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By J. Oliver, J. Cleason

"With maintenance costs rising and an increasing demand for higher production, the use of sealed spherical roller bearings can reduce maintenance costs while providing increased production uptime. The sealed spherical bearing has increased bearing life by slowing the ingress of contamination. This paper will describe how contamination and bearing wear is reduced by reviewing case histories from coal and metal mines that improved their production by decreasing downtime using sealed spherical roller bearings. INTRODUCTION The spherical roller bearing (SRB) is a robust rolling element bearing that can accommodate high radial loads, low to moderate axial loads, and misalignment with no loss in carrying capacity. (i.e. Figure 1) Because of these characteristics, it is a favorite in mining applications and many other heavy industry applications. Primary applications are the various pulleys on conveyors. Coal and hard rock mines have hundreds of miles of conveyor lines and countless head, tail, drive, and idler pulleys. Maintaining the bearings on these pulleys and replacing bearings that have failed can be a huge expense to a mine in components, labor and downtime. Ideally, the bearings are replaced when the pulley is changed out according to the maintenance cycle, about every 4 years. Reality, however, causes the bearings to be changed out every year, or sooner. These premature bearing replacements can cause production shut downs for as long as 8 hours. Conveyor pulleys have been targeted at many mines as the single biggest maintenance pain point.The primary reason for replacing conveyor bearings is internal wear from fine, solid particle and water ingress through the bearing housing seals. (i.e. Figures 2 & 3) Traditionally, the bearing is mounted in a split pillow block housing (SAF, FSAF, SDAF or SNL are typical industry styles) with only labyrinth seals (LER/LOR) for protection against the ingress of contaminants. The bearings are lubricated with grease. A common way to extend the MTBR is to purge the bearing system with large quantities of grease to flush out the contaminants. If the bearing, and thus the housing seals are not frequently purged with grease, fine solid contaminants enter the housing and bearing through the labyrinth cavities."

Jan 1, 2015

By R. J. McLaughlin

Acid mine drainage (AMD) can occur as a natural phenomenon through oxidation of sulfide minerals. Oxidation by air and water can be exacerbated by mining. Because AMD can exhibit low pH, it is capable of extracting heavy metals from mineral deposits. The result is highly acidic drainage with elevated heavy metal concentrations -the combination of which can cause extensive environmental damage. Because of the importance of these factors, this paper is focused on sulfate based AMD. Due to widespread mining, the occurrence of AMD is ubiquitous. As a result, several approaches to treatment have been developed. Chemical treatment using alkali is most commonly employed to treat AMD. There are alternative methods of treatment, as well as alternative ways of handling treatment sludges or concentrates. Dewatering of sludge or concentrates is beyond the scope of this paper.

Jan 1, 1996

By J. E. Murphy, Y. Topkaya, R. E. Johnson, B. R. Palmer, D. K. Dysinger, R. B. Bhappu, A. E. Clark

The U.S. Bureau of Mines utilized a central composite experimental design to investigate leaching kinetics of pure chalcopyrite in the sulfuric acid-oxygen system. The six factors studied were oxygen pressures of 150, 300 and 450 PSI, leach temperatures of 35O, 65O, and 95O C, oxygen-to-chalcopyrite mass ratios of 0.5, 1.0 and 1.5, leach times of 8, 16 and 24 hours, initial acid strengths of 0.6, 1.2 and 1.8 g H,SO,/g chalcopyrite and pyrite to chalcopyrite mass ratios of 0, 1, and 2. Chemical analysis and oxygen consumption measurements indicatedthat leaching occurs through three consecutive reactions. Temperature was the primary determinant of leach rate and solution acidity was the primary determinant of oxygen consumption, sulfur oxidation, and activation energy.

Jan 1, 1993

By L. P. Johnson

Longwall mining has gained acceptance as a safe method of maintaining coal production where conventional systems were unsuccessful or uneconomic due to bad roof conditions (Kuti8). One of the major problems confronting American operators involves the correct choice of equipment which should be selected to suit the local mining and geologic conditions for optimum productivity. To achieve this, factors such as rock strengths, local geology, panel geometry, and panel orientation must be examined to optimize production and minimize down time due to ground control problems. Little quantitative information is available in the literature regarding the effects of geologic conditions on American longwall operations, and past implementations of longwall systems appear to have been based largely on experience. When mining engineers are able to assess the geologic conditions both quantitatively and qualitatively, they can more fully optimize equipment selection and panel design.

