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The purpose and current practice of cable trusses in underground coal mines is briefly outlined and details of various cable truss designs are presented. The main part of the paper describes the in-situ performance of cable trusses installed as secondary supports in gateroads and recovery roads of longwall coalfaces. The performance analysis is based on the results of in-situ monitoring carried out at a number of sites at two underground coal mines within Australia. Truss performance has been assessed as load developed in relation to roof convergence, face position and time. Additionally, the influence of strata conditions, installation pattern/density, roadway layout and geometry have been taken into account. An analytical model has been developed to predict truss loads. The principal variables are listed and the basis of the model is described. Based upon behaviour observed in the field the model has been progressively upgraded and validated against measured performance. Guidelines on truss performance and factors which influence it are presented and the implications of these to the use of cable trusses is discussed. Possible methods for improving truss performance in certain strata conditions are outlined and some comments are made on the value of cable trussing compared to cable bolting in these situations,

Jan 1, 1992

By Bret E. Leisemann

The geomechanical environment in which highwall mining must operate represents a hybrid between surface and underground operations. Mine design must take into consideration the complexities associated with unsupported spans, pillars with extreme aspect ratios and the instability related to the interaction between coal extraction and the highwall stability. The rock mass adjacent to the highwall, and for a significant distance in, is affected by lateral stress relief and blast damage. This paper describes a rigorous site investigation program aimed at providing some of the input parameters for the design of highwall mining layouts at the Moura Mine and focuses upon defining the nature and extent of the relaxed zone behind the highwall. The combination of borehole imaging and C factor analysis allows delineation of blast damage and relaxation of the rock mass immediately adjacent to the highwall excavation, as well as the more subtle relaxation effects deeper into the rock mass. This has led to the formulation of a proposed geomechanical model for the relaxed zone for direct input into highwall instability analysis and mine design.

Jan 1, 1993

By Styles P.

To determine whether 130 felt earth tremors around Edwinstowe, Nottinghamshire U.K, which also experienced severe surface fissuring, were caused by coal extraction, a surface seismometer array was established around Thoresby Colliery. Over the next year, 785 microseismic events were detected. The spatio-temporal seismicity patterns are clearly associated with the commencement, continuing extraction and closure of faces. Of particular note are events which locate at the surface and appear to be related to the active fissuring. Events occur within days of commencement of production and cease when production finishes, with good correlation between face advance and hypocentral position. Naturally occurring microseismic events have also been detected up to 1 km ahead of active longwall faces in the Midlands using triaxial geophone packages grouted into the seam together with a surface seismometer in the top of the borehole. The quality of these data were very high and guided waves can clearly be seen with the dispersive characteristics associated with seam waves. In one experiment more than 2000 events were detected in only two days of monitoring even in a relatively noisy surface environment. This paper demonstrates how very accurate locations ( 5 10 metres) can be generated using three-component digital data from only one borehole. The surface seismometer and the borehole P-wave onset can give extra precision for the height of the event relative to the seam. The event distributions give a dynamic, three-dimensional image of the developing patterns of fracturing above, below and ahead of the longwall face with important implications for roof control, subsidence prediction and gas migration.

Jan 1, 1992

By C. Babcock

This Bureau of Mines report considers some equations for mine pillar design from 1833 to the present. Most of the equations are of essentially the same kind, using pillar width to height relationships. The effects of end constraint and test volume size, that were implied in some of these early equations have received special emphasis in recent years. Experimental testing in the laboratory throughout the period given has been on prisms or cubes of coal or rock. Most of the present concepts have been known since about 1912. Wilson has recently emphasized the effect of end constraint in his confined core concept, but a number of other earlier relationships have indicated an awareness of the effects of constraint as well.

Jan 1, 1981

By Ernest A. Curth

An estimated 49 billion tons or 29 percent, of the coal reserve base to a depth of 1,000 feet in the eastern United States fall in the 28- to 42-inch range. Often left out as a consequence of selective mining, it is a paramarginal source that will become increasingly important. At present, all active thin seam longwalls are located in the Eastern Coal Province, and a number of operations have been suspended due to the erosion of the market for metallurgical coal. Extensive premining studies, appropriate mine planning, reliable equipment that is designed with ergonomics in mind, adequate roof support, intensive training, and strong face discipline are the ingredients that provided a measure of consistency in production for the last 10 to 16 years, particularly in the Pennsylvania bituminous coalfield where single drum shearers dominate. Plows, which were found to be very productive in suitable strata in Europe, are used in the Pocahontas field. Low seam four- and six-legged shields have been introduced recently. The future holds automation at progressive levels. The first step is batch control with the benefits of quickly applied roof support and reduction of workmen exposure to environmental hazards.

