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By Mark Colwell

This paper summarizes the results of a research project whose goal was to provide the Australian coal industry with a chain pillar design methodology readily usable by colliery staff. The project was primarily funded by the Australian Coal Association Research Program and further supported by several Australian longwall operations. The starting point or basis of the project was the Analysis of Longwall Pillar Stability (ALPS) methodology. ALPS was chosen because of its operational focus; it uses tailgate performance as the determining chain pillar design criterion rather than simply pillar stability. Furthermore, ALPS recognizes that several geotechnical and design factors, including (but not limited to) chain pillar stability, affect that performance. There are some geotechnical and mine layout differences between United States and Australian coalfields that required investigation and, therefore, calibration before the full benefits offered by the ALPS methodology could be realized in Australia. Ultimately, case history data were collected from 19 longwall mines representing approximately 60% of all Australian longwall operations. In addition, six monitoring sites incorporated an array of hydraulic stress cells to measure the change in vertical stress throughout the various phases of the longwall extraction cycle. The sites also incorporated extensometers to monitor roof and rib performance in response to the retreating longwall face. The study found strong relationships between the tailgate stability factor, the Coal Mine Roof Rating, and the installed level of primary support. The final outcome of the project is a chain pillar design methodology called Analysis of Longwall Tailgate Serviceability (ALTS). Guidelines for using ALTS are provided.

Jan 1, 1999

By Wenjun Ju

It's very important to determine the limitation of bed separation in entry roof in order to predict roofs stability and take some necessary precautions. Based on numerical modeling and monitoring in the field, the new concept of Limitation of Roof Layer Separation (LRLS) and its method of determination LRLS is given in this paper

Jan 1, 2001

By Houquan Zhang, Yu Wu, Yanlong Chen, Zhenfu Luo, Aihong Lu

"A non-destructive detection scheme for bolt installation quality was designed by adopting the method of non-destructive detection on bolt installation quality. Based on the field measurements, a random non-destructive detection method for determining the characteristic length including anchoring length, free segment length and initial axial load of bolt was discussed, and the characteristics of increasing bolt axial load in the initial stable stage was obtained. The results show that the bolt axial load and deformation of surrounding rocks correlate each other. Compared with the traditional detection technique for bolt installation quality, the non-destructive detection method has obvious advantages such as random, accuracy, and economic. The method provides a reliable basis for bolt support optimization and stability control of surrounding rocks.INTRODUCTIONBolt support in roadway can make the surrounding rocks and bolt form a structure in the early stage of roadway development through the bolt pre-tension, which can maximize the self-bearing capacity of surrounding rocks. Thus it is a form of active support (Kang, 2007). In order to make the bolt play an active supporting role, it needs to ensure the installation quality of bolting. Thus the bolts can supply a high initial anchoring force and control the deformation of surrounding rocks (Shan, 2009; Wei, 2012). Moreover, bolt supporting is a concealed work. Once installation quality problems appear; it will be difficult for effective detection and screening (Xie, 2008).THE PROBLEMS OF BOLT INSTALLATION QUALITY IN COAL MINE AND ITS DETECTIONThe Problems of the In Situ Bolt InstallationDue to the problems in installation environment, machine, and technique, and workers proficiency (Li, 2003), there are many problems exist in bolt installation. If the borehole depth is too long, the anchoring agent will accumulate at the bottom of the bore, which reduces the anchoring length. Furthermore, if the borehole depth is short due to the poor geological condition, the exposed part of the bolt will be too Jong and the pallet cannot cling to the rock face, which prevent the pre-tension from being applied. The pre-tension is very important for bolt installation. If pre-tension is insufficient the rock bolt cannot control the deformation of surrounding rock efficiently and the bolt anchoring force will decrease considerably. The bolt length is another important parameter for anchoring quality, which will affect the control area of bolt supporting. In addition, the cleanness of drill hole has a great influence on the effective anchorage length of bolt. If the drill hole cleaning is not enough, it will directly affect the anchoring agent and the contact area of surrounding rocks, which will decrease the effective anchorage length and initial anchor-force."

