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By Martin McKinnon, Guy Reed, Russell Frith

"It is arguable that the development of reinforcing roof bolting systems has largely stagnated in recent times, primarily due to the prevailing industry view that few, if any, further improvements can be made to what currently exists. However, this paper contends that reinforcing roof bolting systems can be further refined by considering both the specific manner by which horizontally bedded roof strata loses its natural self-supporting ability and the specific means by which reinforcing roof bolts act to promote or retain this natural self-supporting ability. The Australian coal industry insists on minimising bolt-hole diameter to maximise load transfer and on targeting full-encapsulation by any means necessary. This has led to a significant, albeit unintended, consequence in terms of overall roof bolting effectiveness, namely increased resin pressures during bolt installation and the associated potential for opening bedding planes that may have, otherwise, remained closed during the bolt installation process. Given that the natural self-supporting ability of roof strata is strongly linked to whether bedding planes are open or closed, logically, minimising resin pressures should be a significant benefit. This paper focuses primarily on three key issues that relate directly to the function of the roof bolting system itself: (i) the importance of proper resin mixing in the context of maximising load transfer strength and stiffness, (ii) the importance of minimising resin pressures developed during bolt installation, and (iii) the importance of maximising the effectiveness of the available bolt pre-tension. All mine operators should be invested in improving the individual effectiveness of each installed roof bolt, even by relatively small incremental amounts, so this is an important topic for discussion within the mining community."

Jan 1, 2017

By George J. Karabin

The Roof Control Division of the Pittsburgh Safety and Health Technology Center, MS HA, is routinely involved in the evaluation of ground conditions in underground coal mines. Assessing the stability of mined areas and the compatibility of mining plans with existing conditions are essential elements in assuring a safe working environment at a given site. Since 1985, the Roof Control Division has successfully used the Boundary Element Method of Numerical Modeling as a tool to aid in the resolution of complex ground control problems. This paper will present an overview of the modeling methodology established and details of techniques currently used to generate coal seam, rock mass, and gob backfill input data. A summary of coal and rock properties used in numerous successful evaluations throughout the country will be included and a set of Deterioration Indices that can aid in the quantification of in-mine ground conditions and verification of model accuracy will be introduced. Finally, a case study will be detailed that typifies the complexity of mining situations analyzed and illustrates various techniques that can be used to evaluate prospective design alternatives.

Jan 1, 1994

By Denise Y. Dizon

The objectives of this study were to document the hydrologic impacts of longwall mining on streams, identify the factors affecting the extent of stream dewatering, and develop empirical trends to predict the extent of dewatering and recovery of streams in advance of mining. This study on stream impacts is a companion to a stud; conducted on ground water impacts from longwall mining, which was done as part of the same research project at the' same locations and times. The study areas are three mine sites in north- central West Virginia. Mine A, located in Upshur County, has relatively thin mine overburden under the valley streams (105-115 feet). Stream monitoring there continued the earlier work of Cifelli and Rauch (1). Stream dewatering was severe with streams often going dry over the longwall panels during baseflow conditions. The lost water was probably flowing to the mine. Dewatering began in streams over the mine entry areas adjacent to the panels and as much as 380 feet from the nearest panel edge. The angles of dewatering influence ranged 22 to 72 degrees with respect to the nearest mined panel edge. Discharge then typically increased as the streams- flowed across the panel edge, but then decreased again over the panels. These streams were often ponded by local reversals in hydraulic gradient caused by land subsidence, and often dried up after flowing a few hundred feet across the panels. This occurred over panels that were previously mined up to eight months before. Mine B, located in Monongalia County, has relatively thick mine overburden (475-485 - feet) under the major stream. Stream dewatering was severe there during warm season baseflows, with complete dewatering occurring then over the longwall panels. This dewatering occurred over mined panels up to five years old, probably in response to a major anticlinal axis lineament and fracture zone in the stream valley. Mine C. located on the West Virginia -Pennsylvania border, has relatively thick mine overburden (about 585-630 feet) under the monitored streams there. stream dewatering was only partial, and these streams typically recovered their lost discharge downstream of the mine, indicating that lost water was not recharging to the mine. Only longwall panels less than one year old were associated with stream dewatering.

