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By Bruce H. Gardner

This paper describes the application procedure of the Stress Control mine design method. This procedure has evolved over the past 20 years of the practice of this Method in trona, potash, salt, and, most recently, in coal mining. The Stress Control Method of mine design has been defined and amply described in previously-published work (1,2,4) -- likewise, the associated system of in situ instrumentation (3,4) and the computer modelling technique (2,3,4) which are the principal tools of this method. The present paper defines the procedure by which the in situ instrumentation and computer modelling cools are combined in adapting the general concepts of Stress Control to produce site-specific design solutions. This procedure is illustrated by means of case examples from three mines. The Stress Control Method utilizes the geometry of underground excavations, particularly through the time sequence of the creation of yielding pillars, to control stresses. The time sequence is devised such that all the primary stress envelopes are transformed into a single secondary stress envelope, surrounding the group of rooms separated by the yielding pillars. The secondary stress envelope provides protection to the roof and floor boundaries from all the external stresses. This stress relief effect removes the cause of roof and floor failure -- high shear stress in the weakly confined boundaries of underground openings. It thereby replaces artificial support with ground self-support, enabling simultaneous in¬creases in stability and percent recovery. Within limits which are still being explored and extended, the Stress Control Method has proven capable of overcoming the trade-off between stability and extraction ratio which has afflicted conventional mining methods. The application procedure of the Stress Control Method is outlined below in terms of the objectives, structure, and stages of a generic mine rock mechanics program for the site-specific adaptation of the time control yield pillar design technique.

Jan 1, 1984

By Gheorghe Oncioiu

The phenomena developed on the influence of thick coal seam no.3 mining, by horizontal slices and rocks caving roof control, in the Jiu Valley coal basin, are very complex. By analyzing the influence volume of underground faces and defining the stress and strain zones, around the faces, it is possible to evaluate the suitable powered support performances. Also, Jiu Valley thick coal seams involve about 50% of reserve balance in the main safety pillars of social, industrial and natural objectives, situated on the mining surface perimeters. For a correct design and drawing of main safety pillars complex research was done with the aim of establishing the subsidence parameters. Upper face rocks medium could be assimilated as a granular medium, which is in a permanent movement. For describing the rocks movement, respectively the ground surface displacement, was taken into account the stochastic mechanics principles. The theoretical solutions are in accordance within situ measurements.

Jan 1, 1999

By Kenneth Rigsby

Black Mountain Resources' Highlands Mine No. I is a full extraction room and pillar operation located in Harlan County, Kentucky. Black Mountain extracts the 0.9 to 1.1 meter (36- to 42-inch) thick Owl Scam that is situated 60 to 70 meters (200 to 235 ft) above the longwall panels of the abandoned Arch No. 37 Mine, in the 3.4 meter (I I ft) thick Harlan Scam. In the areas where full extraction has taken place, the overburden above the Owl Seam ranges from 60 to 430 meters (200 to 1,400 ft). A short room and pillar panel was developed above an extracted longwall panel to test the selected pillar size, to evaluate support systems, and to verify the shape of the subsidence profile over the recovery end of the depleted longwall panel. It is over the recovery end that Highlands Mine No. I would access the Owl Seam reserves that exist above the No. 37 Mine. Black Mountain selected pillars sizes to be employed above the longwalls based on the ability to fully extract a pillar employing extended cuts. After a mechanical testing program was completed and mine behavior and top disturbance were observed in the test panel, John T. Boyd Company (BOYD) verified pillar stability and designed support systems. BOYD previously designed No. 37 Mine's gate pillars and studied subsidence and gob formation over its longwall panels; this effort aided the analysis of overburden movement and pillar integrity at Highlands Mine No. 1.

Jan 1, 2003

By Wolfgang Schott

The propagation velocity of in-seam seismic (ISS ) waves depends on frequency which is also known as dispersion. This dispersion is not only determined by the thickness of the coal seam and its dirt hand content (i.e. thickness and position), but also by the rock type and its condition (soft or compact) above and below the coal. A coal panel completely surrounded by roadways was investigated using the in-seam seismic transmission survey technique. Along the roadways the immediate roof was either sandstone and shale- The seam thickness varied between 1.20 m (sandstone in the roof) and 1.40 m (shale in the roof) and a fault system (normal faults) crossed the panel with a total throw of about the seam thickness. Below this coal seam another coal scam with a thickness of about 80 cm exists. Its interval shown by exploratory drilling changes from about 60 cm to more than 6 m (scam splitting). Tomographic inversion was applied to the dispersion of the measured ISS waves and velocity distributions were calculated at frequencies found to he most sensitive to changes in the seam thickness and the distance to the coal seam below. By correlating the velocity distributions with known values along the roadways approximate distributions of the coal seam thickness and the seam splitting within the panel were produced. The mining of the coal panel had already been finished and comparison was made between the survey results and the actual situation as far as it was documented. Prediction by the seismic survey was partly in good agreement with the actual held conditions or showed at least the same trends.

