Investment expenditures are assumed in the amount, which does not take into account the possibility of their settlement in O/ZG “Polkowice-Sieroszowice”. – Investment return period is specified in two versions because cost reductions will take place already during the time of executing the project. – If discount rate at the level of 8% is taken into account the investment return period will be longer on average by 40%. Regardless of applied calculation method, almost identical results concerning the investment return periods were obtained. From the economic point of view, creating the sealing barrier from the mine workings would be the most profitable. In such a case the average investment return period is about 3,3 years at 75% sealing effectiveness and about 5 years at 50% sealing effectiveness (discount rate is not taken into account). When technical and geological aspects of creating a barrier are considered, the probability of achieving the sealing effectiveness above 75% is decisively higher when the work is carried out from the mine workings than in case when it is carried out from the surface. This additional factor should be also taken into account while deciding on the method for carrying out the deposit sealing works. REFERENCES Dzbik J., Stachowiecki J., Wodecki J. 2003: Doświadczenia techniczno-organizacyjne związane z działalnością Centralnego Zakładu Odwadniania Kopalń w Czeladzi. Materiały Konferencji Szkoły Eksploatacji Podziemnej, Krakow 2003, 407–413. Houlsby A.C. 1990: Construction and Design of Cement Grouting: A Guide to Grouting in Rock Foundations. John Wiley & Sons, New York 1990.


256

Kipko Y. et al. 1993: Integrated Grouting and Hydrogeology of Fractured Rock in the Former USSR. Society for Mining Metallurgy & Exploration Inc. Colorado 1993. Kuś R., Polozov J. 2000: Features Required from Grouting Solutions—Practical Application. 7th International Mine Water Association Congress—Mine Water and The Environment, Ustroń.2000, 565–576. Kuś R., Popov A. 2000: Ocena wpływu powierzchni jednostkowej fazy stałej w roztworach uszczelniających na ich właściwości migracyjne i przeciwfiltracyjne. “Stabilizacja masywów skalnych w podłożu budowli hydrotechnicznych”—materiały konferencyjne. Instytut Meteorologii i Gospodarki Wodnej, Warszawa 2000, 111–121.

The Attempt to Apply Radar Interferometry InSAR in the Monitoring of the Impact of the Ore Deposit Exploitation in LGOM (Lubin Copper Mining Area)

Edward Popiołek AGH—University of Science and Technology, WGGiIŚ. Krakow, Poland Cezary Bachowski KGHM Polska Miedź S.A. Lubin, Poland Artur Krawczyk, Pawel Sopata AGH—University of Science and Technology, WGGiIŚ. Krakow, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: In the paper the possibilities of using satellite radar scanning of the area with InSAR method in the determination of area subsidence caused by underground mining. It was proved that it is possible to determine temporary subsidence of the area, zones of present threat to the buildings as well as to verify the forecasts for the deformation of the surface of the area. This way there are the possibilities of the monitoring of subsidence without the need for establishing a surveying grid. KEYWORDS: Radar interferometry, satellites, studies on area deformation, information systems

1. INTRODUCTION The purpose of this paper is to present the results of multifaceted analyses of the application of a new monitoring method referring to the changes in the area. This method—satellite radar interferometry—can undoubtedly be used to examine the deformations of the surface area in the regions of the influence of underground exploitation of copper ores deposits.


258

The duty of the observation of the impact of mining can directly be derived from the Geological and Mining Law and the obligations of the Law on Environmental Protection. One of the most important negative effects of mining influence includes the deformations of the surface of the area. They make a direct thereat to the constructions as well as the management of the area. Dynamic development of the technologies of remote sensing and getting information on the terrain led into the construction and implementation for civilian applications a satellite system of the visualisation of the area, realized by special satellites w in the way enabling us to get radar images in the intervals not longer than one month. Getting information from radar images of the area is the transposition of phase differences of two images. The result of this is one interferometric method (InSAR). Apart from the surveys and classic check-ups of continuous surface deformations in the form of subsidence troughs, further possibilities of applying this method appear—for example at the monitoring of decantation ponds. These are particularly important aspects, because they are directly bound with the requirements of providing common security by mining enterprises. Basic advantages of radar interferometry are: the possibility of observing surface changes of the area without the necessity to establish expensive grid of measurement points and the possibility of frequent information gaining in area of subsidence as well as getting information on the results of the exploitation, practically going back to 1990. One can say that radar interferometry eliminates basic drawbacks of surveying methods, moreover, it allows limiting expensive surveying grids. The result of carried out so far analyses is positive and justifies the implementation of radar interferometry into practice, as a new method in the monitoring of mining influence. 2. PRINCIPLES OF SATELLITE RADAR INTERFEROMETRY InSAR Satellite radar interferometry (InSAR) is a modern remote sensing technique to get information on the surface of the Earth in a strictly defined time interval, in the area of tens of square kilometres simultaneously. In this method SAR (Synthetic Aparature Radar) images obtained during repetitive runs of satellites (e.g. European Remote Sensing Satellite ERS-1 or ERS-2) over the same Earth surface are used. Taking radar images by a satellite involves the registration (with the help of SAR) of a radar wave reflecting from the area. During the registration the information on the intensity of the reflection of radar wave (i.e. the degree of the wave absorption) and its phase at the moment of reaching the receiver is obtained. This way one gets the information on the relative values of the ordinate of the area surface or its changes in time. In the field of the protection of mining areas an important property of radar images is the information on the phase difference between the emitted wave and the wave reflected from the same area in a different time. Having two such radar images one can measure the value of phases making the difference between signals and obtain an interferogram as a result (figure 1).

The attempt to apply radar interferometry InSAR

259

Figure 1. The fragment of the interferogram with interferometric lines presenting a local subsidence trough Interferogram makes a picture of phase differences of subsequent pixels of two SAR images (so-called master image and slave image). The differences of phase are presented in a standard way by a proper scale of colours, where a full sequence of colours reflects the change of the phase of radar signal by 360°. The observed on the interferogram change of phase by a full cycle (360°) is equivalent to the half of the length of the radar signal wave. In case of ERS-1 and ERS-2 satellites the length of the wave used by SAR equipment equals 5, 6 cm, i.e. full sequence of interferometric lines (phase change by 2p) presents the transformation of the area surface by 2,8 cm in the direction towards the detection radius. For scanning is done with the angle 23° (figure 2), with the reduction of altitude changes to normal in the area of the radar image, subsequent interference lines occur in places of the difference in area subsidence, which is about 2, 5 cm.

Figure 2. Geometry of radar signal


260

Basically, for the purpose of interferogram radar images showing mutual coherence of a proper value are suitable. This condition can indirectly cause some difficulties in the practical application of this method. Coherence is a parameter defining the proportion of spectral differences of the signal (phase and intensity) for the individual pixels of two SAR images. The value of coherence indicates changes that took place between making subsequent visualizations and the areas of weak coherence (dark) show places where the initial phase of the signal was changed by external factors. The coherence value is the most influenced by the following factors (Pratti et al. 1994): – base distance—the difference in the position of satellites in respective subsequent flights; – the value of the noise of radar signal and changes of phase connected with the measuring instruments; – physical changes of the environment taking place between subsequent SAR visualizations. Decoherence (the loss of coherence) is to a great degree dependent on the time that passed between taking subsequent visualisations (Villasenor, Zebker 1992; Usai, Hanssen 1997). 3. DETERMINATION OF TEMPORARY SUBSIDENCE TROUGHS For the interferogram is merely the image of changes taking place in a given area, no terrain details are visible in it. Thus it is not possible to directly adjust an interferometric image and calibrated topographic map. However, knowing the fact, that an interferogram arises based on radar images, there is the possibility of the calibration of interferogram after adjusting the radar image SAR. This guaranties identical and unambiguous covering the area of a radar image and interferogram by pixels. The calibration of a radar image was carried out in the same way as the adjustment of raster image of a topographic map to the system of co-ordinates, while the reference map was in this case a topographic map. Because of that, the first stage of calibration is fitting into the system of co-ordinates in a topographic map as a raster file. This calibration takes place in the Microstation program. After preliminary joining the raster image of a topographic map with a digital map of the surface, further adjustment is carried out, based on the existing terrain details. Then the radar image is read and calibrated to the system of the co-ordinates of a hybrid topographic map. During the calibration only the objects that allowed no ambiguity in their identification on both maps: topographic and digital are used. They allowed precise identification of the respective points (e.g. road crossings). Preserved parameters of the calibration of radar image were then used in the calibration of interferogram. Based on made in such a way set of parameters on a calibrated SAR image a respective interferogram was put. Calibrated this way interferogram makes a starting product and the base for the digitalisation of the isoclines of the area subsidence.


261

The digitalisation of the image of interferometric lines is carried out manually in the environment of MicroStation made by Bentley. Vector lines are brought in the places of the borders between the areas of the sets of points having the same value. This way—by digitalising subsequent interferometric lines—subsequent subsidence isoclines are obtained. The attributed subsidence value results from the number of colour lines contained in one full sequence of colours (in this case 0,86 cm per colour). The results of the digitalisation are processed in a computer program Surfer from the pack of Golden Software, obtaining a numerical image of subsidence isoclines. 4. THE DEFINITION OF THE REGIONS OF CURRENT MININGCAUSED THREAT The distribution of satellite-made radar images is possible in a short period of time, necessary for the transmission of digital data on a magnetic carrier. Making an interferogram from the last and the previous radar images takes a few days. In total it takes only about two weeks to obtain temporary subsidence troughs indicating the regions of current direct influence of mining exploitation on the surface of the area. Temporary subsidence troughs are obtained in the system of co-ordinates of a digital map of the area and mining map through the process of calibration. As a border of the regions of current threat to the area one should take the limit of the last external interferogram line representing the subsidence growth of the value above 8 mm.

