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By Heather E. Lawson

Recent field data providing measured pillar loading from a deep western mine revealed that an increased load transfer distance existed compared to expected values (Larson et al., 2012). Additionally, recent work by Tulu and Heasley (2012) has shown that the abutment angle in deep longwall mines can be significantly reduced as the depth of mining increases. The load transfer distance and abutment angle are both used as parameters to estimate pillar loading in longwall design. To investigate the potential implications of these observations on pillar loading in deep longwall mines, a parametric study using ALPS loading estimation procedures was performed. A supplemental study was conducted using LaModel to further investigate pillar loading at the tailgate. The study found that the abutment angle is particularly important for panels with supercritical geometries. As panels reach a subcritical geometry with increasing depth, the primary influence of the abutment angle is reduced and matched by other factors, including load transfer distance for tailgate loading. An extended load transfer distance was shown to potentially reduce loading on pillars adjacent to the longwall as the loading is spread over a larger area. The results of these studies indicate a need for further field verification and research toward regional geologic influences on pillar loading and how these may impact ground stability.

Jan 1, 2013

By Kirk W. McCabe

RokLok binder is a two-component polyurethane system consisting of a polymeric isocyanate (Component A) and a polyol resin (Component B). The two chemicals are mixed and injected into the mine rock under pressure. The initial low viscosity enables the RokLok material to permeate many of the smallest fractures. About 3-5 minutes after injection, the reacting material begins to expand in the fractures; the reaction is completed after 1-2 hours, and the material sets to a mechanically strong binder which adheres tightly to the rock. The consolidated rack provides resistance to any further f movement w thin the strata. The RokLok binder injection system is fast becoming an important engineering tool for coping with adverse underground geological conditions. Two case histories discussed in this paper, a longwall mining problem and an entry-roof problem, demonstrate the effectiveness of using the RokLok binder system to help solve complex ground-control problems in a cost-effective manner.

Jan 1, 1981

By WenFeng Li

A 3D roof bolt model with the consideration of rebar, resin, bearing plate, and resin/rock interface was developed for studying the bolt/rock interactions of the tensioned and fully-grouted resin bolts. For the tensioned bolts, 2 ft compressive zone above the rooflineis generated by the 10 tons pretension. Since the free portion of the tensioned bolts displays a higher vertical stress, the increase of the bearing capacity of the rocks in such range is much larger. For the fully grouted resin bolts, despite no pretension is generated during installation, the roof sag can indirectly generate a compressive zone near the roof line. In addition, the analysis on the differences in bolt/rock interactions when employing the tensioned or fully-grouted resin bolts indicates that the fully-grouted resin bolts are preferred as the primary support when weak and thinly-bedded immediate-roof is present. Meanwhile, it is found that both the tensioned and fully-grouted resin bolts may have very limited effect in preventing and/or control the cutter roof.

Jan 1, 2014

By Zbigniew Lubosik, Andrzej Walentek, Stanislaw Prusek, Sylwester Rajwa, Aleksander Wrana

"Underground coal gasification (UCG) provides a solution to supply electricity from high ash coal or deep deposited seams. The coal is consumed in situ, and, accordingly, the ash is left underground. UCG could play a major role in the future of coal, which is expected to fuel 44% of the world’s energy in 2030, in a more environmentally friendly and commercially viable manner. Nevertheless, it is known that the UCG can impact the environment. One of the most important factors in safe UCG process should be the evaluation of rock mass quality and the monitoring of its behaviour. The changes of rock mass parameters and quality have an influence on numerous aspects, such as subsidence and the stability of underground workings in georeactor vicinity, however, they also increase the possibility of pollutant migration through fractures and pores. These issues are especially important in cases of gasification conducted in an operating mine or on areas of high ecological vulnerability. This paper details observations made in two underground trials of UCG performed in Poland recently. The first of which is the Experimental Mine Barbara (under the project entitled HUGE-2 Hydrogen oriented underground coal gasification for Europe – Environmental and Safety aspects, (Wiatowski et al., 2015). The other trial has been conducted at an operating mine, Wieczorek, which is in the framework of a research project on the development of coal gasification technology for highly efficient fuel and electricity production (Czaplicka-Kolarz et al., 2013). During these trials, investigations were carried out aimed at assessing geomechanical parameters of rock mass and support behaviour in surroundings of UCG geaoreactors. Investigation comprises assessment of geomechanical parameters of coal, floor and roof strata, roof displacements, and concrete support strength parameters variation, as well as standing steel support load. Hydraulic penetrometer, endoscopic camera, Schmidt hammer, pull-out method, telltale, and ground penetration radar (GPR) were used in the investigation. Research conducted at both EM Barbara and Wieczorek Collieries revealed that the UCG process did not adversely affected the competency of rock mass (expressed by number of fissures and strength parameters of rocks), the stability of roadways and the support performance in the areas located in the vicinity of the georeactors."