Jan 1, 2003

By Bruce K. Hebblewhite

Tower Colliery is a longwall mine operated by BHP Coal Illawarra Collieries, Southwest of Sydney, Australia It mines the Bulli Seam at a depth of approximately 450m. The surface topography overlying the mine consists of several steep-sided river gorges, up to 68m in depth, which run at oblique angles across a sloping terrain Apart from some light rural development, the surface land is essentially natural bushland, but is traversed by a major freeway. This crosses one of the gorges on twin, six-span, box-girder bridges, with bridge piers up to 55m in height. Consequently, a major surface subsidence monitoring program has been in place for several years now, including intensive conventional, GPS and E DM surveying, plus real-time monitoring of critical components of the bridge structure. Although the bridge and freeway are outside the conventional angle of draw' subsidence influence criteria, and have seen only negligible vertical deformation as a result of mining, there has been widespread evidence of regional horizontal deformation of the land surface, large distances away from the mining area The effect at the bridge has been well within acceptable tolerances, primarily because of the regional nature of the movements, with very little differential movement observed at the bridge piers Gorge closure and evidence of large headlands of land moving `en masse' have been observed. Horizontal movements of hundreds of millimetres have been recorded in some locations where gorge closure has been monitored. Possible explanations of these movements include one or a combination of mechanisms such as pre-mining stress relaxation. valley notch effects, valley bulging, regional joint patterns, movement toward active goaf areas, vertical gorge shearing' and shear failure of horizontal bedding planes below the surface This paper presents a case study of the regional horizontal ?subsidence displacements? measured during mining. and a discussion of the possible explanations for these phenomena

Jan 1, 2000

By Kevin Jinrong Ma

Underground mines often experience roof falls in entries, crosscuts, and intersections of active mining sections, main travel ways, and belt entries. Roof fall heights greater than 20 ft (6 m) make re-bolting of the newly exposed roof dangerous and impractical. To protect mine personnel, belts, moving vehicles, and other facilities, mine operators typically install steel structures such as square or arch sets in the roof fall areas. Wood lagging is usually installed between the steel-sets to enclose the area and protect the entry from recurring falls. Usually the void space above the steel-sets and laggings is backfilled. However, backfilling high roof fall voids is costly, which causes some operators to leave the voids open. In this case, durability of wood over time and the capability of the steel-set and wood lagging to resist falling rocks are unknown. During the last few years, Keystone Mining Services, LLC (Jennmar Corporation affiliate engineering company) designed, tested, and developed an impact-resistant lagging system to protect steel-sets1. Various steel-sets protected with the impact-resistant lagging were approved by MSHA and installed in different underground roof fall rehabilitation projects. This paper presents (1) design, development, and laboratory testing of the patent-pending impact-resistant lagging panel, (2) steel-set design methodology according to the American Institute of Steel Construction (AISC) national standard, and (3) impact-resistant steel-set design method based on Law of Conservation of Energy. Two underground cases are reviewed to demonstrate the successful application of the impact-resistant lagging system.

Jan 1, 2011

By Mikko Ahola

This paper describes recent innovative analyses conducted by the Bureau of Mines to demonstrate the application of the boundary-element method in evaluating the effectiveness of caving chambers or sacrificial entries to relieve the stress conditions initiating cutter roof failure in underground coal mines. The analysis assumes that cutter roof failure results when the shear strength of the roof rock at an entry corner is exceeded by the applied shear stress. The applied shear stress is due to large differences between the horizontal and vertical in situ stresses known to exist in many coal mining districts experiencing cutter roof failure. The results of boundary-element analysis suggest that driving an arched lead entry in advance of adjacent rectangular entries and allowing it to cave significantly reduces the difference between the principal stresses at the corner of the rectangular entry, thereby lowering the maximum shear stress. Slippage of the immediate roof along horizontal bedding planes into the caved entry is shown to further reduce the stress conditions in adjacent entries. Practical applications of this stress control technique to longwall and room-and-pillar mining are discussed.