Jan 1, 1981

By David Bartsch

The Ohio Valley Coal Company (Ohio Valley) has operated the Powhatan No. 6 Mine in Belmont County, Ohio since 1972. Ohio University owns a 455-acre research farm over a portion of the Ohio Valley reserve containing 51 acres of old-growth forest known as Dysart Woods. In 1988, when Ohio Valley applied for a new mining area for its first longwall located several miles away from Dysart Woods, Ohio University claimed that mining at that distance would lower the regional water table and kill the old trees. In response, Ohio Valley began to study potential impacts from longwall mining on hydrology, soil moisture, and tree growth. Early studies examined the impacts of longwall mining on hydrology, soil moisture, and tree growth directly over mined areas compared to woodlots that were not undermined. A subsequent study examined tree-ring growth for a five-year period before and after longwall mining for several woodlots, some directly over panels and others at distances of up to 2000 ft away. No significant differences in growth rate associated with longwall mining were observed. New research was begun that compared longwall-mining impacts on groundwater hydrology, available soil moisture, and tree growth in two stands of mature trees located near Dysart Woods in order to confirm the earlier results. Using a test technique known as "Before-and-After-Controlled-Impact-Pairs," data was collected from one woodlot (test) that was undermined and from another (control) that was not undermined. Data collection began several years prior to longwall mining and has continued for several years after mining. The data for the test site shows that although the shallow and deep groundwater levels were affected by longwall mining, the soil moisture, vigor rating, or the growth rate of trees did not decline significantly compared to the control site where there were hydrologic impacts. The study shows that there are no known significant impacts to trees from longwall mining. The results of these studies were used by Ohio Valley to obtain State approval to longwall to within 300 ft of Dysart Woods

Jan 1, 2005

By Erik Westman, Benjamin Mirabile

"Quantitative evaluation of secondary support performance in longwall tailgates in U.S. coal mines was introduced with the U.S. Bureau of Mines (USBM) Wood Crib Performance Model. This model evaluated crib performance based on load capacity, stability, and convergence characteristics. The introduction of a large variety of alternative secondary supports in the 1990s prompted a full-scale load-displacement testing program conducted by the USBM and, subsequently, NIOSH. The application of the ground response curve concept to secondary supports provided a more complete estimation of secondary support loading in longwall abutment zones. Secondary support loading from floor heave, horizontal stress effects, and pillar yield are captured by the ground response curve. A point on the ground response curve can be measured by instrumenting a secondary support with load and convergence measurement devices. A segment of the ground response curve can be measured by instrumenting two secondary support types of varying stiffness. Numerical models have been used to construct complete ground response curves for longwall tailgates based on measured secondary support performance. The ground response curve concept can be used as a design methodology for secondary supports in the #2 entry, as well as the longwall tailgate. Extending published evaluation and design methodologies for secondary support in longwall tailgates, a ground response curve unique to the loading conditions and required functionality of the #2 entry can be developed. Numerical models can be used to construct full ground response curves based on measured data in the #2 entry.INTRODUCTIONA longwall ventilation system is designed to continuously dilute and move methane-air mixtures and other contaminants from the active face, to bleeder systems, and out of the mine. The #2 entry of the tailgate gateroad serves to transport contaminated air from the longwall face to the bleeder system. The #2 entry is required to function as an air course between two longwall gobs and is subject to significant ground movement and loading. Current ground support design and evaluation practices are based on empirical evidence and little to no quantitative data on support performance exists for these entries. Substantial volumes of research have been conducted on ground movement and support performance in longwall tailgate entries. Following the successful ground support evaluation and design practices established for longwall tailgate entries, a practical and efficient design methodology can be adapted for supporting the #2 entry of a three-entry gateroad system."