Jan 1, 2016

By Thomas R. Scotese

A rock mechanics instrumentation and monitoring program was implemented during pillar extraction at Gulf Mineral Resources' Mt. Taylor Mine, the deepest uranium mine in the U.S. Three types of monitoring were employed: (1) drift convergence around stopes (using a portable tube extensometer), (2) stress changes in pillars (using vibrating wire stressmeters in horizontal boreholes), and (3) load changes in haulage ways under stopes (using vibrating wire load cells in jack stands). Results include convergence-time graphs, convergence contour maps, stress-time graphs, a stress increase contour map, load-time graphs, and abutment load limit maps. Major factors influencing results include (1) size of pillars around the extraction area, and (2) presence and orientation of a major fault adjacent to one extraction area. These factors, combined with pillar extraction, produced evidence of load transfer over 120 m from a stope measuring only 55 m x 45 m. Vertical stress increases of 7 to 17 MPa (1000 to 2500 psi) within 60 m of the stope were obtained using the stress- meter manufacturer's method. The fault apparently prevented load transfer in the down-dip direction but increased load transfer in the up-dip direction. Convergence rates of up to 1.3 cm/day (0.5 in/day) were measured at two stations. Instruments were generally reliable despite adverse underground conditions. The instrumentation and monitoring program at Mt. Taylor provided a warning system against ground control problems and a characterization of ground behavior for future development and extraction of pillars. Recommendations are made for future mining and instrumentation practices.

Jan 1, 1984

By David Hill

Over the last decade, Beltana No.1 Mine has generally been Australia?s most productive longwall operation, with a ROM output of 7.2 Mt in YEJ2007. Characteristically favorable roof conditions and low depths of cover resulted in typically high rates of development drivage and longwall retreat. As a consequence of these high rates of development and extraction, there was a continuing focus on recovering and relocating the longwall face as efficiently as possible to minimize associated production delays and operational costs. This paper presents a comprehensive case study of the conventional longwall recoveries (?take-offs?) at the mine, from Longwall 1 to 14. By examining the take-off monitoring data collected over the life of the mine, the key geotechnical issues are demonstrated, including their relationship to the ground behavior experienced and the issues overcome.

Jan 1, 2011

By Daniel W. H. Su

Thick, massive sandstone present in close proximity of the mining horizon can create difficult longwall caving issues, which may lead to severe face loading and subsequent face roof control problems. Under deep cover, such delayed caving and overhang not only impose additional loading on the gate road pillar system, but they also increase the potential for large seismic events. Such a combination of thick, massive sandstone geology and deep cover was present over a portion of a southwestern Virginia coal mine, which was confirmed via detailed in-mine roof scoping and mapping. Large seismic events were recorded over a district of 1,000-foot-wide longwall panels. This paper presents the ground control design changes implemented to reduce longwall-induced stresses in the sandstone, to reduce longwall abutment pressures in the gate road pillars, and to reduce the magnitude of mining-induced seismic events. A series of two-dimensional finite element models were constructed and analyzed to evaluate the longwall-induced stresses in the sandstone above the gate road pillars and the abutment pressures within the gate road pillars. Results from the analyses indicated that reducing the panel width from 1,000 to 700 feet reduced the longwall-induced stresses in the sandstone by a factor of 2.0. Specifically the probability of sandstone stability improved from 29% for the 1,000-foot-wide panel district to over 95% for the 700-foot-wide panel districts. Also, the subcritical 700-foot-wide panel, coupled with a change in pillar design, considerably reduced the longwall-induced abutment pressures in the gate road pillars, which significantly increased the gate road pillar stability factor. Surface subsidence measurements conducted over the 1,000-foot-wide panel district and two new 700-foot-wide panel districts were in very good agreement with those predicted by the finite element models. In addition, results from the models indicated that wider panels with a smaller district may produce the same probability of sandstone stability. One four-panel district and one five-panel district with the new ground control design changes have been mined successfully. The panel width was found to be the most influential factor in determining the longwall-induced stresses in the sandstone and in the gate road pillars. The thickness and massive nature of the sandstone, the proximity of the sandstone, and the strength of the sandstone were also found to be important factors. Seismic monitoring over the two mining districts that employed 700-foot-wide panels confirmed the reduction of one order of seismic magnitude when compared with those measured over the 1,000-foot-wide panel district.