Jan 1, 1987

By Craig Byington

The stress-field orientation mapping and analysis (SOMA) technique for determining the operative stress field near mine workings and its relationship to various fracture sets is described using Dickenson-Russell Coal Company's Laurel Mountain mine as an example. Carried to its practical application, this technique ultimately defines the optimal orientation for the mine workings wherein pillars, rather than rock bolts or other roof support methods, dominantly shoulder support of the highest-probability, potential, roof-fall blocks. It also provides a predictive tool describing local variability in rock deformation and in secondary roof support methods. Fracture discontinuities within a rock mass, either as pre¬existing natural fractures or as mining-induced fractures, are uniquely necessary for every brittle-deformation ground failure including roof falls. Prediction and mitigation of roof falls requires quantifying the interaction between the fracture sets and the operative principal-compressive¬ stress-axis orientation (olo). The SOMA technique utilizes the interaction between fractures and 61o, specifically their dilation or closure, to calculate the orientation of 610, and to estimate the orientations of 62o and o3o. A stereonet program is used to organize and statistically analyze the fracture data, and then to determine the optimal orientation for the mine workings. The final step in the SOMA technique involves modeling the Laurel Mountain mine using a 2-D, finite-element, modeling program. This assures that the interpretations regarding the optimal working orientation closely match the deformation features observed in the mine. It also provides a predictive tool describing highly variable stress and strain partitioning, seemingly erratic rock-bolt failures and the interactions between different sets of mine workings within different coal seams. The critical importance of including the fracture sets is well illustrated in the Laurel Mountain mine where the effects of earlier mining of an underlying seam is contrasted in both a fractured and otherwise identical non-fractured model to illustrate the consequences of ignoring fractures in a predictive roof-fall model.

Jan 1, 2004

By Kot F. von Unrug

One of the major difficulties in the proper characterization of rocks surrounding excavations is the lack of data. The existing method for strength determination requires core drilling for sample collection, which, when performed from the surface is costly, and when performed underground, in addition to its cost, interferes with the mine operations. I-landling samples tested in the laboratory is troublesome and also influences the obtained results. Over several years. the author has been interested in developing a method of in situ strength determination that could rapidly generate strength data at low cost without a negative effect on mining activities. The recently developed borehole penetrometer is an improved version of the initial design ( 3 inches in diameter), which was used mainly for strength determination of coal pillars. The new penetrometer has been designed to work in 1.5 inch diameter holes. Such holes can be drilled at low cost with either hand held drilling machines or roof bolters. eliminating limitations of the previous system and opening the possibility of broad application. Strength testing of roof rock around longwall pillars is an example of an obvious benefit when designing secondary roof support. In the paper are given the principles of work of the apparatus and examples of the obtained strength data.

Jan 1, 1998

By Thomas Lautsch

After a short description of geology and gateroad roof support technology in the German coalfields the development of roof bolting as primary roof support is presented. Based on geotechnical parameters, different strategies used for roof support at depths of 3,300-4,600 ft. are explained. Strict quality management and a strict monitoring program are part of the entry development. A number of recently finished developments are described such as the introduction of standardized bolts and accelerated resin systems. An outlook of future goals is presented. The paper concludes with a comparison of roof control systems at RAG's mines in the US and Germany.

Jan 1, 2000

By Danny J. Van Roosendaal

An investigation was undertaken to determine hydrogeological effects of subsidence over a longwall coal mine in the Illinois Basin. At this mine, approximately 200 ft of bedrock overburden is overlain by up to 150 ft of glacial till and localized water-bearing glacial sediments within a buried bedrock valley that overlies the longwall panel. A finite-element mechanical model of the subsided longwall panel shows sufficient strain to enhance permeability throughout the bedrock portion of the overburden. However, examination of the actual surface subsidence profile indicates that the overburden deformation is localized in narrow zones at the panel edges. The aquifer units at the base of the glacial sediments are confined by glacial clays (above) and weathered shales (below). Elevated water levels were observed in the glacial aquifers during subsidence, which indicates that the clays and shales maintained confinement as the overburden subsided and deformed. A three-dimensional ground-water flow model was used to simulate the effects of subsidence on the hydrogeologic system and to estimate inflows into the mine. Longwall retreat was simulated by several model runs, each representing a new longwall face position. Simulated heads and mine inflows were calibrated with recorded water levels and observed inflow conditions by modifying the hydraulic conductivity of particular model layers. Model calibration indicates that the permeability of certain lithologies, such as plastic underclays and weathered shales, is not significantly enhanced by subsidence. A worst-case scenario was simulated by inserting a deeply incised, sandfilled valley into the model, which had only minimal impact on the simulated mine inflows.