Jan 1, 2017

By Michael Murphy, Khaled M. Mohamed, Berk Tulu

"Failures of coal ribs are major hazards in underground coal mines. For the period of 2010 to 2014, rib falls were the leading cause of groundfall fatalities and accounted for 35% of all groundfall-related fatalities. Currently, the National Institute for Occupational Safety and Health (NIOSH) is conducting a research project to develop an engineering-based design methodology for rib control in underground coal mines to reduce rib fall accidents and fatalities. Numerical models can be used as part of the design methodology provided the models are appropriately calibrated and verified.This paper presents a coal rib model to simulate the mechanism of coal deformation and fracturing that can lead to stress-driven rib falls. The proposed coal rib model was calibrated against published data from an instrumented rib site at the West Cliff Mine in Australia. Fracturing of the coal rib was defined by critical VALUES (of plastic shear strain of the coal material and rib displacement. A critical plastic shear strain of 2.8% and a sudden increase of rib displacement of 30 to 45 mm were sufficient to develop a fracture plane in the rib model. Accordingly, a discontinuous rib deformation was achieved and it compared favorably against field monitoring data.INTRODUCTIONCoal ribs are the vertical walls of coal formed when entries are developed in a coal bed. Failures of coal ribs are a major hazard in underground coal mines. Furthermore, failing ribs can contribute indirectly to roof and floor instability by increasing opening widths across intersections and entryways. Over the last 18 years, there has been a continual occurrence of injuries and fatalities in underground coal mines due to rib falls. For the period of 2010 to 2014, rib falls (excluding longwall face failure and bursts) were the leading cause of ground fall fatalities and accounted for 35% of all groundfall-related fatalities (MSHA, 2015)."

Jan 1, 2016

By Xicai Gao

With the intensification and expansion of deep coal mining, the surrounding rock stress environment becomes further deteriorated because of the complicated geological environment and powerful mining effect, so the major deformation of roadway appeared frequently. The proportions of floor heave are increasing and , the most serious problem is the floor heave of preparatory workings. The coal seam of Binchang coal area in Shaanxi province is at a depth of 621.6 ~ 817.0 m, the average thickness of the coal seam is 25 m, and the overburden and coal measure strata are the main regional aquifers and are highly permeable. The properties of water-weakening are obvious in the floor strata. The main preparatory workings or chambers were driven in the coal seam. The severe floor heave of the water chamber appeared during early mining and the amount was 180 ~ 910mm, and seriously affected the normal installation and use of the water chamber. Theoretical study and numerical simulation as well as the in-situ measurements were carried out to investigate the deformation and failure characteristic of the surrounding rock with non-support in the floor and the main reasons which caused the floor heave were determined. A new combined support technology was proposed which consists of overcutting, grouting, backfilling and invertedarch and bolt-grouting. The field monitoring showed that the cumulative convergence of the water chamber was 20mm, the maximum convergence speed is 1.2 ~ 2mm/d, the accumulative convergence of two ribs was 36mm and the maximum convergence speed is 2mm/d. The combined support technology will enhance the support intensity between the roof and floor, the two sides of the entry and improve the loading capacity of the floor effectively and the floor heave can be controlled effectively.

Jan 1, 2013

By Michael Karmis

During the last 25 years, technological advancement and subsidence research have resulted in more accurate and diverse prediction capabilities. The work presented in this paper focuses on the development of integrated prediction capabilities for dynamic deformation indices and for strains over sloping terrain. In order to assist subsidence engineers with accurate prediction methodologies, the research also addresses the development and validation of techniques that can enable model calibration using alternative measured subsidence parameters such as horizontal or ground strain instead of the traditional vertical subsidence values. This paper presents the basic concepts and validation of the above mentioned enhanced prediction and control methodologies, which are necessary for effective assessment and control of mining-induced subsidence, using examples and case studies. In addition, a risk assessment approach for evaluating landscape stability over mined areas is presented. The enhanced prediction methodologies have been incorporated into the Surface Deformation Prediction Software (SDPS) package.