Jan 1, 2018

By André Vervoort

To learn more about the fundamental roof behaviour in South African coal mines, accurate underground measurements of roof deflection were conducted in bord and pillar panels during development and pillar extraction. To conduct these measurements a survey levelling technique was adapted. The overall system reading accuracy was established as being within 0,4 mm on a 15 m sighting distance. The underground sites were modelled and, for a roadway development, similar symmetrical curvatures were calculated by a linear elastic model, as were measured underground. In a pillar extraction panel, a significant amount of bed separation was measured and the roof underwent an a-symmetric movement due to the close presence of the mined out area at one side of the roadway.

Jan 1, 1992

By Mictck Rataj

Methods of rock reinforcement in Australian coal mines are being advanced progressively through the better understanding of rock failure mechanisms around roadways and the-development of improved bolting hardware and installation techniques. It has been recognised that to improve the performance of rock reinforcement systems, the stiffness of the system as a whole has to be improved. New style bolts and accessories together with high pre-tensioned roof and cable bolts are being developed as a result. A number of the new techniques used to stabilise the roof, especially in weak roof/high horizontal stress areas have been site proven and are now currently used on a routine basis as a means of improving development rates and gate road stability during longwall retreat.

Jan 1, 1997

Previously, a variety of risk rating systems were used on different mines. However, these were mostly site specific and geology based systems. The purpose of developing a uniform rating system, to be used on all Anglo Coal open cast mines, is to provide an unbiased, standard and quantifiable assessment of the risk from highwall and lowwall failures. This takes into account relative differences in the importance of hazards, as experienced on each mine as a result of different combinations of geotechnical factors and mining conditions. The system is based on critical geotechnical and other parameters that have been identified by mining personnel and rock engineers. The primary advantage of this risk rating system is that all open cast mines in Anglo Coal use an almost identical system, which enables the user to compare their risk to the risk at neighbouring mines and/or other open cast mines. The system can be adjusted to meet local, mine specific requirements. The implementation of this new system (a computer program that automatically calculates the risk) has been made as practical and easy to use as possible. Therefore, this tool can be used by personnel from other mining disciplines not directly related to rock engineering.

Jan 1, 2004

By George B. Quaye

Conventionally, roadways in South African collieries are 6 to 7 m wide. This dimension is chosen so as to allow maneuvrability of appropriate mine machinery and equipment, to meet production requirements and to ensure roof stability. Over the last two decades, a great deal of research effort has been expended to understand the behaviour of roadways in different strata within this width range. The results have improved the understanding of the strata behaviour to such an extent that 6 m wide roadways are developed with a high degree of confidence. With the advancement of technology and the quest to mine at a lower cost, some mines in South Africa have started considering the benefits of other types of mine machinery that will require much wider roadways to ensure optimum productivity. The roadways required will have to be about 10 - 14 m wide. The challenge confronting these coal mining companies is the potential instability that might be associated with such wide roadways. This dilemma is aggravated by the fact that, apart from theory, nothing is known about the expected behaviour of such wide roadways. This led to experimentation at a selected site in a South African coal mine where 10 m and 14 m wide roadways were developed and monitored. This paper highlights the experimental procedure adopted, discusses the results obtained, compares the prediction of elastic theory to reality and set the precedence towards understanding the behaviour of wider roadways in South Africa.

Jan 1, 2001

By John L. Edwards

The underground coal mines in the Lake Macquarie area of the Newcastle Coalfield, New South Wales, Australia are constrained by surface subsidence restrictions as low as 150 mm. The mining environment is shallow, generally between 100 m and 200 m, often multi-seam and commonly overlain by residential development or tidal waters. Most mines use partial or panel and pillar extraction techniques to meet the subsidence restrictions and allow maximum resource recovery. This paper gives an overview of the results of a recent two year government funded research project aimed at developing a mechanistic, geomechanically based mine design criteria for subsidence control. Three extraction layouts were monitored in detail to assess the relationship between mining geometry, pillar behaviour and surface subsidence. Pillars at each site were instrumented with stress cells, convergence monitors and extensometers. Field data was used to initially calibrate, and subsequently, verify stress and displacement predictions made by the three-dimensional displacement discontinuity code, SUBSOL. Back-analysis of the modelling results indicate SUBSOL's capability as a tool to satisfactorily predict surface subsidence magnitudes and pillar loads for alternative pillar and panel layouts.