Jan 1, 1977

By C. Bailey Williams

Aragonite "sand," an oolitic lime-stone, is distributed throughout the Bahamas on the Grand Bahama Bank, adjacent to Southeast Florida. Bahamian deposits hold 75-100 billion tons of recoverable Aragonite. Aragonite's precipitation from seawater over 700 years ago assures low contaminates, making it a pure material for such industries as cement, glass, etc. A 90-acre, man-made production island was built south of Bimini, adjacent to ocean shipping channels mining is by cutterhead, suction dredge, dewatering, and stockpiling for further drainage and shipping. Aragonite is transferred to the U. S. East and Gulf Coasts via bulk carriers to points of sale.

Jan 1, 1978

By Rodrigo Peroni, Euler Victoria

Knowing the depth of contacts between weathering zones in phosphate mines is determinant to define the stripping ratios and the mineralized zones. Contacts are gradational and difficult to define even in the drillhole samples, so the combination with geostatistical simulation techniques helps to define the contact zone and the grade distribution, providing a way to assess the probability of the occurrence and the positioning of contacts between the barren waste and the ore zones. These inputs are used in the scheduling, anticipating risks associated with mine planning scenarios. The results from an application to a case study showed a good improvement when compared to traditional techniques.

Jun 25, 2023

By A. Almeida, D. Castro, K. Ingmarsson, D. Ferreira, R. Tamir

Mine-to-mill improvements can be initiated at any stage, but it is our assertion that drilling accuracy is a practical anchor in any operating mine or quarry for a structured continuous improvement process. This paper will: 1. Define deviations from plan and provide some best practices for maintaining drilling accuracy. 2. Introduce tools and methods to monitor and calculate consequences of accurate versus inaccurate drilling. 3. Demonstrate the use of these tools and methods and their impact on one cost item, the mitigation of oversized material by a hydraulic hammer. 4. Provide recommendations for implementing improvements.

Feb 1, 2023

By W. de Bruijn

Differential reduction and expansion or shrinkage of hematite under the influence of intimately associated foreign oxides may be a source of stress formation and sinter degradation during blast furnace reduction. To study specific effects, synthetically pre-pared hematite powder evenly doped with 1% by weight of several elements present as oxides, was pressed into tablets and sintered at high temperature prior to reduction. Continuous thermo-gravimetric and dilatometric analyses were per-formed on the sintered tablets as these were being subjected to reduction under simulated blast furnace time/temperature/gas composition conditions up to 1183 K. These findings were coupled with observations on the microstructure of tablets removed from various stages of the reduction process.

Jan 1, 1988

By . A. Drouin, J. -M. Lalancette, D. Lemieux, C. Chouinard

"A gold bearing pyrite concentrate was treated at the demonstration scale (15 t/d) for the extraction of gold using a cyanide-free process. The pyrite was depressed in the flotation circuit in order to promote the flotation of copper and to achieve a sufficient copper grade to meet the copper smelter criteria. However, the recoveries of copper were 80% and only about 50% for the gold was found in the copper concentrate. The balance of the gold (50%) and copper (20%) contained in the pyrite concentrate as a reject from an overall flotation circuit. The gold grade in the pyrite concentrate is 5.40 g Au/t and copper about 0.77%. About 170 tons of pyrite concentrate were processed in the DST demonstration plant. The concentrate was first oxidized using a fluid bed to produce a calcine. Copper was then extracted as copper sulphate using diluted sulphuric acid. The residual solid was submitted to the DST chlorination process for gold extraction. Thereafter, the gold was recovered from the pregnant brine by precipitation over silica using DST proprietary process. The maximum gold recovery by chlorination was 90%. The average gold recovery by cyanidation of these calcines was 71%. INTRODUCTION Dundee Sustainable Technologies (“DST”) has developed a chlorination process as an alternative to the conventional cyanidation process for the extraction of precious metals from ores or concentrates. The technology offers viable and efficient gold recoveries and the ability to handle refractory ores and contaminated material (As, Hg, etc..). The process was successfully demonstrated on hundreds of samples provided by various mines, exploration and development projects and has evolved from laboratory experiment to demonstration at an industrial scale. DST’s chlorination process operates in a closed loop circuit, which implies that no liquid effluents are generated. The chemistry and process design allows precious metal deposits containing base metals to be treated viably. The tailings from the process are inert and non-toxic, sulphide depleted and thus non-acid generating. The efficiency of the process has demonstrated excellent gold recoveries within a fraction of the time needed for cyanidation to obtain similar results. The company completed in 2015 the construction of the first industrial demonstration plant operating with the DST technology. The plant has a capacity of 15 tons per day of mineral concentrate, and the ability to extract gold and silver utilizing DST chlorination technology. DST has tested its chlorination technology at demonstration scale on the refractory pyrite concentrate containing 5.40 g Au/t and 0.8 %copper. This auriferous pyrite concentrate, was by-product of copper concentrate respecting the copper smelter norms, is produced in a country where cyanide is prohibited and refractory to cyanidation. The pyrite concentrate was initially processed successfully (gold recoveries: > 90%) by the DST technology at laboratory (2011: 20 kg) and pilot (2014: 5 tons) scales to extract gold. About 170 tons of this pyrite concentrate were processed in the DST demonstration plant located in Thetford Mines."