Jan 1, 2008

By Raymond M. Stateham

The Bureau of Mines has developed a portable, lightweight device for determining the integrity of resin-grouted roof bolts. The instrument functions by sending a known pulse of ulstrasonic energy into the bolt and comparing the amount of energy reflected hack with the original measured pulse. Where the bolt is properly bonded or coupled with the rock, the energy passes through the bolt-resin-rock interface and dissipates in the rock. Where the bolt is not properly coupled, the energy is reflected back along the bolt to the instrument. The ratio of reflected energy to the pulse approximates the fraction of the bolt's surface that is improperly bonded. The instrument is easy to use. The operator holds the sensor head of the device firmly over the head of the bolt to be tested, causing the pulse of ultrasonic energy to be imparted to the bolt. The device then listens for reflected energy. This cycle is repeated about 200 times per second until an evaluation of the bolt installation is made and indicated by colored lights in the display window. Bolts having 50% or less of their surface bonded are considered had and cause a red light to he illuminated. Bolts with 50% to 75% of their surface bonded cause a yellow light to be lit and are considered marginal. Bolts having more than 75% of their surface bonded are considered good and cause a green light to be lit. The indicator lamp remains on for about 4 seconds and then goes off, indicating that the instrument is ready for another test.

Jan 1, 1982

By James H. Fletcher

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

Jan 1, 1982

By R. Karl Zipf

Stress analysis programs such as MULSIM/NL, LAMODEL, MinSim 2000, and EXAMINE TAB are used in the mining industry to analyze stresses and displacements in coal mines, platinum mines, gold reefs, and tabular-type deposits. These relatively simple numerical models can efficiently simulate yielding and failure of a rock mass near a mine opening and subsequent stress transfer. The main input parameters for these models are a family of nonlinear stress-strain curves for the in-seam material. For many applications of these models, such as in coal mining, extensive field measurements and observations exist from which to estimate these stress-strain curves and calibrate the numerical models; however, such measurements are lacking for other types of mines. A three-step method is presented to determine nonlinear stress¬strain curves for boundary-element (BE) programs used in many mining applications. The method requires a suite of laboratory¬scale strength tests at various confining pressures. Note that this analysis uses laboratory-measured strength and modulus properties of alluvium. The dependency of the mechanical properties of alluvium on scale has not been established, although more tests are being planned. However, it is believed that the relative comparisons performed are valuable regardless of scale effects. First, the FLAC2D computer program is used to model the laboratory tests and determine cohesion (c) and friction angle (0) for a Mohr-Coulomb material model. Second, another FLAC2D model uses the c and 0 values to calculate vertical stress and strain around a single opening in the rock mass. This model calculates the stress-strain path of points at various distances from the opening boundary. Finally, based on these stress-strain paths, the stress-strain curves needed for a BE analysis are derived. This three-step method was demonstrated in a large, flat-lying underground test facility in very weak rock. The BE analyses agreed well with observations of failure at the test facility and provided a basis for evaluating the behavior of alternative layouts.

Jan 1, 2003

Design and selection of support systems for longwall faces call for in-depth knowledge of strata mechanics and in particular of strata-support interactions. The paper presents the results of field investigations and analysis of a difficult-to-cave roof using powered supports under a massive dolerite sill in Central India using a state-of-the-art datalogger system. The mechanised longwall face was equipped with 4/680 tons chockshields operated on the IFS mode using a 300 K W DERD shearer. The datalogger system was capable of storing data simultaneously in 32 channels. The paper describes the monitoring system and the monitoring programme. The system provided useful information on support behaviour, build-up of leg pressures and face convergence during an intense dynamic periodic weighting phenomenon which destroyed the integrity of almost half of the face supports. At this point, the face had advanced by about 197 metres from the barrier. The paper outlines a post-mortem analysis of this dynamic event and provides a critical evaluation of the data collected for understanding strata support interaction and the behaviour of the massive sill in the overburden.