Jan 1, 1990

By S. K. Sharma

The present study attempts to critically review the pillar-design as required by Coal Mines Regulations of India vis-a-vis latest rockmechanics formulae, Equivalent material modelling was the approach adopted for testing the safety and stability of the pillar sizes predicted by various formulae. Two parameters, considered here for calculating pillar sizes was the stresses acting over the pillars and the strength of the pillars. Stresses acting on the pillars were calculated by the tributary area method. Based on the studies of Coates and Hustrulid, the stress obtained was reduced by 40 percent. For calculating the strength of the pillars, three formulae, namely Holland and Gaddy, Salamon and Munro and Central Mining Research Station, India, were used. The pillar widths were then determined by equating the product of recommended factor of safety and pillar stress with the pillar strength.

Jan 1, 1992

By Shosei Serata

Serious floor heave of up to 2.4 m in a 2.4-m high mine entry was eliminated by applying the stress control method of mining, as a last resort, at the No. 5 coal mine of Jim Walter Resources, Inc., in the Black Warrior coal basin near Birmingham. Alabama. Underground observation of the first 3-room entry created using the stress control method is discussed here. The behavior of the test entry, which eliminated the heave problem, is analyzed in relation to similar studies conducted in a salt mine under similar ground conditions by utilizing finite element analysis.

Jan 1, 1984

By K. Haramy

Coal mine bursts or bumps involve the violent, rapid failure of coal and rock in or around a mine excavation. Failure is normally associated with high stress and brittle or brittle-elastic materials; in coal mining, bursting may also be associated with desorbing gas, and this type of failure is termed an outburst. This paper discusses factors affecting coal mine bursts and several methods to predict bursts, including microgravity, microseismic, rheological, rebound, and drilling-yield methods. Methods to avoid burst conditions using volley firing, hydraulic fracturing, and auger drilling are also discussed. A case study is presented in which the redistribution of stresses and displacements was found to be associated with a burst at a deep longwall coal mine. The research showed that stresses are redistributed following the burst.

Jan 1, 1984

By Ali Naeimipour

Availability of accurate geological information about the rock mass surrounding a tunnel or mine working is the key to enhance and optimize the required ground support measures. Various approaches have been used to obtain data on ground conditions, including the possibility of characterizing the ground while drilling for roof bolts, by the Measurement While Drilling (MWD) system, via optical, sonic and acoustic televiewers, and using various types of gamma probes. While there are many different systems for characterizing the ground in longer, larger-diameter holes drilled downward, the options for using probes and logging systems in short, small-diameter holes drilled in horizontal or an upward direction are very limited. This paper will present the results of recent field tests using optical televiewer (OPTV) to analyze the joints and voids in boreholes drilled in different directions. The studies show that OPTV is a promising tool; however, getting a good quality image from the upward drill holes still need to be studied. Meanwhile, estimated in-situ strength of the intact rock is needed to complement the jointing pattern for rock mass classification. For this purpose, preliminary steps for the development of a borehole scratcher to estimate rock strength will be discussed in this paper.