Jan 1, 2017

By S. Tandanand

The Bureau of Mines, U.S. Department of the Interior, assessed the geological effects on the ratio of maximum subsidence and extraction thickness, hown as the subsidence factor, to develop a simple method for predicting the maximum subsidence. The study is restricted to the Northern Appalachian Coal basin where data were collected from 13 coal mines. A model previously developed to relate the subsidence factor with the lithology and with the width and depth of working panels was used in the assessment. The result shows the possibility of determining maximum subsidence without resorting to a rigorous method.

Jan 1, 1982

By M. Nambiar

Due to the remote locations of the mining projects and the demand for long transmission lines to provide electrical power, mining sites often operate with poor network power quality and low short circuit capacity. In order to cope up with this situation, more and more mining projects are considering the installation of appropriately designed Synchronous Condensers(SC) together with Power Factor Correction and Harmonic Filter System(PFC&HFS). The installation of these power quality equipments is viewed as a technological advancement in the mining application. An adequate level of short circuit power is essential and has to be ensured for the safe and reliable operation of especially large cycloconverter driven gearless mill drive systems used for grinding application. Increase in plant short circuit power is generally achieved by lowering the impedance of the supplying network. The SC also provides smooth voltage control during mill starting or major load trips, thereby regulating the plant voltage within the required limits. This paper will explore the impact, technical advantage and reasoning for the selection of the advanced hybrid technology mentioned above, in order to cope with the power system challenges for future mining projects. It also highlights the potential problems and how to address them carefully.

Feb 23, 2014

By R. M. Damodaran

The flotation performance of Florida phosphate rock was evaluated in terms of the mixing behavior of reagents with particulates during anionic conditioning of phosphate slurries. The various conditioning variables that influence mixing, namely, solids loading, agitation speed and the conditioning time were investigated to determine their effect on reagent coating on the particles. The phosphate recovery was used to monitor the conditioning efficiency. The mixing characteristics of the phosphate feed were evaluated in terms of the ratio of the mixing power P, to the minimum power Pm' required to keep the slurry in suspension. It was found that mixing conditions govern the conditioning time required to obtain high recovery. Reagentization was slower at lower P /P m ratios. At higher P /P m ratios, the flotation recovery was optimized and the agitation speed was found to have no significant effect on it. Optimum mixing conditions were obtained for a 74 wt% slurry at an agitation speed of 450 (47 rad/s) rpm. Under the optimum conditions of mixing, recovery was improved at low collector dosages and short conditioning times.

Jan 1, 1995

By J. R. Cobble

The Bureau of Mines has developed a process to improve the leaching kinetics of diffusion controlled systems. The system features the Bureau-developed turbomill, an attrition grinder which scrubs the reaction product coating away from the unreacted mineral grains during leaching. The system can be continuously operated at any temperature up to the boiling point of the leaching agent. This paper describes the research on the dissolution of chalcopyrite with acidified ferric sulfate which forms a sulfur product on the unreacted chalcopyrite surface. This method is a hydrometallurgical alternative to chalcopyrite smelter technology which eliminates S)2 emission problems with its associated costs. Batch testing at 90? C with the chalcopyrite/ferric sulfate system resulted in better than 99-pct extraction of copper. At an optimal energy consumption of 1,520 kW h/st, 95 pct of the copper was extracted. Recent work has been directed to development of a continuous system and SX/EW treatment of the resulting pregnant liquor from the chalcopyrite/ferric sulfate leach.