Jan 1, 1992

By André Zingano

Pillar collapses have been studied for several years and can be classified into two types: nonviolent squeeze or violent pillar collapse, i.e., controlled or uncontrolled pillar collapse. Underground coal mines in Brazil are considered shallow mines, because the overburden thickness varies between 20 and 400 meters. For this reason, the coal pillars should be designed as permanent pillars to avoid subsidence and groundwater problems. In the last two years, various pillar collapses occurred, one being a violent collapse and the others being squeeze collapses. All these collapses were considered massive pillar collapses because many pillars were involved. In the violent pillar collapse case, the width¬to-height (w/h) ratio was less than three; although, the pillar safety factor (SF) was more than 1.3. There was a collapse of about 100 pillars in less than three hours. The objective of this study is to determine the mechanism of pillar collapse for this violent pillar collapse. Convergence monitoring and numerical modeling were applied to combine field data and theory to simulate the pillar collapse. Convergence field measurements introduce the arc-effect behavior for the roof. The results showed that other factors besides pillar geometry caused the pillar failure. Other parameters, mainly dip of the coal seam and presence of fractures and faults may affect the strength and failure mechanism of a pillar.

Jan 1, 2004

By Jun Lu, Greg Hasenfus

"In this paper, the mining experience and challenges for the first right-handed longwall panel in the Pittsburgh seam are introduced. The longwall headgate T-junction experienced very high face convergence (up to 2 feet), accompanied by roof sag, floor heave, and rib loading. The headgate convergence was so large that, in a few places, it threatened longwall retreat and ultimately required the bottom to be re-graded.Different underground instruments, such as a roof scope, de-gas drill, tell-tale, laser meter, and Borehole Pressure Cell (BPC), were employed to explore the roof geology and to monitor the entry convergence and the stress changes in the pillar. In addition, the impact of other geologic factors, such as large overburden depth, laminated sandstone roof geology, soft floor, and large headgate equipment, were also analyzed. Subsequently, geotechnical solutions were provided to avoid or mitigate the impact of these challenging geologic factors.INTRODUCTIONIt is well known that high horizontal stress will affect the stability of roof, floor, and ribs within underground mines. In general, the headgate entries will experience stress concentration for the right-handed longwall face. On the contrary, the stress concentration will be in the tailgate entries for the left-handed longwall face (Hasenfus and Su, 2006). Without the stress shadow, the stress concentration in the headgate for the first longwall panel may be more significant (Figure 1).A mine in the CONSOL Pennsylvania Operation was extracting the first right-handed longwall panel in the Pittsburgh seam, which created high horizontal stress concentrations at the longwall headgate. Because of concern over soft wet floor within the syncline, floor coal had been left in place on development. Early on, upon retreat, the longwall headgate T-junction experienced very high face convergence, accompanied by roof sag, floor heave, and rib loading. The floor heave was exacerbated by the presence of floor coal, which buckled upward near the center of the beltline. Roof movement was occasionally aggravated by laminated roof conditions, but was kept in check by aggressive supplemental support including cable bolts, yielding props, cribbing, and roof mesh. Rib bolting and rib mesh helped to further maintain the entryway stability by keeping sloughage at a minimum. Nevertheless, despite the heavy support, headgate convergence was so large that, in a few places, it threatened longwall retreat and ultimately required the bottom to be re-graded."

Jan 1, 2018

By Jeffrey Johnson

To measure the loading behavior of friction bolts, researchers at the Spokane Research Laboratory of the National Institute for Occupational Safety and Health (NIOSH) installed strain gauges on the interior of friction bolts and developed a battery-powered miniature data acquisition system (MIDAS) that fits inside the hollow portion of the friction bolt. The advantages of this system are that it is protected from face blasts and eliminates the need for external wiring. Laboratory pull-out tests showed that friction bolts installed in a concrete block with resin were loaded to about 1.8 tons/ft of bolt length; bolts installed without resin were loaded to 1.3 tons/ft. Three strain-gauged friction bolts were installed at Hecla Mining Company's Gold Hunter Mine, Mullan, ID, in the wall or rib of a mechanized cut-and-fill stope. The object of these tests was to establish an installation procedure for the bolt-and¬MIDAS combination and to evaluate performance. The stope was advanced in three 15-ft increments and strains induced in the rock bolts by stress changes in the rock were recorded every 30 minutes. The MIDAS proved to be rugged enough to withstand the shock and vibration from nearby (5 ft) face blasting. Data from MIDAS showed that the friction bolt installed with resin was the most sensitive to each face advance. It showed a steady increase in loading to a maximum value of 1000 microstrain. The friction bolt installed without resin showed stick-slip behavior with loads to a maximum of 400 microstrain before the stope was completed.