Jan 1, 2014

By Christopher J. Brown

"Proper ground stability in gate roads has been a critical aspect of success in longwall mining. In particular, the tailgate has proven to be one of the more active mine areas that needs to be managed (Barczak, Dolinar, and Mucho, 1999). As longwall tailgate support technologies continue to be developed and improved, Strata Worldwide has worked to design and manufacture a revised Propsetter to provide additional support capacity and enable applications at greater mine heights. The support is targeted for tailgate applications; however, it can also be used for roof support in various types of mining environments. This paper describes the design changes for enhancing the support capacity of the Propsetter to more closely align its performance with other support technologies currently being employed in longwall tailgates. Balancing changes to each design aspect of the Propsetter was key to successfully improving the roof support load capacity while maintaining controlled deformation. Increasing the load capacity would allow a mine to improve support efficiency with a smaller, less intrusive support compared to support alternatives and to reduce the cost per foot of supported entry. The target support capacity was 80 to 100 tons at two inches of deformation to more closely match the performance of competing tailgate supports, such as the 24-inch-diameter CAN®. To achieve this goal, changes were made to all design aspects of the Propsetter: diameter, wedge cut depth and angle, pod size, and confinement rings. Test results from the NIOSH Mine Roof Simulator verified that these design changes were successful in increasing the capacity of the Propsetter to the 80- to 100-ton range at two inches of deformation. INTRODUCTION Roof support is one of the most fundamental elements of the underground mining process. As materials are removed from the earth, geological conditions will almost always require some form of roof support to keep the extraction area open and stable. Coal mining environments often experience difficult geological conditions that require both primary and secondary forms of roof support to keep the mining environment safe and economical. The secondary support system serves as additional support in the case the primary support system is inadequate. If designed properly, the secondary support system will also assist in maintaining integrity of the immediate roof beam (Barczak, 2000)."

Jan 1, 2017

By Keith Heasley, Christopher Newman, Zach Agioutantis

"Previously, in 2013, a laboratory version of the computer code ARMPS-LAM, introduced at the 32""d International Conference of Ground Control in Mining, successfully integrated the laminated overburden model of LaModel into an ARMPS-style pillar stability analysis. The ARMPS-LAM analysis has now been fully integrated into the ARMPS program allowing users to more accurately classify the stability of a given mine design utilizing a laminated overburden model for the determination of pillar loading within theAMZ. ,This paper introduces the new Windows-based ARMPS-LAM program. The program's ease of use, output result generation, and limitations are highlighted through the parametric analysis of three hypothetical case studies where traditional ARMPS output is compared to output generated by LaModel. As mining operations continue to produce at greater depths and in more complex geometric and geological conditions, flexible and efficient analyses of mine stability for underground room and pillar mining has become more and more essential. The ARMPS-LAM program provides the mining industry with an empirically backed, numerical design tool for the evaluation of underground room and pillar mine plans.INTRODUCTIONIn the Southern Appalachian coal fields of the United States, retreat room-and-pillar methods have allowed mining operations to maintain high rates of production in the face of adverse geologic conditions and shrinking coal reserves. Over the years, research has been dedicated to understanding and mitigating mining-related dangers due to second mining practices, such as: pillar squeezes, floor heave, roof falls, pillar bumps, etc. However, as mining operations continue at deeper depths and with more complex geometric and geologic conditions, there is an inherent industry need for more accurate, flexible, and faster evaluations of stability for underground room-and-pillar mining."