Jan 1, 1993

By David I. Hoyer

It is shown that by monitoring longwall leg pressures in real time, warning can be given for significant weighting events and the formation of roof instabilities, such as roof cavities, several hours in advance. Longwall Visual Analysis (LVA) is a software package that continuously monitors shield pressures and shearer position in longwall mines. LVA has been running on 19 Australian longwalls for up to four years, and as a result a very substantial database of shield pressure trends in a wide range of longwall situations has been collected. This database has been analyzed to develop indicators that will give operators and geotechnical engineers advance warnings of developing conditions such as weighting events and difficult roof conditions. LVA displays live charts of data, such as shield leg pressures, loading rates, and yield frequencies. Data analysis techniques were applied to historical data records from multiple longwall sites in order to develop a ?Cavity Risk Index? (CRI). The CRI indicates the risk of roof cavities developing, and it is based on pressure trends that indicate significant yielding and loading rates spanning a region of relatively low support. Case studies from different longwalls are given, showing how the CRI can give real-time advance warning of the formation of roof cavities with their anticipated location on the face. The limited number of case studies done so far indicate that the predictions have a high degree of reliability, but more work is required to quantify this.

Jan 1, 2011

By Ding Xu Chu

There are many methods developed for measuring the crustal stress in China. Among them only two methods have been adopted in practice, i.e., overcoring method (for which the sensors include: Piezo-magnetic stress gauge, steel-ring strain cell, 3-dimensional hollow-inclusion stress ,gauge, and Leeman's-like stress gauge) and hydraulic fracturing method. For the former method, one of those sensors is installed in a borehole, 36 mm in diameter, then prestressed, and finally overcored with a drill bit of 130 mm in diameter. After that the deformation of the borehole wall is measured, from which the in-situ stresses are determined. In order to avoid the errors resulting from the difference in elastic modulus for various kinds of rocks as well as from the installment of the cells, we have developed a technique to calibrate the sensor on-spot using a confining pressure calibrator and produced a very good result. For the later method, the measurements are achieved by isolating a certain section of a borehole and then injecting water to pressurize it till a tensile break is induced. It is identical to the typical technology used internationally. For many years, through field observations and a large number of laboratory experiments, we believe that the most efficient way of stress measurement for engineering purposes is the overcoring method using Piezo-magnetic stress gauge. It is characterized by stable and reliable measurement values, high accuracy and high adaptability. Taking the measurements in the intact rock mass as an example, the measured stress magnitudes are within accuracies of 8% and the errors in orientation are about 5 degrees. It can even be used perfectly in deep water. Up to now, we have performed in-situ stress measurements in more than 70 sites and obtain more than one thousand sets of data. The deepest depth used in the overcoring method is 80 m in vertical borehole and 780 m in tunnels.

Jan 1, 1987

By William Monaghan

Over the period 2000 to 2003, roof falls have accounted for 4 to 14% of the fatalities in underground mining operations. The National Institute for Occupational Safety and Health (NIOSH) is conducting research to reduce the frequency, exposure, and risk of these events through an ongoing program of field and laboratory studies. One area of research involves the evaluation of polyurethane grouting technology that is commonly used to stabilize fractured mine roof strata. Concurrently, NIOSH is conducting work to evaluate the application of ground penetrating radar for advanced delineation of problematic mining areas. In this study, NIOSH partnered with Sub-Technical, Inc. who injected a polyurethane grout (also known as glue) into a roof area of NIOSH's Safety Research Coal Mine. The mine roof area was scanned using ground penetrating radar technology before and after grout injection in an attempt to determine the extent of grout infiltration. A comparison of the pre-and post-grouting radar records showed a significant change at a mine roof depth of about four to five feet. The interpreted radar records were then compared with drill core information, borescope evaluation of roof-bolt holes in the study area, and underground observations. At this site, the interpretations of the radar records correlated with data obtained from the core holes, borescope evaluations and underground observations. This study showed that ground penetrating radar technology can be a useful tool for detecting changes in mine roof due to the injection of the grout.