Jan 1, 2014

By Hamid Maleki

The advance-and-relieve method benefits from lateral destressing associated with mining in laminated rocks and a high horizontal stress regime. This stress control method is based on measurements showing that occurrence of rock failure in the roof and floor of an entry (or caving of roof strata in a panel gob) results in redistribution of stresses in adjacent entries. By locating other entries within the shadow zone of the first entry (or gob), improvements in stability can be achieved. Numerical modeling proved useful in studying the basic mechanics of lateral relief while investigating the sensitivity of results to different geologic and mining parameters using controlled experiments. It was shown that failure of rocks near an entry results in redistribution of horizontal stress and shifting of the stress to higher horizons. Measurements from two mines are consistent in showing significant horizontal stress reductions in comparison with the far-field stress regime within the destressed zone. Although the far-field stress regime is very anisotropic, these measurements show near-equal secondary principal horizontal stresses, or perhaps a switch in orientation, as a result of destressing. Stress relief is achieved through lateral movement and relaxation of rocks along weak bedding planes toward adjacent caved zones (or softened zone). Because of cave geometry in the advancing panel, horizontal stress concentrations occur near the cave line both in front of the face and to the sides. The horizontal stress concentration reaches 1.7 times the far-field stress ahead of the face. This stress increase is significant and may cause structural damage in this zone benefiting from additional support. In the next advancing panel located within the shadow zone of the gob, horizontal stresses are significantly reduced in the roof (by 50%). Thus, the stability of future advancing panel can be improved through prudent layout designs and sequencing. The width of the relief zone is significantly influenced by the height of the softened (or cave zone) and rock mass properties.

Jan 1, 2003

By Wei Wu

A new bench blast design method and its realization with the assistance of the computer is introduced. The basic design principle is to define a rhomb cell, called fully-fractured zone, through a series of full scale bench crater tests and apply this geometric pattern in multi-hole, multi-row blast design. To illustrate the design method, the optimization process and assessment of blast performance are presented in the context of a case study.

Jan 1, 1992

By Olayemi Akinkugbe

As coal production in the United States continues to increase, the availability of unexploited or virgin fields continue to diminish. As a result, mine operators are forced to mine in unfavorable or multiple seam conditions. The safe, productive exploitation of coal in multiple-scam conditions requires specific design technology. At present, there are two general design approaches that the mining engineer can use to analyze the multiple-seam problem in question, either simple empirical relationships or complex numerical models. Often, the empirical approaches are too general for the specific problem and the numerical models require too much time and expertise for the practicing engineer. This paper introduces a new computer program, Lam2D, designed specifically to enhance multiple seam mining operation by quickly and easily calculating, stress, displacement, subsidence, horizontal strain and safety factors associated with multiple seam mining situations. The program implements a simplified, two-dimensional boundary-element method in order to model the complex multiple-seam stress and displacement interactions. Most importantly, the program is greatly simplified by incorporating automatic overburden, interburden, coal and gob property generation thereby reducing typical modeling and solution time to a few minutes.