Jan 1, 2016

By D. W. H. Su

Following a roof fall in the B12 headgate of a CONSOL Pennsylvania Coal Company (CCPC) underground coal mine in February 1998, CONSOL R&D and Exploration initiated an underground roof geology reconnaissance program along the B13 and B14 developments to define potential roof instability areas and to recommend appropriate supplemental headgate support. This reconnaissance program revealed the presence of a thick, massive Pittsburgh sandstone channel overlying the B13 and B14 panels. Surface and underground core holes confirmed the close proximity and the very massive, nonlaminated nature of the sandstone. To mitigate anticipated adverse conditions at the B 13 and B 14 longwall faces due to delayed caving of the massive sandstone behind the shields, a task force consisting of Operations, Exploration, and R&D personnel was formed in May 1998. Several operational adjustments were made by mine management, including repairing damaged lemniscate links, maintaining immediate forward shield support, and implementing measures recommended by the task force intended to boost the shield set load density at the face. After reviewing available geotcchnical information, the task force also recommended fracing of the massive sandstone over the B13 and B14 panels to enhance caving characteristics and further reduce the bending-induced stresses in the sandstone near the face. Horizontal fracs were created at a horizon about 25 feet above the Pittsburgh seam main bench over the two panels in June and November 1998, respectively. To evaluate the effect of sandstone fracing and operational adjustments on the bending-induced stresses in the sandstone, three finite element models were constructed and analyzed. Results from the finite element analyses indicate that one horizontal frac, 25 feet above the Pittsburgh seam, would reduce the bending-induced stresses in the sandstone by approximately 50 percent, maintaining them at a level below the ultimate strength of the sandstone. Operational adjustments to increase the support set load density and support stiffness could reduce the bending-induced stresses in the sandstone by a maximum of an additional 10 percent. Shield pressure monitoring at the B14 midface confirmed the presence of a lower roof pressure zone inby and outby the 1314 frac holes. This lower face pressure significantly contributed to improving the B13 and B14 longwall productivity and safety. The total production downtime due to bad top at the B 12 face within the interval of channel influence was 30,600 minutes compared to about 3,500 minutes each for the B13 and B14 faces. The average longwall advance per machine shift for the B13 and B14 faces within the channel was about 70 percent higher than that for the B 12 face.

Jan 1, 2001

By Stephen P. Signer

A method to assist in the evaluation and selection of roof bolts using in situ measurements of roof bolt loading has been developed by researchers of the Spokane Research Center, National Institute of Occupational Safety and Health. Both axial and bending forces are measured by strain gauges at many locations along the length of fully grouted bolts during various stages of mining. This information is then used to ( I) predict bolt loading for variations in bolt spacing, grade, and diameter, (2) calculate the design changes necessary in zones where bending loads are high, and (3) determine if bolt length is adequate. Results from several case studies of full-column bolt loading in coal mine gate roads are used to illustrate the design method. This knowledge will give design engineers a tool for the selection of roof support systems that will improve underground safety by reducing roof bolt failures.

Jan 1, 1997

By Antoni Kidybinski

Coal mining both with longwall and room-and-pillar method often encounters severe operational difficulties due to had roof conditions. Rooffalls at entry cross-cuts or along the rib line in longwalls as welt as dynamic and excessive loads on supports due to overhanging of strong roof layers in a goab - are well known phenomena to the mine operators. Also, a number of practical examples may be quoted of dramatic improvement of working conditions after changing of coal winning technology or supports type. It is therefore a question to be answered how far we are able to recognize roof strata properties before mining starts and adjust both coal extraction technology and supports of entries to rock stability foreseen. Furthermore, engineering principles and criteria for decision making should be established. In recent years several in situ rock strength and discontinuities detection testing techniques have been developed both indirect (geophysical) and direct thus making possible closer roof-strata examination. In this paper critical roof span criterion is discussed as a basis for selection of proper mining system in terms of desirable stability of mine opening .

Jan 1, 1982

By Ihsan B. Tulu, Keith A. Heasley, Aanand Nandula

"The objective of this paper is to present the development of a software tool that will assist in an area calculation of the Analysis of Roof Bolting Systems (ARBS) support intensity while incorporating more detailed geology, stress, and intersection span input. In this area ARBS tool, a detailed CMRR distribution can be determined from the available geologic database at the mine. In addition, the overburden and multiple-seam stresses, as obtained from the boundary-element program LaModel, can be converted to a “pseudo-depth” which is then used as the depth input to the area ARBS calculation. Finally, the section-specific intersection spans can be determined from the mine design for input to the calculation. All this detailed area input information for ARBS is handled by the latest version of the Stability Mapping (StabMap) program, which now outputs an area ARBS suggested support density given the required input parameters. This final ARBS grid can be plotted, analyzed and overlaid on the mine map for optimum use in support design and for presentation to production personnel.INTRODUCTIONSince the inception of roof bolts in the 1940s, there has been a lack of a universally accepted roof bolting design method (Mark, 2002; Mark and Pappas, 2012). To address this situation, the National Institute of Occupational Safety and Hazard (NIOSH) developed roof support design guidelines, which were incorporated into a computer program called the Analysis of Roof Bolting Systems (ARBS). These guidelines were based on the effects of roof quality, stress, and mine geometry (Mark, Molinda and Dolinar, 2001). The roof rock quality in ARBS is represented by the Coal Mine Roof Rating (CMRR), the depth is used as a surrogate for the stress, and the intersection span serves to represent the mine geometry (Mark, Molinda and Dolinar, 2001). These three input parameters are used in the fundamental, empirically derived equation of ARBS to calculate a required support density factor, called ARBS, (Mark, Molinda and Dolinar, 2001):"