Figure 3. The range of the influence of exploitation in the region of Polkowice (23.10.1995–01.01.1996) As a border of the regions of current threat to the area one should take the limit of the last external interferogram line representing the subsidence growth of the value above 8 mm.


262

In figure 3 an example of determining the regions of current threat to the area for the period 23.10.1995–01.01.1996 is presented. After the location of regions the following stage can involve the assessment of the number of threatened objects and, if possible the scale of this threat. 5. VERIFICATION OF THE FORECASTS FOR THE DEFORMATION OF THE AREA SURFACE 5.1. Determination of theory parameters from elementary troughs To determine parameters r and tgβ included into S.Knothe’s theory of influences—the calculation method based on the elementary character of a subsidence trough was applied (Batkiewicz, Popiołek 1972). During the survey (Popiołek 2002) it was concluded that registered and selected for calculation interferometric subsidence troughs fulfil this condition. The accepted calculation method allows the determination of the vale of the parameter of the dispersion of influences r, defined in the mentioned above theory. In the calculations the following formula was used (Batkiewicz, Popiołek 1972): (1) where w1—maximum subsidence value in a cross-section of the subsidence trough (in place x1=0), w2—value of subsidence on the side of the subsidence trough (in place x2). Based on the calculated this way r, knowing the depth of the deposit H, the value of “rock-mass parameter” tgβ can be calculated using linear function relationship: (2) To define tgβ parameters, for underground exploitation of copper in LGOM, six subsidence troughs were selected. They were generated by InSAR technology. The distribution of subsequently numbered from S1 to S6 troughs are presented in figure 4. Through the centres of subsidence troughs a few longitudinal sections were made. Thus values x2 and respective subsidences w2 (to the formula (1)) were obtained from the whole spatial image of the subsidence trough. As the result of these operations several dozens of parameter values of dispersion and tgβ parameter were obtained from each section (with a knowledge on deep exploitation in a given region). These data were analysed and selected, values significantly different from initially calculated average value were rejected. Selected this way data were averaged and average values of tgβp and rp, based on a defined section of a given subsidence trough were determined. Calculation results as well as the values of calculated tgβ and r in the regions of the occurrence of subsidence troughs are presented in table 1. The values of calculated in the table above parameters tgβ are not different from the values accepted so far in the influence forecasts. In our opinion the differentiation of ultimate tgβ values in the range 1,26–1,69 reflects the changeability of the rock-mass in the area of LGOM. To characterise the rock-mass in LGOM a mean value can be


263

calculated. For the area of LGOM covered by selected for the analysis elementary subsidence troughs (figure 4) it is tgβśr≈1,51. This value is close to the ones obtained from direct surveying observations (tgβ ranging from 1,4 to 1,7).

Table 1. The values of the parameters of the dispersion of influence r and parameters tgβ determined from interferometric elementary troughs Trough S1

S2

S3 S4

S5

S6

Section

rp [m]

tgβp

P1

895

1,20

P2

625

1,72

P3

860

1,28

P1

535

1,72

P2

590

1,54

P3

510

1,79

P1

410

1,69

P1

470

1,64

P2

595

1,60

P1

510

1,32

P2

560

1,19

P1

440

1,50

P2

505

1,36

r[m]

tgβ

793

1,40

545

1,68

410

1,69

533

1,62

535

1,26

473

1,43


264

Figure 4. The distribution of troughs 5.2. Verification of the forecasts of the area subsidence with Knothe’s theory The forecast of the subsidence of the area is possible to obtain by the calculations based on the theories of influences. Theoretical calculations in LGOM are made with the use of a computer system MODEZ. In this system the calculation is based on the knowledge on the situation of fields and the conditions of the exploitation of deposit, as well as the parameters for the modified for LGOM Knothe’s theory. Correct acceptation of the theory parameters for calculation in a definite region of bed exploitation allows obtaining the subsidence growths similar to the determined by e.g. direct surveying measurements. This allows the acceptation of the forecast of subsidence growths as reflecting real temporary subsidence. For the quantitative and qualitative verification of subsidence troughs obtained by InSAR method—their comparison with temporary troughs obtained by MODEZ system (thus troughs close to reality—figure 5) was made. Despite small values of subsidence growths the compliance of troughs was satisfactory both in quantitative and qualitative terms. This confirms great opportunities given by InSAR method and shows its reliability.


265

Figure 5. The comparison of subsidence troughs: forecasted in MODEZ system and determined by InSAR method 6. OTHER POSSIBILITIES OF THE APPLICATION OF InSAR METHOD 6.1. Claims referred to mining-cause damage In the area of LGOM, and in particular on the border of the direct influence of exploitation and beyond, there are many reports of damage to the construction objects while its connection with mining activities is quite doubtful. In these regions there are usually no surveying points to provide information on real influence of mining exploitation. In case of rejecting the claims as not related to mining a problem is solved in the court. The only possibility to obtain the information on area deformations that would be reliable for the court is to use the pair of radar images covering the time of the manifestation of damage. The accuracy of this determination by InSAR method was confirmed by the survey conducted in LGOM and turned out to be fully satisfactory. The applied so far way of solving judicial problems by expensive expertises and opinions of judicial experts has intensively consumed both time resources of KGHM staff and finances of the company, moreover it has not always been fully objective. 6.2. Monitoring of decantation ponds Carried out so far interferograms for the mining area LGOM allowed applying satellite radar interferometry for the monitoring of big decantation ponds. The problem was considered based on several registered by InSAR method images of two ponds located in this region. One of them is an active pond of post-flotation wastes “Żelazny Most”, while the other—“Gilów” in not active (closed in 1984). This influences their different character and related problems and threats.


266

In case of the pond “Żelazny Most” the most important thing is to control high embankments in the aspect of keeping their stability. Properly verified results of an interferometric image of vertical dislocations within the fill slopes—can—too much extent supplement applied so far surveying measurement techniques. In this area InSAR method can provide information on vertical movements of the dyke, particularly in the periods between one and the other surveying measurement. Pond “Gilów” is located in the place of currently conducted mining exploitation. Thus its periodical control is necessary. For the whole surface of the object InSAR method can be applied. Beside the registered with this method mining influence—additionally vertical movements of the surface can be observed. They are probably connected with the continuing so far compression of deposed residues. Carried out so far analyses and speculations show that the application of InSAR method for temporary monitoring of decantation ponds of LGOM is useful. 7. CONCLUSIONS Carried out in this paper analyses and speculations showed wide range of the applications of satellite radar interferometry method—InSAR in solving problems of the monitoring of negative effects of underground exploitation of the copper ores deposit on the surface of the area and its objects. This is a modern method of satellite remote sensing, extending the possibilities of the monitoring of the influence of copper ore deposit exploitation in the whole mining area. Its application has also important economic aspects. The work has done within the framework of Polish State Committee for Scientific Research (KBN), AGH, number 10.10.150.665. REFERENCES Atlantis: EarthView InSAR version 1.1.0 User’s Guide, Atlantis Scientific Inc., Ontario, 1997. ESA 1995: Satelite Radar in Agriculture. ESA raport SP-1185, (red: Tan-Duc Guyenne) 1–71, Noordwijk. Batkiewicz W., Popiołek E.: Prognozowanie wpływu eksploatacji górniczej na powierzchnię terenu w warunkach LGOM. Prace Komisji Górniczo-Geodezyjnej PAN, Geodezja 14, Krakow 1972. Fanelli A., Santoro M., Vitale A., Murino P., Askne J. 2000: Understanding ERS Coherence Over Urban Areas. ERS-Envisat Symposium, Gothenburg., CD-ROM, ESA SP-461. Fisher P., Perski Z. 2001: Qualitative und quantitative Erfassung von Hangrutschungsphanomenen im Bereich “Hoxberg” mittels radarbasierter Fernerkundungsverfahren—Machbarkeitsstudie im Auftrag der Deutschen Steinkohle AG (DSK), Fachhochshule Trier, Umwelt-Campus, Trier. Lichtenegger J., Raney R.K., Schumman R. 1993: Radar Imagery: Theory and Interpretation. Lecture Notes, FAO Remote Sensing Centre Series. No. 67, 103, Rome, pp. 103. Massonnet D., Feigl K.L.: Radar Interferometry and its Application to the Changes in the Earth’s Surface. Reviews of Geophysics, Vol. 36, No. 4, 1998. Perski Z.: Applicability of ERS-1 and ERS-2 InSAR for Land Subsidence Monitoring in the Silesian Coal Mining Region, Poland. International Archives of Photogrametry and Remote Sensing, Vol. 32, No. 7, 1998.