Jan 1, 2004

By Russell C. Frith

Risk-based roadway roof support design is now a critical part of the Australian Coal Industry. Safe and efficient mining demands that roof support be tailored to the prevailing geotechnical conditions and coal mining legislation in New South Wales, Queensland, and Tasmania is clear in requiring that formalized roof support design be undertaken. The assumption of dead-load suspension of an otherwise unstable roof has been used in roadway roof support design for many years. However longevity does not necessarily translate into either a best or even reasonable practice in the current industry, particularly when pro-active strata management systems are now routinely used and reinforcing support design methods are available. The paper discusses why the assumption of dead-load suspension is fundamentally incorrect in almost all instances when pro-active or reinforcing roof support is applied and how it can easily result in misleading and potentially under-designed roof support systems under various circumstances. The commonly used design methodology of balancing the installed axial capacity of long tendons with the assumed ?weight? of a future roof fall is summarized, and several fundamental flaws are identified. The critical area of selecting the design Factor of Safety is also discussed in detail. Accepting that under certain circumstances the roof of a roadway will require it to be ?suspended?, suggestions are given for how a robust support system can be developed, designed and applied.

Jan 1, 2011

By Terry Medhurst

Geophysical logging is routinely undertaken as part of most Australian coal mining exploration programs. Typically, the main application is the determination of coal seam depth and the qualitative estimation of coal quality, lithology and rock strength. For rock strength estimation, sonic transit times relate to the elastic properties of the rock and provide a general correlation with uniaxial compressive strength. However, experience has shown that site-specific relationships are necessary to account for different geological environments. In this paper we describe results from a series of recent studies that have provided a more comprehensive method of geotechnical strata characterisation using detailed lithological interpretation and rock property estimation from borehole logs. This has resulted in an improved method of estimating rock strength via the determination of the proportions of quartz and clay plus the porosity. We also propose a rock classification scheme based on geophysical log analysis. This builds on the existing acceptance of the sonic log as an indicator of rock strength and adds scores according to the moisture, cohesion, porosity and quartz content inferred from the geophysical logs. The new scheme can be related to existing mapping based schemes such as CMRR and the Q-system.

Jan 1, 2005

By E. Bane Kroeger

For many years, researchers around the world have been investigating coal pillar stability. Many have focused on trying to optimize the size of the pillars by examining stable and failed pillars in underground coal mines. For mines that are already in operation or companies that are opening a new mine adjacent to an existing mine where the pillar dimensions and stability can easily be determined, this is the preferred technique. However, when a new mine is planned for an area where there has been relatively little mining, then samples of the coal need to be drilled and tested for strength. Because of the fractured nature of coal, this lab testing is often problematic and may often yield poor results because of the limited size and number of samples that can be obtained from the core. To remedy this problem, research was conducted in the labs at SIUC to determine if improvements could be made when optimizing the size of pillars for underground coal mining in Illinois. Through a greater understanding of the relationship between dimensions and strengths of coal samples tested in the lab, it was hoped the in situ strength of coal pillars could be determined with greater accuracy. This could allow the extraction ratio to be increased and more of the coal resources in Illinois recovered without impacting the safety of the mine workers. Many different agencies around the world have published recommended minimum numbers of samples for strength determination based on the size of the coal samples. For two inch cubes, uniaxial testing of at least seven samples is recommended and this minimum number decreases to four when testing four inch cubes. However, the laboratory testing of coal from two seams thus far has suggested this number should be almost doubled to accurately determine the coal strength. Testing also showed there is a pronounced decrease in the variation in the strength of the samples tested as the dimensions of the sample were increased. From the testing results thus far, it is recommended that a sample size versus strength curve be constructed for every coal seam for Illinois. Creating such a curve will better determine both the minimum number of samples and minimum sample dimensions for that seam or region of a particular scam. A second set of tests were conducted in the laboratory to determine the increase in strength with reduction in height for samples with the same width and length. In Illinois, typical pillar dimensions are 40 to 80 feet (12.2 to 24.4 m) wide and long, with coal thickness varying from about 4 to 9 feet (1.22 to 2.74 m). This provides width/height ratios ranging from 4.5 to 17.5. Coal samples having width/height ratios ranging from 0.5 to 13 have been tested thus far. As with most pillar design formulae, the strength of the Murphysboro coal versus width/height ratio was linear and closely matched shape factors published by other researchers. The laboratory testing thus far has identified three factors that affect the strength of laboratory samples, the number of discontinuities, the angles of critical discontinuities, and the shape of the samples. The factor that probably had the most pronounced effect on the strength of the coal samples was the angle of critical discontinuities. This testing has confirmed that there is still a gap that needs to be bridged between the strength obtained from laboratory testing and the in situ coal strength predicted.