Jan 1, 2018

By John Shepherd

Pillar extraction carried out using various methods (split and lift, Wongawilli and old Ben) involves the formation of narrow fenders (commonly 7 m of coal) formed by driving "splits" for lifting off with a continuous miner. The strata above a split and fender adjacent to the goaf consists of a cantilever beam (usually >3 m thick) that extends from the solid coal and bridges across the fender. Lifting off the fender from the split in safe conditions depends upon the bridging capability of the cantilever and the fender strength. The nature and competence of the immediate and near roof, the support capability of the coal fender and the presence of faults and or joints can adversely affect the bridging behaviour of the cantilever. Insufficient knowledge of these parameters can lead to extraction under dangerous conditions. Remnants of fenders ("stooks") also support the cantilever and their dimensions are critical to obtain regular caving and goaf falls. Stook area, goaf roof span and roof stand-up time after lifting can be related and only a relatively small increase in Btook area of 10-15 m2 can lead to a significantly increased stand-up time, especially under a massive roof. Such conditions strongly influence the redistribution of abutment stresses and can create increased outbye pillar loading. Based on detailed underground geotechnical investigations, it is concluded that strata caveability, roof cantilever behaviour, fender H:W ratio and remnant stook size are all critical parameters for successful and safe pillar extraction.

Jan 1, 1992

By Larry Howe

Second, or retreat mhng with Mobile Roof Supports (MRS) has now been part of coal mining in the United States for a decade Their utilization has evolved into mining applications which vary in seam heights from 38 inches to 170 inches, and overburden depths from outcrop to beyond 1200 feet This paper will detail the pillaring plans, lift sequences, and mnchky utilized in these operations and the rational for the plans A comparison of the current productivities achieved will be outlined

Jan 1, 1998

By C. D. Peters

The Bureau of Mines has been researching remote sensing analysis of an underground coal mining area in Utah and computer-aided methods of defining ground control hazards in this mining area. As a result of these studies, a preliminary potential hazards map was developed for the test site. This map is based primarily upon a lineament (linear feature) analysis of Landsat satellite data. Other data used were paleochannel location data, resistivity data, and information from surface and underground field checking of the lineaments and ground control problems in the mine. Lineaments proved to be reliable indicators of hazardous areas, although the type and degree of hazard varies with different lineament trends and the presence of other conditions such as lineament intersections and intersections with other geologic structures or with paleochannels. Lineaments were observed to match faults, joints, and/or joint trends in the area. The paleochannel data, derived from drill hole data, provided basic information about potentially hazardous areas in the mine by suggesting where paleochannels could cause instability in the roof rock. The ground control problem data served largely to define the relative degree of hazard associated with lineaments and the types of hazards encountered in the mine. Resistivity data helped confirm the presence of fractures where they were expected to occur both from surface geology and from the lineament data.