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Perski Z.: ERS InSAR Data for Geological Interpretation of Mining Subsidence in Upper Silesian Coal Basin in Poland. Second International Workshop on ERS SAR Interferometry FRINGE ‘99, http://www.esrin.esa.it/fringe99/ (30 March 2000), 1999. Piwowarski W., Krawczyk A.: Koncepcja Geoprzestrzennego Systemu Informacji o Terenie Górniczym. Mat. V Konferencji – “Dni Miernictwa Górniczego i Ochrony Terenów Górniczych”, Szczyrk 1999. Popiołek E.: Ochrona terenów górniczych. Skrypt AGH, nr 1172. Krakow 1989. Popiołek E. et al.: Analiza rozwoju wielkopowierzchniowej niecki obniżeniowej terenu na obszarze LGOM wywołanej odwodnieniem warstw trzeciorzędowych i czwartorzędowych. AGH, Krakow 1997. Praca niepublikowana. Popiołek E. et al.: Zastosowanie interferometrii radarowej do wielkopowierzchniowych pomiarów odkształceń pionowych terenów górniczych LGOM w celu ograniczania klasycznych pomiarów deformacji—Etap I. AGH, Krakow 2002. Praca niepublikowana. Popiołek E., Hejmanowski R., Krawczyk A., Perski Z.: Application of Satellite Radar Interferometry to the Examination of the Areas of Mining Exploitation. Anwendung der Radarinterferometrie für die Untersuchung von Bergbauregionen. Surface Mining: Braunkohle & Other Minerals 2002. Vol. 54. No. 1, pp. 74–82. Pratti C., Rocca F., Monti Guarnieri A., Pasquali P. 1994: Report on ERS-1 SAR Interferometric Techniques and Applications. ESA Report 10179/93/YT/I/SC 1–122. Frascati. Reigber C., Xia Y., Kaufmann H., Massmann F.H., Timmen L., Bodechtel J. and Frei M. 1996: Impact of Precise Orbits on SAR Interferometry. ESA FRINGE ‘96 Workshop. Strozzi T., Tosi L., Wegmüller U., Galgaro A.: Monitoring Land Subsidence in the Euganean Geothermal Basin with Differential SAR Interferometry. Second International Workshop on ERS SAR Interferometry FRINGE ‘99, http://www.esrin.esa.it/fringe99/ (30 March 2000), 1999. Solaas G., Gatelli F., Campbell G. 1996: Initial Testing of ERS Tandem Data Quality for InSAR Applications. ESA RS/ED96.D002/1.0, 1–54, Frascati. Ulaby F.T., Moore R., Fung: 1981. Microwave Remote Sensing, Active and Passive, Vol. I, Microwawe Remote Sensing Fundamentals and Radiometry. Artech House Inc. Norwood, Norwood, 1–456. Wagner S.: Porównanie pogórniczych obniżeń terenu z obniżeniami określonymi metodą interferometrii radarowej InSAR w rejonie LGOM. Praca dyplomowa niepublikowana, AGH Krakow 2002.

The Control of Mining Damage in China

Yu Xueyi & Zhao Binchao Xi’an University of Science & Technology. China International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: This article contains an analysis of the types, scale and situation of engineering methods of control of mining damage in the coal industry of China. Furthermore, some solutions for preventing and controlling mining damage of natural resources, environment, traffic system and buildings are presented. The authors also propose a method of co-coordinating the development in coal industry and environmental protection using both mining and engineering methods of control. KEYWORDS: resource development, mining damage, environmental protection, land reclamation 1. INTRODUCTION The structure of energy resources will drastically change in the 21st century, and the proportion of coal in the structure of energy resources will decrease. However, owing to the fact that reserves of petrol and natural gas are limited in China, the leading position of coal should not change radically in a short time. With the growth of the population and the lasting development of the economy, the total output of energy will increase at high speed in China during a long period of time. Therefore, although the proportion of coal consumption will be decreasing in the whole of energy resources in the 21st century, the total output still can increase. According to some official forecasts, the total output of coal will increase to 1340 mln tons in China in 2010. Vast the coal reserves have been developed since the 1970’s, and this tendency has brought a series of potentially negative effects at the same time. Due to especially the damage of the environment and resources, this problem cannot be ignored. In recent years, people have realized problems of the environment in coal industry still exist. In order to understand these problems, there is a great need to continue studying the regulations, managements and theory of technology. The government of China is interested in reconstructing the northwest regions of the country, especially the traffic system, resources, and ecosystem. However, the coal


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industry development is one of the important foundations in the development of these regions. Thus mining damage cannot be ignored. Control of the environment from mining damage with the use of engineering methods appears to be the only way to foster development of the coal industry and economy. In the traditional meaning, the mining damage is not limited to buildings, railroads and water systems. It also includes the immovables of the industry and agriculture, traffic systems, landscape, forest vegetation, land resources, water resources, atmosphere and potential mining damage (e.g. mine goaf and cavity of subsidence and destruction). 2. THE MINING DAMAGE AND THE ANALYSIS OF ENGINEERING CONTROL METHODS 2.1. The mining damage of buildings The mining damage of buildings constitutes a considerable part of mining damage. At present, the total quantity of the coal reserves located under buildings has been estimated at 8760 bln tons in China (Li Fengming 1999). Among that, total quantity of the coal reserves under villages constitutes 60%, amounting to 5221 bln tons, and total quantity of the reserves under buildings amounts to 65% of the total quantity of coal under villages. Total quantity of the coal in Hebei, Henan, Shandong, Jiangsu and Anhui provinces constitutes more than 55%. These provinces are the fertile agricultural regions. For example (Huang Leting 1999) in the mining region of Kailuan there are 63 villages, the total quantity of the coal under them reaches 550 bln tons. The safety of the villages has seriously limited the output and development of the mines, especially in some old traditional mining centres in China. The coal seams below the villages in China are exploited using the method of continuous resettlement of people and partial extraction. At present, 75% of the coal located under villages is extracted by moving the settlements elsewhere. Unfortunately, the cost of resettlement has risen considerably in recent years. For example, in the Yanzhou mining area the change was from 28,000 RMB/family to 120,000 RMB/family in 2002. The extraction rate when partial extraction is utilized is about 50%. Not only the cost of mining is high, but also it takes a large resources loss as the price. With mining depth increasing, ineffectiveness of such practice has become more and more been obvious. The use of partial extraction, if the pillars are not left in a responsible way, may potentially cause subsidence on the surface. Such incident took place in the Benxi mining area. However, in order to protect the cities, backfilling method was used there in the 1980’s. Unfortunately, due to the problems with backfill materials, water resources and mining cost, backfilling method is seldom used when mining under the villages. In recent years, in the mining areas with high level of ground-water table (for example: Yanzhou, Datun, Kanluan-tangshan and Fushun), the separated strata fill technology has been used in order to reduce subsidence caused by partial extraction and it has proven itself to be effective (Zhao Jingche and the China…1997), (Lu Lian 1997). The separated strata fill technology is applied to the middle and supreme part of the overlying strata (0, 5 H-0,7 H), and the rate of the surface subsidence reaches up to 50%– 70%. For example, the rate of the surface subsidence of the 1403 working faces in

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Yanzhou-dongtan along the strike reaches 56%-82%, and along dip direction reaches 42%–82%. The Xuzhuang mine in Datun Coal-Electricity Corporation uses the separated strata fill technology in 7215 working faces, the rate of the surface subsidence reaches 51%–73%. The Hua Feng mine uses the coal preparation plant to be floated with tailing water as annotating the thick liquid material—this reduces surface subsidence from 2310 mm to 1483 mm, the rate of subsidence is 36%, maximum subsidence speed at 2,86 mm/d. This enabled extracting coal under the village. The separated strata fill technology not only allows filling a separated strata area, but also increases swelling of rock, which regenerates and reinforces the overlying strata. As the separated strata fill technology uses tailings and mine wastewater at the same time, this method makes the waste material useful and being utilized in the mine. Longwall extraction has been successfully utilized in the Yanzhou mining area, and has achieved best application effect especially as a method used when mining under villages. The depth of mining of huge partial extraction is more than 500 m, partial extraction can increase working face length and make it up to the general length of 120– 150 m of working face in ordinary condition. This basic method enables full-seam mining and reduces deformation. Huge area synchronised extraction is one of the methods extensively used in the world in the early stage of mining exploitation under buildings. In China it is only used if extraction takes place under railroads or single buildings. Utilization of huge area synchronised extraction is a dynamic problem that includes deformation of the overlying strata and surface displacement during mining, the major influence factor that involves mining sequence, time, direction and range of mining, etc. Its application—using theory of the dynamic forecast and full basin evaluation method as the supplementary means— can definitely optimize the mining sequence. This kind of method is especially effective for villages concentrated in a relatively small area or scattered over a relatively big area. The application of this method depends on different building types, building distribution, mining geology and topography change. This method is used for protection of important buildings or architectural complexes. As for scattered villages or buildings of minor importance, relocation of settlements or control after mining is used in order to achieve the best technological and economic effect. Xi’an University of science and technology has developed Ylh-12’s (Yu Xueyi, Liu Chunguang 1995) software that can intelligently estimate, evaluate and resolve the problem of surface subsidence. 2.2. Mining damage to important motorways and railroads Many important motorways built in recent years pass through the areas suffering from mine goaf. For instance, there are five motorways in the Shanxi region that run through many mine goafs. Without suitable control, it would lead to great traffic problems even though the motorways were built using special technologies that protect them against mining damage. Therefore, nowadays in China a lot of attention is given to the studies of the influence of mine goaf on motorways of the advanced grade and the methods of controlling it. For example, there is a pioneer study of controlling structures situated under the mine goaf in Shanxi Taiyuan to Hebei Jiuguan (Sun Zhongdi 2000). At its base is a complex integral theory of probability, evaluating the application of the method in