Jan 1, 2004

By D. Newman

An underground limestone quarry, developed in the 1940's, transitioned from a surface quarry to an underground room-and-pillar operation. The quarry expanded to a second underground level when the reserves on the first level reached the property boundaries. The historical basis for the sill pillar thickness between the first and second level, pillar dimensions, and mining height was lost during the course of several changes in ownership. When reserves on the second level reached a five-year life, and plans were initiated for a third level, quarry management embarked upon a thorough evaluation of i. current ground control practices for the support of the immediate roof, ii. roof support in the declines between the surface, Level 1, and Level II, iii. pillar centers and pillar safety factors, iv. pillar alignment between levels, v. mining height within a level, and vi. the locations of declines to access future reserves on level three. The evaluation began with detailed mapping of the location and orientation of calcite veins, joints, fractures, and ground conditions in the immediate roof of Level II. The orientation and location of calcite veins, joints, and fractures on Level II were compared with similar mapping on Level 1, completed in the 1970's to establish whether these features were continuous and would be expected to affect conditions on Level III. Rock strength, physical properties, and fracture spacing were obtained from two (2) diamond core holes drilled from Level I to Level II and two (2) holes drilled from Level II to define the vertical extent of future reserves on Level III. The effort to characterize geological conditions was expanded to a rock mechanics investigation of the: i. performance of the current roof support, ii. optimization of roof and rib scaling, iii. calculation of safety factors for 416 pillars on Level II, iv. evaluation of future pillar centers and mining height within a level based upon pillar strength and pillar geometry, v. development of beam safety factors for the immediate roof on Level II, vi. identification of the most suitable strata for the roof in the future level three, and vii. use of LAMODEL to develop a large numerical model of overburden, inter-level, and total stress on the pillars in levels one and Level II with a stress "footprint" on the proposed level three. The characterization of engineering parameters and geological conditions resulted in changes to the: ? active operations in Level II regarding scaling ? equipment and scaling practices, ? pillar centers, ? mining height, ? raising the immediate roof horizon to avoid ? laminated strata in the current roof, and ? a change in rock bolts from a friction based system ? to active tensioned rebar. The planning and design of level three is ongoing and relies heavily upon the geological and engineering investigations conducted during the past year and one-half.

Jan 1, 2001

By J. M. Girard

Ground control problems at surface mining operations can occur for a variety of reasons. Stress, gravity loading, rock strength, geology, pore pressure, weather effects, underground workings, and many other factors contribute to slope instabilities that range from small rock falls to massive slides of material. While some of these failures can he predicted or controlled by preventive measures, each year many completely unexpected failures occur. Current methods for monitoring generally involve measuring displacements at a few, selected points in and around the suspected area of instability. While most of the displacements along these points will be in a downslope direction, freeze-thaw cycles of water-filled joints, horizontal stresses or pressure, buoyancy in saturated soils, human measurement errors, or other situations can produce deformation in almost any direction, even without any instability in the slope. Determination of which, if any of the observed movements represents a potential hazard is essential.