Jan 1, 1995

By Chris Strawson

Moranbah North is a modern longwall operation mining premium coking coal in the Northern Bowen Basin, Central Queensland. Australia Moranbah North Coal (MNC) is owned and operated by Anglo Coal Australia. Extracting approximately 4.5m of the Goonyella Middle seam using a single longwall, the mine produces 5 to 6 Mt per annum. The weak coal (4.5 MPa), low strength laminated immediate roof, and strong upper roof sandstones, make operation of the longwall at depths in excess of 250m geotechnically challenging. In the past, there have been a number of production delays due to geotechnical issues. Roof instability resulted in a long stoppage on Longwall 105, and the eventual relocation of the face further outbye in the block. (Figure 1). When mining commenced in LW106, there was a requirement for uninterrupted production, particularly in the more demanding inbye area of the panel. Following discussions with a number of Geotechnical experts, it was decided to narrow the face, and test the theories of reduced face weighting due to spanning of the goaf across the panel. A new on line chock monitoring system, and microseismic monitoring was installed in order to characterise the difference in the rockmass response, as the panel width changed from 175m, to 265m, and then 300m. The nature of the goaf formation and impacts on face loading and production are discussed, having the unique opportunity of three different face widths within the same area.

Jan 1, 2007

By Liane J. Terrill

A 9O) ft (275 meter) section of a highly fractured beam of halite was mined from the roof of a supported thirteen year old drift at the Waste Isolation Pilot Plant (WIPP). The operation was conducted with a continuous miner and allowed the unique opportunity to collect spatial information on roof beam geology and to observe ground support conditions during the mining advance. Photographs and fracture maps of the mining face document the progress of the operation and the deteriorated condition of the strata. This case study details the drift's excavation and support sequences. the geomechanical response, and the conditions found in the exposed roof beam.

Jan 1, 1997

By David A. Newman

Severe roof control problems have plagued a West Virginia underground mine since its initial development in the late 1970's. Adverse roof conditions in the Eastern portion of the reserve result from a combination of mining: - perpendicular to the major principal horizontal stress direction, - near lineament intersections, - in areas where the ratio of horizontal stress to vertical stress exceeds 2.00 to 4.00, and - where laminated strata are present in the immediate roof. The focus of this study is to identify the engineering (mine orientation, roof safety factor) and geological parameters (lineament and horizontal stress orientation, horizontal to vertical stress ratio, and immediate roof lithology) causing poor roof conditions and to quantify their influence on projected mining conditions in the Western portion of the reserve Rose diagrams were used to illustrate the correlation between the orientation of eight hundred thirty-six (836) roof falls with that of forty-four (44) lineaments and the principal horizontal stress Similarly a composite map was used to demonstrate the relationship between roof falls and the ratio of horizontal stress to vertical stress The immediate roof strata across the property were categorized as either laminated or non-laminated based upon a thorough examination of one hundred sixty-two core logs. The presence or absence of fireclay, a rider coal, and carbonaceous shale was also noted because of the historic correlation with poor roof conditions. Rock strength and physical property tests were conducted on selected immediate roof strata recovered from coreholes drilled along the mains projected to the Western portion of the reserve Potential roof control problems in the Western reserve were quantified based upon the correlations developed by analyzing engineering and geological data The main conclusion is that laminated roof strata, the rider seam, and its associated fireclay will have the greatest influence on mining conditions. Roof problems attributable to horizontal stresses will not be as severe in the West due to an increase in overburden depth and corresponding decrease in the stress ratio. Recommendations are provided for mining orientation, the layout of mains and submains, panel location, and roof support based on intended service life.