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two-dimensional integral transformation and building the model simulating structural deformation due to mining damage. The technology embraces makes use of localised high-pressure build up in mine goaf and selecting proper reinforcement of the foundation using thick liquids. The experiment has achieved a good technological and economical effect. Furthermore, the method and theory has achieved the international advanced standard. At present, successful forecasting and evaluating of mine goaf influence have been popularised and applied in the case of important motorways in Shanxi, Henan, Shandong and Xinjiang. However, there are some limitations. Namely, this method is inadequate for controlling the deformations caused by mine goaf in the overlying strata. 2.3. Landslides and mining damage Mining damage often causes landslide disasters. In some mountain mining areas of such regions as Shaanxi, Sichuan and Guizhou etc, landslides are found everywhere due to mining damage. Not only the forests and farmland are damaged, but mining damage also leads to the occurrence of mud-rock flow disasters, which seriously destroys the environment, and threatens the lives and property of local people and institutions. Surface cracks that are caused by mining not only damage building installations and lead to traffic problems, but also together with erosion caused by rainwater destroy huge pieces of farmland soil by mud sliding from overlaying slopes. Applying the proper mining technology can effectively control the hazards of landslips in such regions. For example, the Xiaobaoding mine in Panzhihua mining area achieved outstanding success. The mine applied monitoring systems under the buildings endangered by landslip (Yu Xueyi 1996). The complex control of lateral movement deformation of surface in the slump area and applying safety pillars solves the problem of mining exploitation of coal seams in these complicated conditions. 2.4. The mining damage of land resources Surface subsidence and cracks caused by mining activity are the most common problems for environment and farmland in mining areas. For example, 100 years of mining in the city of Tangshan has formed the area of mine goaf more than 60 km2 big. The areas suffering from surface subsidence caused by mining amount to more than 17,500 ha. The farmers try desperately to yield crops from only 3,800 ha. The subsidence created more than 50 water holes in the area of 1,150 ha. There are 75 villages relocated, and more than 60 thousand peasants lost their land in Tangshan. The situation is very difficult. What is more problematic, it is getting worse—every year the area of damaged surface increases with the speed of 75 km2 in China. Furthermore, mining damage coincides with erosion, especially in the regions with the huge thick loess overburden. Huge surface cracks caused by mining are eroded by rainwater. A wet pitfall loess crack can form great subsidence that creates deep holes or ditches. This is a serious problem in Shaanxi Weihei, Shanxi and Henan, where many mines utilize caving method at present. The level of destruction caused by cracks of this kind has been aggravated and seriously reduces the output of farmlands, more than 50% on average.


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Another problem is waste and tailings amounting to more than 35,000 bln tons in the world. As people do not find the problem of waste and tailings important, these useless materials have been stored up year by year. The quantity of solid waste of mining origin has reached 11,150 bln tons in China, which covers an area of 8,57×103 km2, among them farmland 6,7×103 km2 in 2000. Only one of the large-scale coalmine waste in the Shanxi covers an area of 1364,5 a. What is worse, this amount increases with astonishing speed. Not only mine discarded waste and tailings occupy large pieces of lands. Also smoke and dust have become a serious source of pollution. There are 237 coal waste dumps in danger of occurrence of spontaneous combustion in China. Therefore, the Chinese mining industry must introduce strict control of a large quantity of harmful gases. Integrated processing of mining waste and tailings seems to be a perfect solution here. 2.5. The mining damage to water resources Surface subsidence caused by mining has not only destroyed buildings and cultivated land, but also destroyed landscape and ecology of river systems. This causes enormous problems in irrigation. Most of mining area of Shanxi suffers from this problem. For example, the silt caused by erosion in the Yellow River goes into in Wulanmulun River every year (e.g. last year the increase was up to 239 mln tons from only 1, 7 mln tons). The environment problem worsens with each passing day. Therefore, water resource protection is an important component of this area environment ecology protection. In addition mine water pumped out from underground and poisonous substances from mine waste dumps penetrate into soil, which can cause surface water pollution. System planning and implementation of mining measures and comprehensive regulations on land reclamation in mining areas should be stressed nowadays. 3. THE WAY OF CONTROLLING MINING DAMAGE Since the 1950’s the western advanced mining countries have been paying a lot of attention to comprehensive regulations in mining damage. The Chinese government announced (environmental protection law) in 1979 as a comprehensive set of laws regulating the work to be done in case of subsidence damage caused by mining activity. Reclamation of land and planned cultivation has achieved excellent results in some modern mining areas. According to the Chinese specialists, the mining industry should follow a principle of “controlled hazard-comprehensive utilization” in order to reduce mining damage. 3.1. Controlling mining technology in order to reduce mining damage The application of controlling mining technology is a major method to reduce mining damage. The bearing ability of the environment should fully be taken under consideration when thinking about the environment and ecology. In my opinion, the industry should popularise and apply separated strata fill technology in difficult conditions. There are also


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some other possibilities, such as huge partial extraction; backfilling, limiting mining width or huge area synchronised extraction. These methods have been utilized with great success with huge economic benefits in experiments carried out not only in China. But the aspect of product system needs further theoretical studies. The following aspects should be given a lot attention: non-linearity theory on the foundation of key layer of establishment, dynamic estimating theory about the overlying strata in mining. This analysis should also include such factors as mining rate, direction, sequence, overlying strata lithology and mining depth, etc. The relations between trend changes and the change of speed in surface displacement and deformation should also be considered. 3.2. Comprehensive control ecology Before constructing new coal mines, ecology should play crucial role in planning. Furthermore, environmental protection should be taken under consideration in the first place in mining areas where specialists plan future exploitation. Utilization of systems of controlling mining damage and environmental protection is of crucial importance here. There are some simple solutions that may be applied in this respect: – using coal waste to generate electricity (utilization of coal waste amounted to 66 kt in China in 2001, it should reach 80 kt in 2005; the share of coal waste in generating electrical energy should increase from 43% to 60%), – using waste rock and mine tailings as backfill material, – using waste rock, tailings and backfill to reduce surface subsidence in order to cultivate again in the damaged farmland, – using waste water to fill mine goaf and applying suitable methods of controlling water division layers.

4. CONCLUSION Comprehensive control of the mining damage should be a long-term action. Unfortunately, the considerable cost for the national economy is the main reason why the government has to postpone many projects for the future. Even though there are some projects or successful experiments, there is a big difficulty in implementation. This implementation must, of course, be reinforced by a strict system of regulations. To sum up, a very telling proverb may be quoted “the one who destroys, should be the one who controls, and everyone will benefit”. REFERENCES Li Fengming 1999: Research Present Situation Mining Coal under the China Village and Follows the Problem Development Trend. Coal Science & Technology. Vol.: The Present Situation Mining Coal under the 27 (1) 1999. Huang Leting 1999: China Villages is with Developing the Focal Point and Measures in the Mine and the Coal Resources Can Go On to Mine the Strategy Building. (4) 1999. Zhao Jingche and the China Mining Industry University Publishing House. Publishes 1997.


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Lu Lian 1997: Wins and Mines Coal and Cultivates Present Situation and His Development Utilization Countermeasure Centre Nation University Learned Journal. Vol.: With Subsiding Again in the Tangshan City. 6 (1) 1997. Yu Xueyi, Liu Chunguang 1995: “Coal Mining under Building, Railroads and Water Body” are Mined and are Protected Estimating Evaluation Software with Ground, and is Economized in the Mining Office in Shaanxi, 1995. Sun Zhongdi 2000: Ect & Poineering under the Advanced Grade Highway in Goaf. Estimating and Appraising and Reaching Research Reports & Science Publishing 2000. Yu Xueyi 1996: The Technology of Mining under the Building in Xiao Baoding. The Scientific Research Committee in Kapok City 1996. Tumle S. & Sisson R. 1999: Closure and Reclamation Plan’s for a Proposed Expansion Project. World Mining Wall Bulletin. Vol.: 15, (6) 1999.

Limiting Inflow of Water to Operating Shafts by Application of Permanent HydroInsulating Screens

Roman Kuś G.Janik & R.Kuś PRGW Sp. jawna. Sławków, Poland International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: This paper presents the manner of limitation of underground water inflow to the shaft, on the example of works designed and performed in the mine “Zofiówka”. It presents the principles related to the design of such works, and afterwards, the performance of the same in the conditions of continuing operation of the shaft and it touches upon the necessity of the application of the screen in loose soil. KEYWORDS: Shafts, hydro-insulating solution, plugging

1. INTRODUCTION Przedsiębiorstwo Robót Geologiczno-Wiertniczych (The Geological and Drilling Company) in Sławków has developed and implemented the author’s technology of loose soil plugging by means of injection (grouting) of a binder, the fundamental components of which are modified polymineral clays. So far, in practice, the following works have been carried out: – Packing of the ground and earth embankments of water reservoirs. – Repairs and rectification of failures in hydro-engineering facilities (soil compacting and building of hydro-insulating screens). – Compacting of subsoil for residential buildings in the mining damage area. – Neutralisation of flooded faulting zones in mines by means of a clayey cement solution. – Stopping the inflow of salt water to mining excavations. – Quenching of thermoactive mine waste dumps by injection of a modified hydroinsulating solution.