Jan 1, 1998

By Paul C. Hagan, Jianhang Chen, Serkan Saydam

"Modified cable bolts are commonly used in underground mines due to their superior performance in preventing bed separation when compared with plain strands. To better test the axial performance of a wide range of cable bolts, a new Laboratory Short Encapsulation Pull Test (LSEPT) facility was developed. The facility simulates the interaction between cable bolts and surrounding rock mass, using artificial rock cylinders with a diameter of 300 mm in which the cable bolt is grouted. Furthermore, the joint where the load is applied is left unconstrained to allow shear slippage at the cable/grout or grout/ rock interface. Based on this apparatus, a series of pull tests were undertaken using the MW9 modified bulb cable bolt. Various parameters including embedment length, test material strength and borehole size were evaluated. It was found that within a limited range of 360 mm, there is a linear relationship between the maximum bearing capacity of the cable bolt and embedment length. Beyond 360 mm, the peak capacity continues to rise but with a much lower slope. When the MW9 cable bolt was grouted in a weak test material, failure always took place along the grout/ rock interface. Interestingly, increasing the borehole diameter from 42 mm to 52 mm in weak test material altered the failure mode from grout/rock interface to cable/grout interface and improved the performance in terms of both peak and residual capacity.INTRODUCTIONFully grouted cable bolts have been used in underground mining industry for many decades. Initially, plain cable bolts manufactured by winding seven flexible steel wires together were commonly used (Thorne and Muller, 1964). Laboratory and field tests lead to a better understanding of the transfer process between grouted cable bolts and the surrounding rock mass, which is beneficial to improve the reinforcing performance of cable bolts in engineering practice (Fuller and Cox, 1975; Cox and Fuller, 1977; Schmuck, 1979; Stillborg, 1984). Nevertheless, as in many cases the initial cable was sourced from discarded steel with a smooth surface along the strand, plain cable bolts did not always meet performance expectations (Windsor, 1992). Over time, it was realised that cable surface geometry plays an important role in determining anchorage performance. Consequently, modified forms of cable bolts were developed. For instance, twin-strands which were made by using spacers to combine two plain cables together improved anchorage capacity two-fold (Matthews, Tillmann and Worotnicki, 1983). However, this required a large borehole to be drilled. Epoxycoated strands had remarkable performance in practice but were difficult to install without damaging the epoxy layer (Dorsten, Hunt and Kent, 1984). In the 1990s, there was a rapid development in the cable bolt design. A typical example was birdcaged cable bolts. They were initially suggested by Hutchins et al. (1990) via unwrapping traditional seven-wire strands and generating a number of nodes together with antinodes along the strand. But it was difficult to make birdcaged cable bolts that could be coiled for transporting."

Jan 1, 2015

By John C. Stankus

At the 19th International Conference on Ground Control in Mining, a paper was presented detailing new types of Jennmar INSTáL tensionable cable bolts (Stankus and Taylor, 2000). Since that time, significant testing and trials of the various types of tensionable cable bolts have been conducted. Several MSHA districts and state regulatory agencies have now approved these systems as primary support This paper will detail the use of tensionable cable bolts as primary support at two longwall mines in Virginia (MSHA District 5).

Jan 1, 2001

By Bruce K. Hebblewhite

BHP Billiton Illawarra Coal operates Dendrobium Mine in an area 10-20km west-northwest of Wollongong in the Southern Coalfield of New South Wales, Australia. The mine recently completed longwall mining in Area 3A adjacent to an overhanging rock feature known as Sandy Creek Waterfall, which had been identified as a significant, and potentially sensitive natural surface feature of the environment. This waterfall lies within the Sydney drinking water catchment area, in a surface terrain of variable topography where non-conventional subsidence behaviour such as valley closure, and far-field horizontal movements are known to occur, over and above conventional subsidence responses to mining. As a part of the conditions of approval to mine, Illawarra Coal undertook to protect the waterfall and the section of Sandy Creek immediately upstream of the waterfall from the effects of longwall mining using an innovative management process and an array of very high resolution monitoring systems. This paper outlines the type of non-conventional subsidence effects that can occur; the range of different monitoring regimes adopted in response to such effects in order to develop an understanding of the valley closure phenomenon in proximity to the waterfall and provide suitable early warning data. The paper also describes the management processes that were adopted, incorporating the use of these monitoring regimes in order to successfully protect the waterfall from the valley closure effects of mining four adjacent longwall panels in close proximity to the waterfall while continuing to maximise recovery of the coal resource in the area. The management structure adopted involved a technical committee, a steering committee, and an external independent reviewer. The technical committee was responsible for design of the monitoring systems, interpretation of the results, and making timely recommendations to the company steering committee who were then able to make informed decisions on when to cease mining each adjacent panel. The steering committee comprised Illawarra Coal management and technical personnel. Although the steering committee took advice from the technical committee, all decisions relating to mining were made by the steering committee. An external reviewer was engaged by the steering committee at the end of each longwall panel to review the results, interpretation and management decisions. This management structure was effective in successfully protecting the very sensitive structure of Sandy Creek Waterfall from potential impacts of nearby mining in the midst of active ongoing natural erosion processes.