Jan 1, 1991

By M. Rieber

The recent history of the mercury market results almost entirely from U.S. Government actions. Its derived demand has been curtailed by U.S.E.P.A. mandates; its supply has been augmented by U.S.D.L.A. controlled stockpile releases. The resultant prices have altered its market structure, impacted mercury mining, refiners and recyclers, and affected international production and markets. From a marketplace product mercury is rapidly becoming a disposal problem. Other minerals may follow in its wake. This paper reviews developments in demand, particularly in relation to environmental constraints. Supply side impacts are considered. The impact of stockpiles, both as a market overhang and as a potential supply is reviewed. Government stockpile disposal activity is emphasized, but the size, estimated disposal rate, and quality of the private stockpiles are considered, as is the potential availability of additional material from shut-in mines.

Jan 1, 1995

By Timothy J. Matthews, Timothy J. Batchler, Daniel J. McElhinney

The National Institute for Occupational Safety and Health’s (NIOSH) Mine Roof Simulator (MRS) at the Pittsburgh Mining Research Division was commissioned for operation in 1981. The MRS was originally designed for longwall shield research and was the only hydraulic press in the world with the capability to apply biaxial loads to these large structures. The results of this forty-year research program have contributed to improved performance and increased knowledge of longwall shields, mobile roof supports and hydraulic cylinders in the mining industry. Also, considerable research and development have been conducted to develop standing support technologies for the mining industry. The Support Technology Optimization Program (STOP) was developed to evaluate how these various technologies can impact and improve both ground control conditions and mine safety. Additional MRS research has included the performance of ventilation stopping walls including the development of a new testing protocol to evaluate the transverse loading capabilities under arching load conditions and evaluating the performance characteristics of welded wire screen as used in underground mine roof surface control. The NIOSH Mine Roof Simulator has been a reliable and valuable tool for the past forty years for the evaluation of safety structures used in mining, and this research has contributed significantly to the advancement toward safer mines. This paper will summarize those research results and provide a glimpse of planned future activities.

By S. Chatterjee, S. Bandopadhyay

A spatial modeling technique based on the Bayesian neural network (BNN) is proposed. Incorporation of the Bayesian method for posterior probability of the output parameter helps calculate the uncertainty associated with the estimate. Parameters of the BNN model were selected using a grid search algorithm. The expected value and the variance of block support were calculated by Markov chain Monte Carlo sampling of the posterior distribution at discretized points within the block. The BNN model was validated by applying it to the Walker Lake data set and comparing the results with ordinary kriging results. The comparison revealed that the proposed BNN method performs marginally better than ordinary kriging. The variance map is less smooth than with ordinary kriging, and the proportional effect is less in the BNN-based model than in the ordinary kriging model. The developed BNN model was applied to the Coho block of a placer gold deposit in Nome, AK. Results reveal that the BNN model is globally biased for the Coho block and underestimates the sample mean. The grade volume curve was developed by calculating the block mean and block variance after discretizing the block into a number of points and applying the BNN model to each discretized point.

Jan 1, 2011

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Victor S. Sazhin, Alexander Kostenko

The hydrochemical method, which is based on the separation of aluminium oxide and silicon oxide by the hydrothermal decomposition of aluminosilicates by alkaline solutions in the presence of calcium oxide, can be effectively used for the complex processing of high-silicon aluminium ores. In this case, the silica is bound to a slightly soluble compound: Na20.2Ca0.2Si02•H20 (under the conditions: [&k]= 9, CaO/SiO2 = 1.05, alkali concentration in the solution: over 300 g/l Na20, T = 240...300 °C). If alkaline solutions of 100...150 g/l Na20 concentration are used, Ca0•SiO2•H20 or 2Ca0•Si02.0.5H20 compounds, depending on calcium oxide dosage, are formed at 280...320 °C. New methods for the crystallization of Na20.Al203.2.5H2O from concentrated alkaline solutions rand low-concentration solutions are proposed to isolate aluminium from solutions. The hydrochemical method, which involves no sintering process allows one to decrease the expenditure of heat by a factor of two, to reduce the investment and to exclude the pollution of the ambient air by dust-laden gases. The expansion of the raw-material resources base of