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Properties of clay-based hydro-insulating solutions are determined by the composition and amount of individual components. They are made by the combination of a clayey mixture with binding agents. The choice of proper binders depends on the chemical composition of the underground/ground water. Clay-based hydro-insulating solutions, when set, take on a viscoplastic consistence. Properties such as: plasticity, stability, resistance to washout, resistance to corrosion and low filtration ratio are related to specific properties of clayey rocks (polymineral clays). The task of a filtrating screen may be: – Insulation of water-bearing strata around a structure, e.g. a shaft. – Reduction of the water filtration ratio to the value 1×10−8 and 1×10–11 m/s. – Improvement of geotechnical conditions and creation of “normal working” conditions for a hydro-engineering structure, building or shaft.

2. METHOD OF STOPPING OF UNDERGROUND WATER INFLOW TO EXISTING MINING SHAFTS In the underground mining technology, the basic role is that of shafts, the flow capacity of which determines the effectiveness of the work of a mining plant, since this is the way for: taking the excavated ore out to the surface, delivering machines and equipment and taking the miners up and down the mine, as well as ventilation of mining excavations. One of the main factors making exploitation of the shaft considerably complicated is inflow of underground water, particularly if these are aggressive waters. Water inflow exceeding permissible levels adversely affects the lining and reinforcement of the shaft and its engineering use, forming unacceptable hygienic and sanitary working conditions. The shafts should be in good technical condition during the whole period of operation of a mine, thus, it is inevitable to successively eliminate effects of aggressive water activity, which is usually connected with considerable financial expenditure. To ensure a good technical condition of shafts during its sinking and life, hydroinsulating by means of clay-based hydro-insulating materials is widely used. 2.1. Materials used for shaft hydro-insulation The sealing material used, depending on specific geological and hydrogeological conditions as well as mineralization (aggressiveness) of underground waters, may be: – Cement grouts. – Polymineral clay-based hydro-insulating solutions. – Cement-silicate solutions. – Clayey and clayey-silicate solutions. – Organic polymer-based solutions (bituminous, latex, synthetic resins). The essence of hydro-insulating screens using clay-based materials consists in the making of watertight screens around the shaft in places where underground water flows


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in, by means of injecting of a packing solution through a hole drilled according to a properly selected technology. 2.2. Sealing methods Hydro-insulating works are normally carried out through holes drilled in the surface, shafts or mining excavations existing in the direct neighbourhood of the shaft barrel. Plugging of flooded rocks through holes drilled from the shaft is generally performed downwards, including upwards plugging of individual aquifers (water-bearing layers). Lengths of shaft sections being plugged are assumed to be equal to thickness of the aquifers intersected by the shaft. The aquifer is normally plugged in a few sections when its thickness is bigger than 20 m. Depending on the purpose and the flow chart, holes are distributed in alternating rows (chequered arrangement) or in a spiral. The length of injection holes depends on their purpose. These may be 0, 5–1, 0 m long holes for insulation of space behind shaft lining only (between the lining and the rock-mass), or up to 10 m and longer holes, when fractured rock-mass outside the unstressed rock zone adjacent to the shaft lining is to be plugged. Depending on the orientation of fracturing, the holes may be drilled horizontally or diagonally. Horizontal holes are drilled in case of steep and vertical fractures, whereas diagonal holes are drilled when fractures are slightly sloping and horizontal, in cases of plugging of top parts of the rock-mass around the shaft and existing favourable technical drilling conditions. Depending on the thickness of the aquifers, the horizontal distance between the holes is from 1–3 m, and the vertical distance is from 1–4 m. Despite wide application of hydro-insulating screens from inside of the shaft in practice, these have a number of disadvantages, mainly: – Impossibility to use the shaft for mining purposes during plugging works. – Necessity to erect a special platform in the shaft for performance of plugging. – Necessity of drilling of a considerable number of holes due to the impossibility of increasing of the distribution radius of plugging solutions through the use of higher injection pressures. – Possible damage to the shaft lining and reinforcement during plugging. – Difficult conditions of drilling equipment operation and performance of plugging with a high inflow of water to the shaft. If the shaft is being hydro-insulated through holes drilled from the surface, vertical holes or holes slightly sloping towards the shaft lining are drilled, forming a “band” screen around the shaft outline. When the technical conditions are favourable, 6 to 12 injection holes are drilled, their positions being marked out on the circle around the shaft outline, with spacing not less than 4–5 m and maximum not more than 10–12 m from the shaft barrel outline. The number of injection holes is proportional to the spacing between the drilling posts and the outer edge of the shaft. The bigger the distance from the shaft perimeter, the bigger the planned number of boreholes, which is caused by the necessity of interpenetration of injection range zones of individual holes. The biggest spacings between injection holes are assumed for eliminating water inflow from water-bearing strata of low hydrostatic pressure and a medium or high degree of rock-mass fracturing, and the smallest—for high hydrostatic pressure of underground water and for capillary fractures. Plugging from


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the surface usually aims at eliminating water inflow to the shaft under construction. For existing shafts, injection holes are drilled around the shaft, whereas, depending on the designed structure of the holes and plugging technology used, the solution is injected by means of special heads mounted in the outlet or by means of packers.

Figure 1. Schematic view at process of area injection Extremely important are adequately precise location of the hole in relation to the shaft and injection of the hydro-insulating solution at a pressure safe to the shaft lining. In practice, in view of difficulties concerning location of drilling equipment on the surface within the shaft top in operation, conditions allowing the application of a hydroinsulating screen through holes drilled from the surface are rare. The main disadvantages of the technology of aquifer plugging through holes drilled from the surface are: – It is possible only for small depths of deposition of water-bearing strata (200–300 m), due to inevitable, considerable deflection of very long holes that has an adverse effect on the quality of hydro-insulation. – Formation of filtration “windows” in the rock-mass being plugged due to failed interpenetration of screens formed around individual holes. – As a rule, it is not possible to drill the optimum number of holes with because of the buildings and mechanical equipment situated near the shaft, land development and existing transport roads.


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– Necessary drilling of a considerable number of expensive injection holes. Technologies of hydro-insulating screening of existing shafts do not usually take into account the presence of an intensely fractured rock-mass zone around the shaft, formed during shaft sinking by drilling and blasting. This way, when the hydro-insulating solution is being injected, it may leak into the shaft, which requires technological breaks in injection operations for the period of solution stabilisation (not less than 8 hours), causing unjustified additional consumption of the packing material, and it also has a negative effect on the quality of the hydro-insulating screen. Depending on the hydro-insulating screening technology, the units designed for preparation and injection of the packing solution may be on the surface or in the shaft— on a special platform. The solution may be injected through a single hole or through a few holes simultaneously. Whereas, depending on the designed structure of the hole and plugging technology, the solution may be injected through a special head in the outlet or by means of packers. 3. DESCRIPTION OF CONDITIONS IN THE REGION OF SHAFT U-1 KWK “ZOFIOWKA” – Shaft U-1 is the main skip shaft, at the same time performing the role of a downcast shaft. – The depth of the shaft is 982,76 m (−702, 61 m below sea level). – The internal diameter of the shaft is 7, 20 m. – Lithological structure of Quaternary formations.

Table 1.1 Stratigraphy

Depth

Quaternary

0,00–4,00

Scatterable yellow sand dust

4,00–9,20

Fine- and medium-grained sand and locally gravel

9,20–12,0

Grey sand dust

12,0–19,5

Fine- and medium-grained flooded sand

19,5–21,2

Plastic grey clay

21,2–29,5

Fine-grained flooded sand (quicksand)

29,5–34,0

Plastic grey clay

34,0–41,0

Fine-grained flooded sand

41,0–49,0

Mine-run and flooded gravels

49,0–50,0

Plastic grey clay

50,0–50,80

Varigrained sand

below 50,8

Grey loam

Tertiary

Designation of strata


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Type and technical condition of the lining. Up to the depth of 50, 0 m there is a 90, 5 cm lining of clinker brick class “350”, on cement mortar. From the depth of 50, 0 m to 830,1 m there is a monolithic lining made of concrete class “400”. Shaft U-1 up to the depth of 295 m was sunk in a frozen rock-mass. In the course of sinking, at the depth of 132, 9 m the floor had heaved by 1,6 m. On 10.09.1963, at the depth of 134, 5 m the initial lining of the shaft was damaged at the section 85,0–133,0 m. To restrain further side pressure on the lining, the shaft was backfilled up to the depth of 88,4 m. At the same time, there was a discharge of water from the defrosting rock-mass amounting to 1,8 m3/min. Hydrogeological conditions. Following the sinking of the shaft and putting it into operation, the rock-mass was defrosted and water has been discharging from behind the shaft ginging at the section of 15, 0–50, 8 m until now. Water discharging from behind the ginging comes from the Quaternary water-bearing level (flooded varigrained sand). This water is collected by water rings and drained by means of a pipeline Ø 150 mm into a fire control tank situated nearby the shaft—580 m level. The volume of inflow to the shaft from the section in question amounted to 1,0 m3/min in 1967, ca. 0, 34 m3/min in 1996, and in 2000 it increased to 0, 56 m3/min. It is clean water with a low content of suspended matter and a low degree of mineralization that, according to PN-80/B-01800 standard (concrete and reinforced concrete structures. classification and designation of media), shows no aggressiveness to concrete. Hydrogeological parameters of the Quaternary aquifer. The thickness of permeable deposits nearby the shaft is 19,1 m. The aquifer is built of a top part developed into fine- and medium-grained sand and a bottom part built of varigrained sand with gravel interlayers. These two layers are separated by a ca. 14, 5 m thick clay book. The finding from observation of the shaft lining was that water exudates on the lining in shafts 1 and 3 occur in two zones, starting at the depth of 15 to 30 m and from 36 to 49 m (at the section where permeable formations occur). The highest discharge at the main lower level occurs at the depth of 36 m (+244, 5 m above sea level) and the main discharge—at the depth of 49 m.