Jan 1, 1999

By James H. Fletcher

Much data and experience has been accumulated, especially in the last 5 or 6 years, to show that modern roof trusses, both of the Birmingham type and the bolt-and-channel type, perform well to support certain kinds of coal mine roof. For some time now, roof trusses of the Birmingham type have been effective in controlling thick shale roof, which tends to sag and fall over a period of about eight weeks. These falls leave dame-shaped cavities which are reminiscent of the pressure arch theory of roof failure. Lately there has also been significant interest in supporting some longwall panel entries with channel and-bolt systems. These systems utilize two vertical center bolts and two 45 degree end bolts through a heavy gage channel. Both types of truss systems have long bolts, which require 8 to 9 Foot holes on a 45 degree angle, and point anchor resin bolts, which utilize a 2 or 3 foot rebar grouted in a thin annulus of polyester resin. Consequently, many coal mining people have been impressed with the very good anchorage characteristics and tensionability features of these systems. However, the same people have been proportionately dismayed to find that very often the most difficult part of introducing a trussing program to a mine is deciding on the machine specifications which determine the truss installation rate. The purpose of this paper is to indicate aspects of truss bolter design and use which are not shared by roofbolter design and its vertical pattern of bolts. Of the two types of drill units conventionally accepted, arm feed units and mast units, only one 15 judged suitable. While the arm feed unit offers low collapsed height with a long drill feed length, a great attribute for low seam use, it has not lent itself to being rolled over to a 45 degree angle. One reason for this inability is that many items which are conventionally mounted on the arm feed unit cannot be allowed to rotate with the unit. These items include drilling controls, tool Cray, lights, canopy or safety post, panic bar, and stab jack(s). If these items were independent of the arm feed drilling unit, they would have to be out of the way of the boom arms, links, and feed cylinder as the unit moves. To ease this problem, the arm feed unit can be made very short and even mounted across the machine. However, the resulting short feed length is undesirable. Several short drill steel extensions are required to drill a relatively long hole, making this procedure time-consuming as well as increasing the likelihood of drilling a crooked hole. Another reason for abandonment of the arm feed unit is that a very stiff drill guide (or "centralizer") is required to start the hole. Using the mast tilt control, the roof bolter operator will force the bit against the roof and will rotate the bit without feeding to cut sideways into the roof to form a hole face. Omission of this step often results in the skidding of the bit toward the rib. To date, no such sturdy drill guide has been designed for an arm feed drill unit. The mast style unit has been the most successful design for angle drilling. It is easily rolled over to a 45 degree angle or to 90 degrees and, on single head models, often has been made to roll through 180 degrees. The mast style is compact, in plain view and is easily fitted with heavy duty manual or hydraulically operated drill guides. It can be mounted directly on a machine chassis, or onto a swinging boom, or onto a slider arrangement in a broad variety of configurations which will accommodate controls, tool trays, stab and roof support jacks, lights, canopies, and panic bars. The only significant difficulty with mast drilling units occurs in lower mining heights when desirable low mast height is accompanied by undesirable short feed length. While the lead extension in the drill string can be as large as the mining height will allow, a short feed length will permit only the use of shorter middle extensions and drivers, and perhaps will. require the use of two different length wrenches. The associated problem is the relatively short distance from the drillhead to the drillguide. This distance is further shortened when raising the drillhead to get the bit to the roof. One might keep longer lead extensions on the machine; drill a vertical pocket for the bit, starting the angle hole at this pocket; or modify the mast. If the mast is made so that the distance from the drillhead to the drillguide remains constant while the mast extends, after which the drillhead travels up the mast, the shorter mast can be made to start the hole more easily. However, short teed length will still result in longer drilling time. It is obvious from these points that a proper choice of mast is simply the highest one which will be moved easily in the given mining height. This, of course, goes against the grain of most experienced coal mining people who realize that a relatively small decrease in mining height can create a problem correctable only with the purchase and installation of shorter masts. If the mining

Jan 1, 1982

By Chris T. Franke

The Waste Isolation Pilot Plant is a deep underground facility mined in bedded salt interspersed with persistent clay inter- beds. The clay seams delineate beams of salt above and below the excavations which strongly influence excavation performance. This paper reviews available rock mechanics information relevant to improving drift stability by removing the roof beam. Conceptual deformation mechanisms, relevant field data, and numerical analyses were used to investigate the effects of mining the roof of the excavations along a clay seam. Finally, a case study that applies the results of this investigation is presented.

Jan 1, 1997

By J. Marc Coolen

Marrowbone Development Company operates a large drift mining complex in the central Appalachian coal field. In 1997, five continuous miner supersections produced close to 9 million tons of raw plant feed. All production is from the Coalburg seam, a 5 to 10 ft thick coal seam with multiple coal benches and shale binders. The immediate roof lithologies consist of laminated shales, sandstones, or a mixture of sandstones and shales. The mine area is bisected by the regional Warfield anticline and associated Warfield fault. Mining conditions vary considerably due to frequent changes in seam and roof geology. The following geological categories can be distinguished: 1) presence of coal riders in the immediate shale roof, 2) excessive seam slopes around the Warfield fault, 3) splay faults associated with the Warfield fault, 4) zones of abundant slickensides, 5) abrupt roof lithology changes (sandstone - shale), 6) splitting and merging of different benches of the Coalburg seam, 7) sandstone washouts, 8) excessive roof dilution, 9) pillar sizing problems due to seam anomalies (soft fireclay binders). Several examples of these features are discussed in detail. Also, their relationship to roof fall occurrences is evaluated. Early recognition of potentially adverse geological conditions is of prime importance for the economic viability of the mining plans. Detailed core drilling in advance of the mining faces and coordination of the core data with underground mapping are the main forecasting tools.