Jan 1, 1984

By H. G. Stephenson, D. Castendyk

Open pit anthracite mining at the Canmore Creek Mines, Canmore, Alberta, south of Banff National Park, resulted in several pit lakes by the early 1970’s. To enable future mining opportunities in the Rocky Mountains, operators decided to reclaim the site into a non-polluting landscape which would provide a memorable outdoor experience for residents and visitors alike, and would preserve wildlife habitat. Reclamation began with the stabilization and revegetation of spoil piles adjacent to Canmore Creek, and the backfilling and revegetation of Canmore Creek Mines No. 1 and 2 to create “Salamander Pond.” The company dammed the outlet of an open pit at Canmore Creek Mine No. 3 to raise the water level and create Quarry Lake. Reclamation concluded in 1980 when the Town of Canmore took ownership. Quarry Lake is now the centerpiece of 293-acre Quarry Lake Heritage Park, a popular destination for swimming, hiking, and fishing, visited by about 400 people on summer days. The Alberta Environment Conservation Authority described this as “The best mine reclamation in Western Canada.” Quarry Lake is an exceptional example of walk-away, pit lake closure, where the mining company created a valued post-mining asset and was released from future liabilities. INTRODUCTION One of the most important environmental issues associated with contemporary open pit mining is the closure of mine pit lakes, and the perpetual legacy these lakes can potentially generate. Ideally, successful pit lake closure includes the following: (i) the water quality within the pit lake, plus surface water and groundwater discharging from the pit lake, meets pre-mining, or other acceptable water quality conditions of the receiving environment, (ii) the termination of all active treatment processes, (iii) the return of the pit lake to the government for long-term care and maintenance, (iv) the release of the mining company from all future liabilities associated with the pit lake, (v) the return the environmental bond to the mining company, and (vi) the reuse of the pit lake for some post-closure, beneficial purpose. The achievement of these goals is known as “walk away, pit lake closure.” Peer-reviewed, publications of successful walk away, pit lake closure provide mining companies the ability to address public concerns regarding the safety of pit lakes which commonly occur during the permitting, operations, and closure phases of mining. To date, most published work on walk away, pit lake closure work has occurred in the coal mining sector in Germany and Australia (McCullough et al., 2009; Jones and McCullough, 2011; Kumar et al., 2013; Schultz et al., 2013). However, in North America, examples of walk away, pit lake closure are not widely discussed in the literature despite there being excellent examples from the coal mining sector in Alberta, Canada (e.g. Pleasure Island, Black Diamond, East Pit, Sphynx, Robb, Lovette, Siltstone, and Sterling), and from the iron mining sector in Minnesota in the United States (e.g. Portsmouth Lake, see MNDNR, 2018). Although seldom included in discussions of pit lakes from the metal and coal mining sectors, the aggregate and fabricated-stone mining sectors also produce pit lakes, and there are numerous examples throughout North America where these pit lakes have been successfully closed and transformed into post-mining resources. The difficulty of finding these good examples in the literature makes it challenging for companies to respond to public and regulatory questions regarding pit lake closure and reuse. As such, there is a pressing need document the motivations and approaches used in examples of successful, walk away, pit lake closure in North America so that the mining industry at large can learn from these examples and can better address public concerns.

Jan 1, 2019

By Grażyna Dzik, Adam Mirek, Krzysztof Król

The Upper Silesian Coal Basin is located in the area of the Upper Silesian Syncline and belongs to one of the most complex tectonic units in Europe. It occupies 5,650 km2. In 2020, there were 20 hard coal mines operating in the area. Seismological research in this area consists of observation of precursors/symptoms of shock generation and understanding of their effects. Studies (Teisseyre R. 1972, Mutke G. & Stec K. 1997, Stec K. 2007, Dubiński J. & other 2010, Idziak A. 2010) of complex impacts of rock mass provided a background for determining the conditions for the formation of exploitation and regional seismic shocks. Issues related to tectonic and exploitation stresses, successive extraction of coal seams at higher depths, operation in the vicinity of fault zones, often complex geological and organic conditions result in the activation of a discontinuity surface due to stress caused by deformation of rock layers, which, as a result, causes the emergence of strong shocks above 107 J. The article was based on analyses of extraction value, high-energy shocks, both of mining and tectonic-mining (so-called regional shocks) origin, against the background of changing operating conditions, which, as a result, creates an image of periodically changing seismic activity and location of strong shocks. Observation results show that, with an increase in exploitation depth and declining extraction, the number of strong shocks caused by tectonic stress (107 - 109 J) and exploitation stress (≥105 J) has been steadily increasing over the last 25 years.

Apr 26, 2022