Jan 1, 2014

By Amirhossein Molavi Tabrizi

In this paper we introduce a realistic model for mine openings that explicitly considers the effects of local irregular topography. Also, we consider the effect of entry orientation and rock anisotropy on the induced stress fields around mine openings under the given in-situ stress field. A triangular mesh formulation is derived for representing complicated irregular surface topography and underground openings. Each element contains three internal collocation nodes. Parts of the surface which are far from the mine openings are discretized by infinite elements. The rectangular infinite elements with three internal nodes are introduced specifically for this problem so they are compatible with the triangular elements. With the corresponding boundary element method (BEM) formulation, we then investigate the combined effect of the topography and entry orientation on the stress fields induced by openings in anisotropic rocks under in-situ stresses. A single rectangular opening in the anisotropic ground with topography and in-situ stresses is simulated using data from the Carroll Hollow mine in Ohio. The results of these simulations show that topography, rock anisotropy, orientation of the regional stress field and mine opening geometry all have appreciable effects on the stress fields around underground openings, yet these factors can all be accounted for in an efficient code with no need to discretize the model domain. Future applications using this new BEM formulation will provide useful guidance on selecting the optimal opening orientation in the mine in order to minimize undesirable stress concentrations around underground mine workings.

Jan 1, 2014

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 Guy Reed, Russell Frith

"Current coal pillar design is the epitome of suspension design. A defined weight of potentially unstable overburden material is estimated, and the dimensions of the pillars left behind are based on holding up that material to a prescribed or user-defined Factor of Safety. In principle, this is seemingly no different to early roadway roof support design. However, for the most part, roadway roof stabilisation has progressed to reinforcement, whereby the roof strata is assisted in supporting itself. This is now the mainstay of efficient and effective underground coal production. Suspension and reinforcement are fundamentally different in their roadway roof stabilisation approach and, importantly, lead to substantially different requirements in terms of roof support hardware characteristics and their application. In suspension design, the primary focus is the total load-bearing capacity of the installed support to ensure that it is securely anchored outside of the potentially unstable roof mass. In contrast, reinforcement recognises that roof de-stabilisation is a gradational process with an ever-increasing roof displacement magnitude leading to ever-reducing stability. In a reinforcing situation, key roof support characteristics relate such design issues as system stiffness, the location and pattern of support elements within the roadway, and mobilising a defined thickness of the immediate roof to create (or build) some form of stabilising strata beam. The objective is to ensure that horizontal stress acts across the roof of the roadway and is maintained at a level that prevents mass roof collapse. This paper presents a prototype coal pillar and overburden system representation where reinforcement, rather than suspension, of the overburden is the stabilising mechanism via the action of in situ horizontal stresses within the overburden, the suspension problem potentially being an exception rather than the rule, as is also the case in roadway roof stability. Established principles relating to roadway roof reinforcement can potentially be applied to coal pillar design under this representation. The merit of this assertion is evaluated according to documented failed pillar cases in a range of mining applications and industries found in a series of published databases. Based on the various findings, a series of coal pillar system design considerations and suggestions for bord and pillar type mine workings are provided. This potentially allows a more flexible and informed approach to coal pillar sizing within workable mining layouts, as compared to common industry practices of a single design Factor of Safety (FoS) under defined overburden dead-loading to the exclusion of other potentially relevant overburden stabilising influences."

Jan 1, 2017