Table 1.2. Breakdown of hydrogeological parameters of the aquifer in shafts 1, 2, 3, 4, 5 Specification Test well for shafts I-II

Thickness of the aquifer [m]

Filtration ratio (k) (m/s)

23

4,6·10−5

Shaft 1

23,3

Shaft 2

22,6

Shaft 3

16,5

Shaft 4

16,8

Shaft 5

14,0

H115

21,5

6,7·10−4


H120

16,5

II/1

17,5

Average

19,1

282 3,4·10−4 3,52·10−4

Development of inflow in time. Water inflow to the shaft is stable and depends on the condition of the lining. In 1997 the lining of shaft 2 was sealed, whereas an inflow of water to shaft U-1 has been observed to increase slowly up to 0, 56 m3/min.

Table 1.3 Date

Shaft U-1

01 Jan. 90

0,30

01 Jan. 91

0,26

01 Jan. 92

0,30

01 Jan. 93

0,24

01 Jan. 94

0,32

01 Jan. 95

0,36

01 Jan. 96

0,34

01 Jan. 97

0,40

01 Jan. 98

0,40

01 Jan. 99

0,36

01 Jan. 00

0,49

01 Jul. 00

0,56

6 Sep. 00

0,54

4. TECHNOLOGY OF STOPPING WATER DISCHARGE FROM BEHIND THE LINING OF THE SHAFT U-1 KWK “ZOFIÓWKA” AT THE SECTION FROM 12,0 M TO 51,0 M Plugging operations comprise: – drilling, – preparation and injection of the hydro-insulating solution, – inspection of the quality of packing.


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4.1. Drilling 12 injection holes were designed. Injection and test-injection holes were drilled successively according to numbering as in figure 2. The holes were drilled alternately from both sides of the shaft. Compacting holes performed the function of test and injection holes (wells). The holes were solid drilled with the use of a drilling liquid made from the solution of base clay. The location of the holes was determined by the presence of shaft equipment, ancillary shaft facilities, documented and verified in the course of work and by the necessity to preserve the maximum permissible injection pressures. 4.1.1. Injection holes drilling technology The location of the holes as well as directional parameters were changed in the course of plugging, as information on hydrogeological and lithological conditions was provided, in connection with hydrodynamic observations carried out in the course of plugging. Drilling was carried out according to the following pattern: – Determination of the place of drilling by land surveying methods according to the project, corrected in relation to the model of hydrodynamic conditions in the region of the shaft updated during performance of the works. – Assembly of the drilling unit and setting of the direction of drilling. – Drilling of the hole for the casing pipe, fitting and grouting of the casing pipe (using a hydro-insulating solution). – Technological break of 16–24 hours for hardening of the insulation solution. – Expanding of the grouting pin and inspection of the quality of hydro-insulation. – Drilling of a hole up to the projected depth of the zone to be plugged. – Packing of the water-bearing (aquiferous) zone. – Technological break for stabilization of the hydro-insulating solution in the rock-mass. – Expanding of the zone being plugged, hydrodynamic observations in the borehole. – Drilling down to the next planned level determined for plugging. – Packing of the water-bearing (aquiferous) zone. – Technological break for stabilization of the hydro-insulating solution in the rock-mass. – Expanding of the zone being plugged, hydrodynamic observations in the hole. – Continuation of successive plugging cycles until the projected target depth of 51 m is reached. Drilling and plugging were carried out concurrently in two holes situated alternately in relation to the shaft. In the outlet of plugging holes, a valve limiting the volume of injection pressure was installed. An increase in pressure above the projected one blocked out the valve and washed the hydro-insulating solution outside the installation. For drilling of injection holes drilling units MDR-06e and DeltaBase 520/140 were used.


284

Figure 2. Layout of borehole placement around the U-1 shaft 4.1.2. Preparation and transport of the hydro-insulating solution The base clay solution was prepared in the direct neighbourhood of the shaft U-1. The area of preparation and injection of hydro-insulating solution was indicated in agreement with the Shaft Department of the Mine. The hydro-insulating solution was fed into injection holes by means of a high-pressure hose, using an NB-3 pump with capacity control. The injection capacity was ca. 0,02 m3/min. 5. RESULTS In the course of drilling, very difficult conditions were come across due to the presence of reinforced concrete and steel elements in the ground as well as a very small amount of compacted rubble under the shaft curb at the depth of 3–12 m, manifested in uncontrolled sinking of 5–15 cm long sections of the drilling pipe. Difficult drilling conditions and the necessity to change the location of holes in connection with the disclosure of a number of technical facilities in the ground (basements, pipelines, power cables) not revealed before, made it necessary to increase the number of injection holes up to 19.


285

Also in the zones of planned injection, at the depths of 12–19,5 m, 21–29,5 m and 34– 51 m, a very low degree of compactness of soil was noted, the result of which was a rockmass absorbing capacity (1390 m3) considerably bigger than the planned one (510 m3). Drilling and plugging did not collide with the operation of the shaft. As a result, all works were completed within 3 months, while the lead-time estimated by the mine was 11 months. The works aimed at limiting the volume of noxious to the mine water inflow. Thanks to the utilization of plugging methodology developed by PRGW Sławków, the inflow of water was considerably limited—from 520 dm3/min to 80 dm3/min. Such a low inflow had never been recorded before. Apart from the technical effect, a favourable economic effect was achieved since limitation of water inflow to the shaft generated annual savings of ca. PLN 300 thousand for the mine, only by virtue of saving of energy necessary for pumping out of water onto the surface. Thus, in a short time, all expenditure related to this investment was quickly returned. Experience gained in the course of plugging of the shaft U-1 KWK “Zofiówka” allows us to state that these works may be even more effective when the contractor is given more freedom with regard to modification of the previously developed project. In the design phase it is impossible to predict all technical and geological restrictions affecting location of holes, and thus, the volume of hydro-insulating solution necessary for hydro-insulating screening.

Problems and Prospects of Development of the Coal Mining Enterprises in the Donetsk Region

A.R.Vovchenko, V.G.Grinyov The Regional state administration. Donetsk, Ukraine D.M.Zhitlyonok State enterprise “Dzerzhinskugol”. Donetsk, Ukraine International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: The crisis in the coal mining industry of Ukraine will increase unless the reforms continue to be implemented with the aim of its privatisation. The problems of the coal mining industry are discussed in the context of Ukraine joining WTO and integration into EU. The productive potential of the coal mining enterprises of the Donetsk region is represented at present by 77 mines and 4 mine administrations with the total productive capacity 48,4 million tons a year including 35 mines producing steam coal. Long working life of most mines without reconstruction, technical re-equipment and financial misbalance in the industry has resulted in the situation when only 10 (including one producing steam coal) of them produce more than 1 million tons a year, 59—up to 300 thousand tons and 51 are at different stages of closure. Most of the coal mines are in difficult financial situation because the average cost of producing 1 ton of marketable coal is more than 1,4 times the coal realization price. Moreover in 2003 the cost increased by 10% and the price increased only by 3% in comparison to the corresponding period of 2002. Such misbalance causes acute shortage of capital financial assets. A simple analysis of financial & economic activity of the coal mining enterprises shows that even if all the coal products were to be sold and paid for in full by the customers, the deficit would still make up 15% of the amount necessary to balance the sheets. The debt of the coal mining enterprises’ creditors equals to 4, 5 billion grivnyas and is nearly 3 times bigger that the debt of the industry itself. The arrears of wages continue to increase. As on January 1, 2003 the arrears of wages made up 4 months’ fund and they have increased by 6% in the first 10 months of this year.