Jan 1, 2004

By C. J. H. Brest van Kempen

The techniques developed should provide a useful tool, not only during the initial formulation of a suitable bolting plan for a new section, but also for periodic check¬ing on the utilization factor of hardware installed in the roof. Indications are that with present practice, utilization seldom exceeds 507.. Retro-fittable bolter equipment now available (FCBS?) allows doubling of the utilization factor in many cases. We are also introducing a computer pro¬gram based on the material outlined in this paper. This program can save a great deal of time in examining the real effect of implementing alternate bolting plans or using different installation techniques. It allows fine-tuning of bolt lengths, spacing and installation tension to mini¬mize roof support cost, while maximizing safe life of the mine opening. We have developed a procedure which allows easy, site-specific quantification of the need for reinforcement of a mine roof. The concept is based on the forma¬tion of a direct link between theory and empirical data to custom-fit each mine section. In its simplest form, the proce¬dure requires only recording of roof bolt torques in a sample working place (16 to 25 bolts) several times during the first week or so after initial bolting. The way in which the torque pattern changes during this time can reveal whether or not the bolts carry enough tension to maintain stability in that specific roof. From the same data, the nature of the anchorage horizon is derived. Specifically, shale or limestone can be distinguished from sandstone. If the procedure is repeated a few times (within the same mine section) with different bolt installation tensions, we can numerically characterize the roof in terms of one single value: an "effective angle of internal friction". This number can then serve to calculate trade-offs in bolting plan design, such as bolt length, bolt spacing and optimum bolt tension. Additional refinement is possible if a sagmeter is installed in the sample work¬ing place being monitored and if bolt anchorage capacity is determined. The lat¬ter can be done automatically on each bolt as it is being installed, using a system we developed and patented. Alternately, a number of pull tests could be performed. Using the sagmeter, bolt torque and anchor¬age data, it is possible to project the safe life of the supported opening. The details of the procedures described apply specifically to tensioned roof bolting, but a brief discussion paints the way for a similar analysis of full-column resin bolting also.

Jan 1, 1986

By Syd Peng

It is necessary to fully understand roof geological features in order to design a proper roof support system for underground coal mines. These geological features include: rock type, rock strength. rock layer interfaces, voids, cracks, and bed separations. Traditionally, engineers have relied heavily on the use of surface borehole log data to identify geological features. Since surface boreholes very expensive to be drilled in close spacing, log data are often insufficient to provide detailed information on the roof geology. Recent advances in control technology have made a large amount of roof bolter drilling data available to mine personnel. These data can be obtained during normal roof bolt installation cycle with little or no interruption in mining operations. This research project has developed several methods to identify geological features in a mine roof from measured drilling parameters. Power trend line analysis can be used to locate discontinuities and interfaces in mine roof. Power averaging can be applied to determine the relative strength of roof strata. Thrust trend line analysis can be used to locate fractures in mine roof. The results of these methods are combined to produce a mapped roof geology In a production environment, manually interpreting and managing collected drilling parameters is not practical since hundreds of holes will be drilled during a production day. A new software package, MRGIS, has been developed to allow mine engineers to make use of the large amount of roof drilling parameters for roof support design. An actual underground test site is presented to demonstrate the analysis and display capabilities of MRGIS.

Jan 1, 2003

By W. A. Cincilla

The use of resin-grouted rock anchors continues to gain popularity in underground wall mines in the United States. It is estimated that during 1985, more than 30% and as many as 35% of all roof bolts used employed sane from of resin anchor. The so-called 'combination' bolting systems, which generally utilize a resin pint-anchor, have also found increased application. This increased application is particularly true in poor ground, where more 'traditional' support measures have performed poorly. Virtually all resin bolts in use today in U.S. mines employ cartridge-type delivery systems, which in general, have produced good results under a variety of conditions. However, a host of inconsistent and sometimes uncontrollable variables may significantly alter the optimum bolt performance. These variables have been identified and are well documented in the literature (Snyder, et al., 1979; Kwitoweki and Wade, 1980; Serbcusek and Signer, 1985). The beau of Mines, Denver Research Center (DRC) conducted a major research project aimed at investigating the effects of selected aspects of installation procedures on resin-bolt performance. One objective of the project was to establish anchorage criteria for full column resin bolts installed in actual mining conditions. This piper describes the results of a particular study which focused on quantifying, through field tests, possible reductions in bolt anchorage due to installation of less-than optimum resin columns. These columns might result from losses of resin to surrounding strata, or non-compliance with recommended installation procedures. Over 1000 resin bolts having grout columns ranging from 12 inches up to 48 inches (full column) were installed and pull-tested at 13 underground test sites located in a total of 11 mines (Fig. 1). Four basic roof-rock types were included: shale, top coal, sandstone, and limestone. The majority of the bolts employed 3/4-in-diameter, 48-in-long, type 40 rebar having a minimum '~ated' yield point of 17,600 lb (8.8 tons). Actual yield strength may vary by manufacturer.