Jan 1, 2003

By Kevin Jinrong Ma

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

Jan 1, 2011

By Rafal Misa

In order to prevent the formation of mining damage or to reduce its reach, it is advisable to look for technical solutions which would enable effective protection of construction sites from the impacts of underground mining. So far, other authors have mainly focused on developing methods of securing construction sites in order to minimize the damage resulting from underground mining; the issue of geotechnical methods for protecting buildings against mining damage has not been thoroughly covered in those articles. It should be noted that there are relatively few publications directly related to this issue, nor are there practical examples of geotechnical protection. In this paper, a geotechnical solution available for securing building structures is discussed. Trenches presented in this research reduce the size and range of mining deformation. The results of numerical modelling for sample protective solutions, aimed at reducing the impact of mining activity on the terrain surface under various conditions, are presented. The calculations were performed with the aid of the Abaqus 6.10-2 software, based on the Finite Element Method.

Jan 1, 2014

By Dennis R. Dolinar

Roof falls, even of supported roof, still constitute a major hazard in underground mines. However, associated with any fall or instability is a pattern of roof movement. Therefore, the National Institute for Occupational Safety and Health, Pittsburgh Research Center, has been conducting research into the development of practical extensometer systems that can be used to measure this displacement and to determine if a roof fall may occur, thus increasing miner safety. For extensometers to be used in the detection of potential roof falls in coal mines, an understanding of the movement and displacement rates that will lead to roof failure is required. Because roof falls are unpredictable, it is often very difficult to obtain this information. Therefore, in a coal mine in the lower Kittanning seam in Pennsylvania, a room widening experiment was conducted to intentionally cause a roof fall while roof movement data was collected. In this experiment, two adjacent rooms 5.5 m (18 ft) wide supported with 1.5 m (5 ft) resin-grouted bolts and instrumented with extensometers were widened to 9 m (30 ft). In one of the rooms, 3.6 m (12 ft) cable bolts were installed as supplemental support. In the room with no additional support, a roof fall occurred 72 h after the room was widened. Much of the 3 cm (1.2 in) of movement in this room resulted from the instantaneous and transient response to mining with displacement rates up to 7.6 cm/d (3 in/d) being measured. This was followed by a steady state displacement phase with a rate of 0.43 cm/d (0.17 in/d). Just 12 h before the fall, movement deeper in the roof caused the rate to accelerate to 1.1 cm/d (0.44 in/d). The roof in the room with the supplemental cable support did not fall even though the total roof deformation was over 7.6 cm (3 in). Significant roof movements and strains in the cabled room were detected to a depth of 3 m (10 ft) while maximum cable loads were over 180 kN (40,000 lbf). Essentially, the roof was in postfailure with the cables maintaining the material, while roof stability was indicated by a steady state displacement rate of only 0.025 cm/d (0.01 in/d) at the end of the test.

Jan 1, 1997

By Clarence O. Babcock

Since Vicat in 1833, the width to height ratio of a mine pillar has been taken as the fundamental variable in pillar design. From work in the laboratory by many investigators testing rock, Coal and concrete samples, it has been concluded by some that pillars with a width to height ratio of 10 to 1 are indestructible. It has also been concluded that a constrained core exists within the pillar which increases the pillar strength. The role of the roof and floor, if mentioned at all, is often noted in an incidental way. In the present work on laboratory testing of model pillars on concrete, coal, and rock instrumented with special strain gages, it was concluded that the end constraint and not the width to height ratio is the significant variable in determining the pillar strength. In tests where the end constraint is steel, a large compressive stress is produced on the horizontal plane of the pillar, and this effect increases as the width to height ratio is increased, resulting in an increase in pillar strength. If the end constraint is not steel but something with a small Young's modulus, the state of stress in the horizontal plane may be tensile, increasing as the width to height ratio increases, resulting in a decrease in pillar strength. Related to mining, the results are that a large width to height ratio pillar is strong for strong roof and floor and weak for weak roof and floor. That is, the roof and floor and not the width to height ratio determine the pillar strength. The confined core condition should be verified by testing if large pillars are to be used. If there is no confined core or the pillar is in tension, smaller pillar sizes should be used. The final conclusion is that the geometry alone is meaningless in pillar design.