Problems and prospects of development of the coal mining enterprises

287

The negative financial balance of coal products production caused a great number of mines to become unprofitable. One fifth of all the mines in the region are subordinate to the Ministry and receive no state subsidy for covering cost losses. In 2002 privately owned mines produced 64% of the total coal output in the region and the rest of the mines (i.e. 80% or 68 technical units) produced only 36% of the output. The existence of such polarization of the technical capacities and great differences in financial and economic indicators of coal production are a clear proof of the necessity of adjusting the mines’ funds. In 2003 the state-aided mines received on the average 22 copecks for 1 grivnya of the final product to cover for the incurred losses. However, the cost of production of 1 grivnya worth of product on these mines in 2002 varied from 1,1 to 3,6 grivnyas, and in 2003 this indicator on the mines in Snezhnoye town, for example, reached 6 grivnyas. Certainly the losses can’t be covered by the state subsidy in these conditions. But what could be the criterion of determining which of the mines are unpromising? It is necessary to define the formula of self- sufficiency taking into account the internal investments and the budgetary and investment resources. The programme of reducing the budget burden must be clearly drawn for the nearest 3–4 years. The programme of closing down the unpromising coal mines in Ukraine has been in place since 1996. In 1996 there were 135 mines in the region, now there are 77. The idea of the programme is to enable the operating mines to make up for the reduction in coal output, which resulted from closing some mines down. But in real life this has not happened. During the decade capital investments volumes in the coal mining industry have reduced 12 times, coal output during this period has decreased 1,4 times and the decreasing tendency has been continuing. The steam coal output has dropped from 17,1 to 10,8 million tons in the last three years, and now the coking coal output is decreasing. For the first 10 months of 2003 this output decrease was 2,2 million tons in comparison to the previous year. The present day situation in the coal mining industry confirms an earlier conclusion that increasing state subsidy does not result in improving economic situation at the coal mines. Absence of the mechanism ensuring objective and independent distribution of the state financial support does not stimulate the mines into increasing the overall performance and decreasing unprofitability. The present day situation in the coal mining industry requires cardinal changes. Much was promised in this regard by the programme “The Ukrainian Coal”, which was approved by the Cabinet of Ministers of Ukraine in 2001 with only one reservation—to take into account the actual budget potential to bear the programme realization costs, while at the same time making out the drafts of the state programme of economic and social development of Ukraine. As the approval of the Programme was dragging out in 2001 the deadlines for fulfilling the immediate tasks on the organizational-legal level turned out to be unrealistic and the tasks unfeasible. The actual state budget for the industry for the year 2002 was approved in the volume of 50% of that necessary for the Programme realization. Under these conditions the basic points of the Programme were corrected. But as the result of that the Programme lost its purpose and its realization was brought to formalities. The mines’ financial misbalance aggravates their economic situation, causes social tension in the collectives, decreases the prestige of the miners’ labour, leads to exodus of


288

workers and reduction in coal output. In the year 2003 alone the number of workers in the coal mining industry decreased by 5%. On the other hand the number of working pensioners is on the increase and it already stands at nearly 36%. The mechanism of solving the problems in order ‘to normalize the coal mining industry has been known for a long time—it is the necessity to restructure the finances, which means reorientation of the state subsidy resources. The state subsidies for covering losses are to decrease gradually with simultaneous increase in financing of the major construction work and technical re-equipment of the enterprises. Along with chronic insufficient financing, the coal mining undergoes different kinds of reorganizations more often than any other branch of industry. This is not incidental. Those trying to analyse work of the coal mining industry constantly come across the low efficiency of managerial control and usage of the available resources. The present administrative structure of the coal mining industry took shape in 1996–1997 and its adaptation to conditions of the market economy is not yet complete. In the current year the structural transformations in the coal mining industry of Ukraine have started, but the set of legal instruments ensuring that the reform of the branch’s administrative structure would be carried out, was not prepared. The interaction of the newly established state enterprises with some organizational bodies appeared to be beyond the legal field and the main purpose of the reform— ensuring the receipts and expenditures balance in the coal mining enterprises—has not been achieved. As aresult of this coal output is continuing to fall. The scientists at the Institute of the industrial economy of NASU (city of Donetsk, Ukraine) processed the data of technical and economic indicators of the coal mining industry in the Donetsk region for the period between 1990 and 2002 using the probabilistic forecasting technique on the basis of the auto-regression model. The results of the model were unfavourable. The total output in 2004 could be 37,3 million tons, in 2005–35,3 and in 2010 it could fall to 26 million tons. Such results are logically explained by the fact that statistics with its indirect methods takes into account shortage of the branch budgeting, lack of attention to major construction work, wear and tear of the basic industrial assets, operating obsolete and worn equipment, labour turnover and so on. A study of these problems inherent to the process of mine closure shows that closing down of these coal enterprises that form settlements, without foreseeing the social consequences and without undertaking rehabilitation measures and developing the devastated areas, leads to social degradation of the settlements. The establishments of the social sphere collapse, miners’ settlements are left without telephone & medical services and are on the decline, and their residents migrate to other regions for lack of jobs. The Ministry of coal industry considers these problems to be of national importance, not limited only to the affected branch of industry. On the basis of the above we can come to a conclusion: in spite of the measures taken, all the crisis phenomena in the coal industry failed to be overcome and they tend to aggravate. We can also say with confidence that neither the previous way of managing the mines nor the current one nor even the one offered by the new reform, can eliminate the reasons behind the negative processes in the industry which have always existed even in the period of production growth. It is obvious that the state is not an effective owner.

Problems and prospects of development of the coal mining enterprises

289

The enterprises, which are least dependent on the Ministry, do not experience such problems. Privately owned enterprises already produce about a third of the total coal output. These enterprises’ executives concentrate their effort on research, development and applying state-of-the-art technological solutions and organizational and economic measures. As an example of effectiveness of the model of getting the final product in the system “mine—preparation plant” the experience of an open joint-stock company “Coal Company” mine “Krasnoarmeiskaya—Zapadnaya № 1” may be presented. The fact that it was able to attract investments from private investors proves that the Company’s economic indicators are higher than those of the best mines in Russia and Germany working in similar mining and geological conditions. Without a doubt the most favourable scenario of the branch development is such where the mines are managed and run by their owners or users of the coal products who are the connecting links in the coal-coke-metal or coal-electric power chain. Without this it is difficult to speak about any transformations or the positive dynamics of the processes of the coal industry development. Negative characteristic of the situation in the coal industry is aggravated by its problems in the context of Ukraine joining WTO and integration with the EU. It is necessary to increase the stimulating function of the state subsidy, which is now mainly of distributing character, by regulating the state subsidy of the coal industry with special laws passed on the basis of thorough study and analysis of WTO norms. Integration of Ukraine with the EU will also require undertaking radical measures in order to increase competitiveness of the coal industry at home by accelerating its restructuring, technical modernization and reorganization of finances. It is necessary to increase budgeting of the branch substantially for some period of time so that it could function normally on the eve of joining the EU and then to decrease its financial dependence on the state budget. As far as financial reorganization goes, preserving and developing the coal mining industry of Ukraine would require a law regulating the coal industry’s functioning in the conditions of free market economy, and providing the coal-mining enterprises with special rights and powers. It’s necessary to create an administration system of the coal mining enterprises, which would clearly define the role of the state structures in the industry’s restructuring, the laws concerning the coal mining companies and the procedures of appointing administrative bodies of the coal company. Yet another problem, for a long time on hold, must be solved. It concerns reevaluation of the available reserves of operating coal-mining enterprises of Ukraine. The objective and realistic evaluation of geological coal reserves adjusted for the market economy requirements, can radically change our idea about the role coal can still play as a fuel and source of energy. REFERENCES Amosha A.I.: The Problems of the Home Coal Industry in the Context of Ukraine Joining WTO and Integration in EU/Materials of the International Conference “Coal of CIS”, Yalta, 19–20 September 2003.


290

Conceptual Fundamentals of Scientific and Technical Development of the Region./Editorial board: A.I.Amosha, A.M.Bliznyuk, M.G.Chumachenko, V.P.Shevchenko, V.F.Yanukovich/NAS of Ukraine. The Institute Of Industrial Economy. Donetsk 2002. P. 172. Grinyov V.G., Vovchenko A.R.: The Coal Industry of the Donetsk Region on the Way to Market Relations. Economy of Ukraine. № 8, 1999. Investment Attraction to the Coal Industry of Ukraine: Situation, Problems and the Ways of Solving. Monograph/L.V.Baisarov, M.A.Ilyashov, A.V.Korzun, V.I.Logvinenko, S.V.Yanko/— Kiev. The base, 2002. P. 288. Vovchenko A.R., Grinyov V.G.: Economic Aspects of the Coal Industry Restructuring on the Way to Market Relations. Coal of Ukraine. № 10, 1999.

Organizational—Economic Models of Investment Activity at the State-Owned Coal Mining Enterprises of Ukraine

V.I.Logvinenko, O.Yu.Kuzmich Donetsk regional state administration. Donetsk, Ukraine International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6 ABSTRACT: The main results and prospects of the investment activity in the coal industry have been considered. At present the coal industry of Ukraine is represented mainly by state enterprises condition of production and economic activity of which is extremely difficult. As one of the main reasons of such situation, great deficit of funds both internal and budgetary aimed at support of coal industry, should be mentioned. The way out of this state of affairs may be attracting private capital to the state coal mining enterprises. Investments can be done in various forms. Analyzing the prospects of investment activity in different branches of industry it should be noted that the coal industry position is notable for its low attractiveness to private capital. This is connected to a number of factors of legislative, macroeconomic, ecological and social nature. The development of investment activity in many industries is restrained by absence of a normative control of these problems and lack of proper legal protection for investors. Uncertainty as to the return on the invested capital has considerably reduced the inflow of private capital into the coal industry in the past. The coal mining enterprises are distinguished by their considerable capital intensity and low profitability and most them are unprofitable. At a result the time of return on the invested capital is much longer than 5 years and many investment projects result in a loss. This negative situation is aggravated by a number of factors of purely technological character. They are connected to low probability of accurately predicting production outputs from coal seams developed and mined in conditions of constant worsening of geological conditions. Taking into account the situation that has taken shape, a complex intensified programme of attracting capital, including private investments has been under way for