Jan 1, 1986

By John L. Hill

Over 50 mines of the Appalachian and Illinois Basins are presently experiencing poor ground conditions believed to be caused by overlying stream valleys. The Bureau of Mines is conducting research into the causes of this problem with the aim of developing a predictive method for determining the safe limits of mining beneath stream valleys. In addition, the factors that control instability in these regions will be analyzed to provide design guidelines for adjusting support requirements. This paper presents the theorized causes of poor ground conditions beneath stream valleys and compares the theories with the results of a Bureau study of the Birchfield Mine No. 1 and a preliminary investigation of the Davidson Mine, both of southern West Virginia. The Birchfield Mine excavated four main entries beneath a 50-ft wide stream with only 50 ft of overburden. The valley flood plain is over 500 ft wide, flanked by valley wall slopes of 25 deg and relief of over 950 ft. In-mine monitoring of strata movement and detailed geologic mapping revealed that roof rock strength was greatest under high cover and weakest along outcrops and beneath the valley, with localized areas of decreases in rock mass quality. To improve the rock mass quality beneath the stream and reduce the threat of inundation, an extensive drilling program was undertaken to inject grout into the overburden from the surface. As a result, mining beneath the stream was practically routine from a ground control perspective, and water was not encountered until mining had advanced to a point beneath the floodplain that had not been grouted. The only ground control problems which were encountered were entirely a result of localized poor rock mass quality. In contrast to the Birchfield Mine, the Davidson Mine experienced failure of intact roof rock at a depth of 300 ft beneath a V-notch shaped valley with no change in rock mass quality. Finite element analysis was conducted to study the premining in situ stress at the two sites. The results are discussed within this paper and provide insight into the causes for the types of failure seen beneath valleys of varying shapes and depths.

Jan 1, 1988

By Alan A. Campoli

"Pumpable crib support for longwall gates and ventilation bleeder entries is becoming more prevalent in United States coal mines (Figure 1). The pumpable crib selection and spacing design has been greatly refined by the full-scale testing conducted at the National Institute for Occupational Safety and Health (NIOSH) Mine Roof Simulator. The peak load on 24-, 27-, 30-, and 36- inch diameter J-Cribs is, on average, 300, 400, 450, and 650 kips, respectively. Standard J-Cribs have useful deformation of approximately 8 inches. The maximum useful deformation can be extended to 18–20 inches with an exterior reinforcement layer or an interior double wall, for all four J-Crib diameters. J-Crib diameter is primarily limited by a maximum 4.5:1 height-to-diameter aspect ratio. Spacing is controlled by immediate roof stability, anticipated pillar deformation, and roof bolting practice. Approximately 500,000 J-Cribs have been successfully placed in 18 different United States longwall coal mines. The selection and spacing of the J-Crib utilized to date is summarized for each mine.IntroductionLongwall productivity and safety depend on ground stability in gate and ventilation bleeder entries. Numerous support technologies have been developed and employed in United States longwall mines since the 1990s to replace traditional wooden cribbing. Many, if not all, of these technologies have been evaluated at the NIOSH Pittsburgh Research Laboratory since the mid-1990s (Mucho, et al., 1999). The NIOSH Mine Roof Simulator has a 3,000-kip vertical load and 24-inch displacement capacity, which has facilitated in-depth analysis of complex multimaterial support technologies like the pumpable crib. The extreme cost and human exposure risk associated with modern longwall mining prevent the trial and error development that would have been required absent the Mine Roof Simulator research. As a direct result of this research, pumpable cribs are becoming a major component of the standing support industry, providing the peak capacity and extended deformation characteristics necessary for successful gate and bleeder support. Recent research by Zhang, et al. (2012) and others have defined the required performance characteristics."

Jan 1, 2015