Jan 1, 1984

By C. P. Mangelsdorf

The success of the Birmingham roof truss (Figure 1) in supporting some difficult roof conditions, particularly in the Illinois coal basin, has given impetus to the development of a number of alternate schemes. These schemes attempt to alleviate two major problems with the Birming-ham truss, namely, the time required for installation and the cost of material. Because of the difficulty of drilling angled holes and the number of different pieces of hardware which must be handled manually, most operators have been unable to install Birmingham trusses in the cycle and have had to schedule truss crews during the maintenance shift or whenever they would not impede production. Due to this delay the truss has traditionally been viewed as a supplemental form of roof support. In fact, the truss might qualify as primary support and replace at least two bolts if it could he installed in the cycle (and MSHA would accept it as an alternate in the roof plan.) [ ] Figure 1. The Birmingham truss. Two attempts to deal with the first difficulty have sought to streamline the truss hardware while preserving the essential character of the Birmingham truss. One of these, by Peabody Coal, is reported elsewhere in these proceedings by C. Bollier. Another is a contract issued by the Bureau of Mines to Bendix Corporation for the development of a machine, particularly suited for use in low coal, which will automate many of the operations currently executed manually. This machine is expected to be ready for field testing by fall of 1982. Once in place, a truss installed by either of these schemes will interact with the roof in the same way as a conventional Birmingham truss. An alternate approach has been proposed by several different manufacturers independently. Although the hardware details differ. the essential idea consists of two angle bolts, the heads of which are connected by a horizontal chord tightened to some pre-determined tension. Figure 2 shows the generalized concept and does not depict any one of the four systems proposed. The advantage of this scheme is that the angle bolts can be installed and tightened in the cycle, and as bolts, immediately satisfy the roof control plan. The horizontal chord can be added later to achieve almost any initial desired roof loading, the limits of which are defined below this scheme is referred to here as the angle-bolt truss. Because of a fundamental difference in the way in which load is transferred between chords, this design does not interact with the roof under subsequent loading in the same manner as a Birmingham truss. The purpose of this paper is to examine this fundamental difference and its consequences. [ ]

Jan 1, 1982

By Xinzhi (Thomas) Du

"In order to monitor the overburden strata deformation over a longwall mining area in the Pittsburgh Coal Seam, three boreholes were drilled across the longwall panel. Each borehole employed multiple-point borehole extensometers (MBE) to monitor overburden strata movement during and after the longwall face passed under the study area. The longwall face in the study area was 1,428 ft wide and 4,000 ft long. The average mining height was about 7 ft. The overburden depth ranged from 550 ft to 650 ft with an average of 600 ft. The final extensometer readings indicated that overburden strata initially showed sign of movement when the longwall face was 700 ft inby the MBE. The rate of strata movement accelerated when the face was 300 ft inby MBE. When the face was directly under MBE, a sudden anchor displacement occurred, indicating rapid strata movement once under-mined gob. The readings showed insignificant strata movement after the longwall face was 400–700 ft outby MBE.INTRODUCTIONThe longwall mining method is an efficient mining method of underground mining. Surface and subsurface strata are inevitably disturbed above the longwall mined out gob. Figure 1 shows the four zones of disturbance in the overburden strata in response to longwall mining.A longwall shear cuts the coal seam and allows the immediate roof strata to cave into irregular blocky material to fill the void. This is called the caved zone. The caved zone extends from the mining horizon to six or eight times the mining height, depending on the bulk factor of each caving strata layer. The fractured zone is located above the caved zone, in which the strata breaks into regular blocks by vertical or sub-vertical fractures caused by bending forces due to the weight of the underlying strata. The surface soil and rock crack zone extends to approximately 50 ft below the surface. The surface cracking is caused by downward movement of the surface. Between the fractured zone and surface crack zone is the continuous deformation zone. The strata in this zone deforms gently without causing any major cracks that extend long enough to cut through the thickness of the strata, as in the fractured zone (Peng, 2006).The objective of this project is to provide fundamental rock mechanics research about overburden strata movement response to longwall-related activity. The findings aid in better understanding pillar stability, gas well subsidence management, hydraulic impacts, and numerical modeling."

Jan 1, 2018