Jan 1, 1999

By Christopher Mark

Mine planning for a new reserve is based on information obtained from exploratory coreholes. A critical component of an exploration program is the geotechnical evaluation. Poor assumptions about roof conditions greatly add to the risks of mining. Rock mechanics testing is central to a geotechnical exploration program. Typically, 3 to 5 uniaxial compressive strength (UCS) tests are made to characterize a particular roof unit at a given corehole. The average (mean) of these tests is taken as "The UCS" for that location. Isopach contour maps are then used to show spatial trends in roof strength. Two issues are raised by this traditional approach. The first is due to the large variability in UCS values that is typical even within a single unit from a single hole. The average UCS might be higher at Corehole A than it is at Corehole B, but the difference may not be statistically significant. The second issue is whether widely spaced coreholes can identify valid spatial trends in rock strength. The answer depends upon whether rock strength changes over distances that are longer or shorter than the corehole spacing. This is a classical geostatistical problem. While geostatics have been used to investigate many coal quality parameters, they have seldom been used to evaluate rock strength. This paper describes an extensive investigation of these issues conducted by the National Institute for Occupational Safety and Health (NIOSH) in collaboration with Peabody Energy. The study employed the Peabody Rock Mechanics Data Base which contains more than 10,000 individual test results. Data from four important roof units were subjected to statistical analysis: ? Brereton Limestone above the Herrin 6 seam (Illinois) ? Turner Mine Shale above the No. 9 Seam (Kentucky) ? Sandstone above the Eagle Coal (West Virginia) ? Shale above the Eagle Coal (West Virginia) The study did not find significant spatial trends in rock strength in any of the cases. Perhaps there are none, or perhaps the exploration coreholes were just too far apart to see them. These results have valuable implications for the design of geotechnical exploration programs.

Jan 1, 2004

By Wen H. Su

Multiple seam mining and its associated problems are very serious in Southern West Virginia where poor planning or lack of knowledge in seam interaction often results in complete loss of coal properties. Several measures have been proposed to alleviate the interaction prob¬lems under various multiple seam conditions, but few of them are spec¬ific in terms of mining plan. Two parallel approaches are adopted for this paper. One is field investigation which consists mainly of mine site visits and gathering of data. The other is the two-dimensional finite element simulation of mine workings using the data gathered as input. Two typical cases which represent the past and present multi¬ple seam mining practice in West Virginia were investigated. Much attention is focused on the interaction problems experienced and the important parameters controlling interaction. Several key parameters controlling interaction in multiple seam mining were analyzed. Proper mining plans, by which the interaction problems can be minimized, are proposed for the individual cases after a better understanding of the interaction mechanism was realized through finite element analyses.

Jan 1, 1984

By K. Y. Haramy

Rock burst and coal mine bump research using static and dynamic rock mechanics instrumentation has been conducted for several decades. Research efforts typically have been conducted using static instrumentation such as pressure cells, convergence measurements, multipoint borehole extensometers, and shield loading pressure measurements; dynamic rock mechanics instrumentation, for example, microseismic monitoring, has been used to help provide precursors to those violent failures. Unfortunately, these different styles of study have never been used in tandem. We report in this study the unique opportunity to have both static and dynamic instrumentation available for monitoring a longwall panel in a deep western coal mine and show that the two methods used in tandem provide information to understand the causes of a major bump. Several static rock mechanics instrumentation sites were carefully chosen along the longwall panel in the coal seam, packwalls, roof, and floor to provide the stress history during mining, including the stress redistribution that occurred before and after the major failures. Monitoring of the shield pressure on selected shields of the longwall face revealed that the cyclical loading anticipated during normal caving processes was not occurring. In addition, by locating each microseismic event (rock noise), the areas that were not adjusting to the changing stress field were clearly delineated by a lack of microseismic activity. The area of the longwall panel that did not adjust to the changing conditions ultimately failed. The combination of static and dynamic rock mechanics instrumentation data analysis provided the necessary information to determine the interaction of the structural members and the cause of the major failures at this mine.

Jan 1, 1988