292

the coal mining enterprises of the Donetsk region—the main coal region of Ukraine, during a number of years. The task boiled down to three major problems: 1. Developing organizational chain of interaction between the investor and the enterprise. 2. Adjusting business planning methods taking into account specific features of the mining technology. 3. Providing additional measures for protecting investors’ interests. Major role in achieving the set task was played by the legal act: “About special economic zones and special mode of investment activity in the Donetsk region”, which was passed in 1998 and which granted investors considerable tax privileges. Enjoying these privileges gave the opportunity to increase the effectiveness of economic activity of the enterprise and thus to enhance the certainty of return on the invested capital. 19 investment projects with the total investment amount of more than 1,1 billion grivnyas were developed in the coal mining industry in recent time under the abovementioned Act. This is about half of all the capital invested in the coal mining enterprises of the Donetsk region during a corresponding period. 9 of the projects have been already completed, the rest are at different stages of realization. Until the present day, implementation of the project created 3263 new jobs and preserved 2582 in the existing working places. About 5,5 million tons of run-of-mine coal has been produced out of which more than 3,5 million tons of final coal products worth more than 600 million grivnyas have been received. Realization of investment projects in preferential tax conditions gave coal mining enterprises an opportunity to make available more than 85 million grivnyas, which were reinvested in production activity. At the same time the enterprises paid more than 135 million grivnyas in taxes to budgetary and non-budgetary funds. New projects are mainly focused on introducing modern mining technologies and state-of-the-art mining equipment. As a result their economic efficiency is considerably higher than that of traditional production methods. It is sufficient to note that on certain projects the output from production faces has been increased more than 2,2 times and the cost of production of 1 ton of runof mine coal has reduced by 5–8%. All the investment projects may be divided into two groups. In the first one the realization of investments is done by the coal mining enterprise itself on the basis of internal or loaned funds. At present this is the most typical case for enterprises of the state-owned sector. The idea of this model is that a section, which becomes a target of investment, is artificially separated within an operating coal mining enterprise to form an economical and organizational structure, but does not become a new business entity. As a rule such structures are mining sections having separate financial and economic accounting. Such approach gives the opportunity to present such sections in their best condition for estimating economic efficiency of the investment project and its adaptation to the standards of the Ukrainian Act “About special economic zones and special conditions of investment activity in the Donetsk region”. The coal products mined in these sections have preferential tax treatment. Based on the above-described model “Krasnolimanskaya” mine has prepared and now is introducing 9 projects with the total investment capital of more than 700 million grivnyas. At this enterprise more than 1800 new working places have been created and

Organizational—Economic models of investment activity

293

more than 1845 existing working places have been preserved, there are no budget debts and no arrears of wages. An average wage of essential jobs is 1650 grivnyas. Funds to the amount of 49 million grivnyas, released due to the preferential tax treatment, have been fully reinvested in equipping a new mining block. 5 projects are currently under way at “Krasnolimanskaya—Zapadnaya № 1” mine. At this enterprise a maximum stope output of 3200 t/d has been achieved in similar conditions. A distinguishing feature of investment activity of “Komsomolets Donbassa” mine is utilization of capital loaned from its foreign partner—Slovak company KIMEX in a form of BMV—14 (P3) mechanized longwall set. The project has been successfully completed with considerable economic efficiency. In second group of investment models the activity of a mine is combined with production activity of a private investor within the minefield of the coal mining enterprise, but the investor forms an independent and separate business enterprise. They contract on rendering production services (transporting run-of-mine coal, ventilation, leasing of the surface infrastructure etc.). Two variants of this model of investment activity have been developed. The first one envisages the investor’s coal mining activity in stopes within the scope stipulated in the mine license. With this the investor uses his own material and labour. The main object for investment is powered equipment of the stopes. As an illustration of realization may serve an investment project which is being successfully realized at the mine “Zhdanovskaya”. The investor is OSS “Financial industrial group “Nedra”. The second variant of the model envisages the investor’s production activity after getting license for the part of the minefield allotted for developing by the investment project. The license gives the investor the right to own the coal mined. This in turn guarantees return on the invested capital. Economic independence is one of the main conditions of increasing the attractiveness of coal mining if production infrastructure of operating mines is to be utilized. The project introduced by the mine “Rassvet” can serve as an example of realization of this model. Special attention should be paid to the investment project of SS “Ugolinvest”, envisaging construction of a new mine in the town of Snezhnoye creating 242 new working places. The above mentioned investment projects of the second group are quite clear examples of the effectiveness of attracting private investments into production activity of coal mining enterprises. Based on their experience a mechanism of privatization of the coal industry of Ukraine can be worked out. They convincingly prove that even under the existing legislative and economic conditions investing in coal mining activity may be attractive. As to the methods of calculating business plans of investment projects, characteristic features of mining production must be taken into account. The first one has to do with the necessity of conducting mining work, face extraction in particular, continuously. This is caused by high rock stress manifesting itself at large depths. So, provision for sufficient material, labour and financial resources must be made at the planning stage in order to meet the planned production targets. Any delays and disruptions in sequence can lead to essential economic losses. The second feature is that the place of performing the main technological operations of mining changes. It brings about low probability of accurately


294

forecasting production levels and thus economic and financial indicators of coal mining enterprise activity. That’s why the method should introduce into the analytical model of a business plan some additional adjusting parameters, taking into account the error of forecasting of its basic indicators: period of realization, total capital required, time of return on the capital and other indicators of economic efficiency. Production activities of coal mining enterprise can vary. The choice of a process sheet of reserves developing, scheduling, face cleaning, using different types of mining equipment and other technological processes predetermine the degree of economic efficiency of the investment in the long run. The criterion of choice of the final model variant is the maximum volume of discounted money, earned by the enterprise for a minimum project realization period. That’s why the essential condition for qualitative substantiation of a scenario is optimization of basic parameters of mining production. As the experience shows it is necessary to consider more than 30–50 different variants of production and economic indicators, forecasting development of the mining enterprise, in order to have the opportunity to choose the most expedient for the given conditions. Speaking about the prospects of investment activity it must be admitted that the main direction of its development must be to increase consolidation of productive potential of the investor-enterprise and coal mining enterprise. Such a conclusion is the result of an elementary analysis of the way finance-production groups are formed in the world practice. Development of relations between different market players in coal industry of Ukraine confirms the basis laws of these processes. At the initial stages there was only a financial interaction between the investor with coal mining enterprises, in a form of credits and leasing. But this practice did not provide an investor with satisfactory assurance of recovery of the invested resources and participation in on-line control of the object of investment and has become outdated now. That’s why at the next stage, a system allowing investor’s participation in making expedient from his point of view administrative decisions, including distribution of financial resources made available in result of business activity, found a widespread application. This does not apply to the enterprise as a whole, but only those faces that attract investment resources. With such an approach interests of the investor and coal mining enterprise are equally taken into account and responsibilities for activity of the object of investment are distributed in the same way. This system of interaction is most popular at present. But at the same time evolution of economic transformations in Ukraine requires further deepening of the integration processes between the investors and production enterprises. Typical examples of realization of the third stage are the above-mentioned investment projects based on joint production activity. Production activity of the investor on the basis of the existing infrastructure of the mine is beneficial not only for him in the form of a greater probability of recovery of the invested resources and higher degree of the production process control. It is also economically sound for the operating mine. Even in condition of its unprofitability, rendering production services to the investor enables the mine to get additional resources, compensating partially for the losses. Apart from this the investor has the opportunity to test its production capability and on this base to adjust its strategic plans.

Organizational—Economic models of investment activity

295

This way is the last but one borderline before privatization of the coal mining enterprises when the investor turns from a partner into an owner. Such a transformation cardinally changes the mine’s management views concerning future destiny of the enterprise forcing them to make long-term plans. The approaches to production and economic activity planning change from aiming at getting a quick profit to making the whole enterprise activity effective for a long term. The above analysis presents key stages of the investment activity development in the coal industry. Now, in order to enhance its efficiency, essential adjustment of the legislative base and providing corresponding programs of attracting investments are required from the government.

Author Index

International Mining Forum 2004, Kicki & Sobczyk (eds) © 2004 Taylor & Francis Group, London, ISBN 90 5809 607 6

Bachowski, C. 225 Balykhin, G.A. 71 Binchao, Zhao 235 Bondarenko, V.I. 33, 43 Bruniany, C. 21 Buchwald, P. 169 Butra, J. 131 Carmona, J.M. 105 Carvajal, A.R. 105 Chekushina, T.V. 71, 119 Cybulski, K. 179 Dychkovsky, R.O. 33 Dubiński, J. 197 Fernández, C.G. 105 Gajos, S. 157 Gatnar, K. 83 Gayko, G. 65 Gladush, A.D. 71 Grinyov, V.G. 249 Horoszczak, L. 147 Ivkovic, M. 47 Jakubów, A. 83 Jaskólski, Z. 169

Author index Kaledina, N.O. 189 Karginov, K.G. 119 Konopko, W. 197 Kovalevska, I.A. 43 Krawczyk, A. 225 Kuś, R. 213, 241 Kuzmenko, O.M. 33 Kuzmich, O.Yu. 253 Lamot, T. 157 Lebecki, K. 179 Litvinsky, G.G. 11 Ljubojev, M. 47 Logvinenko, V.I. 253 Oraee, K. 57 Paraszczak, J. 1 Perendic, S. 47 Popiołek, E. 225 Porotnikov, V.V. 43 Pourkhandani, R. 57 Simanovich, G.A. 43 Slastounov, S.V. 99 Sopata, P. 225 Stjepanovic, M. 47 Szlązak, N. 83 Szulik, A. 179 Tor, A. 83 Urbaś, M. 157 Vorobyov, A.E. 71, 119 Vovchenko, A.R. 249 Xueyi Yu 235 Zhitlyonok, D.M. 249 Ziętkowski, L. 147

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