Jan 1, 1986

By Dion Pastars

Kestrel Coal has undertaken a surface-to-seam consolidation program of faulted ground prior to longwall retreat in the 304 panel. Micro fine cement is used to improve the rock mass quality and therefore reduce the risk to personnel and equipment whilst increasing previously assumed sterilised reserves. The program is based on information attained from previously successful grouting campaigns in the Queensland coal area and follows the decision making process to ensure there is sufficient spatial permeation of cement across the faulted area. This data is used to assess the residual risk to the business in terms of health, safety and production

Jan 1, 2007

By L. Z. Wojno

Tunnels in South African gold mines are developed at depths down to 3 600 m below surface where the virgin rock stress approaches 100 MPa and, on occasions, through rock where the field stresses exceed 150 MPa. In addition, many tunnels are subjected to increases in field stress of over 50 MPa during their useful lifetime. As a consequence, the walls of most tunnels are fractured causing the rock to dilate into the tunnel. Quantitative data on the extent of this fractured zone are presented. The function of support in these circumstances is to reinforce and maintain the integrity of the fractured rock. In 80 per cent of the 800 km of tunnels developed annually, the support methods used are capable of maintaining the stability of the tunnels. However, the remaining 20 per cent of tunnels is subjected to severer conditions where dilation in excess of 250 mm occurs and/or where the probability of strong seismic ground motion is high. In these tunnels, additional support is required. The type and pattern of support for both normal and severe conditions is described. In spite of the improved support used, damage to tunnels may occur. To minimize this damage, the concept of a rock reinforcing tendon that will yield and do work, when subjected to quasi-static or rapid deformation, has been put forward. Specifications for such a tendon are presented and the development of certain yielding tendons are described.

Jan 1, 1987

By Andre Zingano

The coal seam strength and the failure criteria that represent the behavior of the coal seam are very important for mine section and pillar design. Additionally, mine section and pillar design take into account roof and floor layers, which form the roof-pillar-floor system. Therefore, underground ground control design begins, most importantly, with rock mass characterization. Another issue is the distribution or behavior of the coal seam strength for the whole mine area. Sometimes, the coal seam strength is assumed to be the same for an entire area, but this is not true because the coal seam is heterogenic and anisotropic, and also the coal strength is affected by faults and spacing between fractures. The objective of this paper is to demonstrate the use of the methodology for geomechanics characterization to determine the strength for the Bonito coal seam, and discuss the coal seam strength distribution around the entire mine area. The intact coal strength was determined by laboratory tests that were completed during the last 17 years, including triaxial and uniaxial strength tests. It was noted that there is a strong correlation between all laboratory strength tests with 90% confidence interval. The explanation for this coal seam strength variation is the presence of the discontinuities and their variation in quantity and quality at each location in the mine area, because the coal seam rock classification and its behavior will be affected by the quantity and quality of fractures.

Jan 1, 2014

By Yoshiaki Fujii

Displacement Discontinuity Method (DDM), which was originally developed by Crouch,S.L., is a powerful tool to calculate mechanical disturbances due to tabular excavations. Three topics on DDM applications to strata control problems in Japanese deep coal mines are described. First, is numerical simulation of microseismicity induced by longwall mining in Horonai Coal Mine. Microseismicity in the mine had been monitored by a mine-wide seismic array. Location of microseismic events and intensity of seismicity were well simulated. In the simulation procedure based on DDM, first, zones where a failure condition is satisfied by the advance of mining are detected. Then, the seismic moments, which indicate the intensity of seismicity, at the failure zones are evaluated. Second, is prediction of the shaft damage due to mining in Sumitomo Akabira Colliery. In this mine, the limit angle of the shaft pillar was changed from 600 to 800 to increase the amount of minable coal volume without increasing mining depth. Displacement of the shaft axis has been measured to assess the damage due to mining around the shaft pillar. In the calculation based on DDM, a fault intersecting the shaft and the ground surface are considered. Observed and predicted displacement of the shaft axis agreed well. Third, is the roadway deformation in Taiheiyo Coal Mine. Attainable maximum stress for a roadway was calculated considering three faults intersecting the roadway. It was realized that the roadway's convergence is related to the calculated stress by simple equations.

Jan 1, 1992