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By Hongwei Zhang, Sen Yang, Shimin Liu

"This laboratory study investigates the damage evolution of sandstone specimens under two types of cyclic loading conditions through monitoring and analyzing the variations of elastic moduli and ultrasonic velocities. We adopted the rigidity degradation method for evaluating the damage accumulation through the rock dynamic ultrasonic velocity measurements. Compared to the ultrasonic test, the method of rigidity degradation shows deficiencies in the detection of damage development. The cyclic loading within the elastic limit led to a stiffening trend for the loading process but a softening trend for unloading process. The method of rigidity degradation was unable to describe the damage accumulation in this situation, while P-wave velocity clearly reflected the damage development. Moreover, a simplified model is established for interpreting the behaviors of the modulus variations and ultrasonic velocity. The cyclic loading within elastic limit enhances the anisotropy of rock, which causes a denser packing pattern in the direction of axial stress. The persistent damage accumulation under the cycling with an increasing upper limit is the consequence of the tensile failure in the microstructure of sandstone. INTRODUCTION The damage accumulation due to dynamic (cyclic) loading is highly detrimental in many engineering structures, such as mine openings, petroleum and natural gas boreholes, tunnels, rock foundations, and underground chambers. The sources of cyclic or repetitive loadings can be roughly divided into two types: (1) periodic operation, including drilling, blasting, and mining; (2) random vibration, including earthquakes and traffic loads. The repetitive loading-unloading process opens and closes the micro-pores and micro-cracks within the rock, further inducing the growth of cracks. As expected, the accumulating damage becomes a potential factor to trigger the rapid and violent failures of large-scale engineering rock structure."

Jan 1, 2017

By Michael J. Castner

"INTRODUCTIONMultiple-seam mining is an ever growing concern for engineers and mine management in every portion of the mining industry and is especially true for coal mining operations. The goal of this report is to identify the location and severity of multiple-seam interactions, provide AMSS calculations performed on the worst of these locations, and follow up with a visual examination of the underground physical effects on the roof and ribs in the locations identified for analysis. Multiple-seam mining is increasingly being studied because:1. Multiple-seam mining can create a multitude of problems such as roof falls, pillar crushing, pillar failing, rib spalling, floor heave, and inundation of water, gas, or oxygen-deficient air that can disrupt mining operations and threaten the safety of the miners.2. Multiple-seam mining is becoming more prevalent due to the depletion of virgin coal reserves.Currently, coal mining in the United States has progressed to the point that a growing number of mining companies are operating in coal seams located above, below, and/or adjacent to previously mined out coal seams. As a result, the state and federal governments are enacting higher safety standards on mining operations. For example, in years past, pillars may have been properly designed to handle single seam interactions for a particular mine or mines; however, today, where mining is now being considered above or below a previously mined area, these loading conditions may not be adequate to handle the removal of additional seams above or below the previously mined reserve. Studies from 1981 estimated that two-thirds of the mineable coal reserves in the United States would be impacted by multiple-seam implications. Today, over three decades later, it is obvious that most of the remaining mineable coal reserve will be influenced by previously extracted seams. Thus, thefocus on multiple-seam mining interactions on current and future mining operations. In 2012, underground mining in the United States produced 342,287 thousand short tons of coal. Northern Appalachia accounted for over 30% (103,301 thousand short tons) of this total (U.S. Energy Information Administration, 2013). Rosebud Mining Company and Western Allegheny Energy (RMC and WAE) have a combined total of 23 actively producing underground mining operations and at least four new mines under or ready for development. All of these mines are located in coal seams (Upper and Lower Freeport; Upper, Middle, and Lower Kittanning) of northern Appalachia (Pennsylvania and Ohio). Currently, more than half of the RMC and WAE mines are dealing with multiple-seam interactions. Seven mines are undermining previously mined out coal seams, and five mines are overmining previously mined out coal seams. Also, between these companies, five mines have inner-seam sloping possibilities (2 completed, 3 planned), most of which will result in mining over or under previously mined areas."

Jan 1, 2015

By Sahendra Ram, Arnn Kr Singh, Dheeraj Kumar, Rajendra Singh

"Inherent existence of different rooms (openings) along the extraction line in a depillaring panel provides an easy path for the goaf to encroach the working area. Each of these openings in a mechanized depillaring (MD) operation along the extraction line is supported by the roof bolt-based breaker line support (RBBLS) to restrict the goaf encroachment. Different field studies by Council for Scientific and Industrial Research - Central Institute of Mining and Fuel Research (CSIR-CIMFR) found that the RBBLS works effectively during the depillaring under the shadow of stable surrounding ribs/fenders only. It is observed that, generally, these natural supports experience high value of induced stresses after a large amount of void creation inside the goaf. The induced stress creates spalling/loosening of the sides of the natural supports. Field studies also found that the positions of the RBBLS at the goaf edge need to be adjusted according to the depth of this spalling/ loosening. However, a systematic study in field for the correct positioning of the breaker line as per the spalling/loosening under different site conditions is difficult. Therefore, on the basis of the data and guidance provided by these field investigations, a detailed parametric investigation was conducted in the laboratory on simulated models for an effective design of RBBLS in a MD panel. FLAC3D is used for the simulation of the depillaring, and the available empirical formulations for Indian coalfields are used for subsequent calibration. An analysis of stress redistribution for different stages of the MD in simulated models showed the significance of the nature of overlying strata/induced stress for the RBBLS design. Discussing some results of field studies, this paper presents results of the laboratory investigations for an effective design of the RBBLS.INTRODUCTIONContinuous Miner (CM)-based mechanized depillaring (MD) is an important issue for Indian coalfields, where huge amounts of coal are locked up in standing pillars (Dixit and Mishra, 2010). Considerable variations in site conditions of these developed pillars bring a number of technical challenges into consideration during their final extraction. One of these challenges is goaf encroachment/overriding during caving of the competent roof strata. Goaf encroachment/overriding can effectively be controlled, mainly, by suitable design of natural supports (rib/snook/fender). However, there is a requirement of applied supports against the immediate roof as a considerable portion of this roof is exposed due to the presence of different openings along the line of extraction. These applied supports act as a fulcrum at the goaf edge between the working area and the goaf to facilitate the break out of the hanging roof inside the goaf. This is called breaker line support. An effective breaker line support is important for the efficiency of a depillaring operation. Generally, MD adopts roof bolt-based breaker line support (RBBLS) at the goaf edge (Singh, et al., 2014). For the first time in India, roof bolts were applied at the goaf edge as RBBLS in a CM-based MD (Leeming, 2003). In the beginning, three rows of roof bolts were installed as primary and secondary breaker lines support. On the basis of the observations of instrumented bolts, only two rows of roof bolts were used as secondary RBBLS (Ram, et al., 2014). Now, most of the MD operations in India are using only two rows of bolts as RBBLS (Table 1). However, the strata mechanics studies at different MD operations showed interesting characteristics resulting from this support (Ram, et al., 2015)."

Jan 1, 2016

By Essie Esterhuizen, John Ellenberger, Dennis Dolinar

"Underground stone mines in the United States use the room-and- pillar method of mining. The stability of the roof between pillars determines the mining dimensions and size of equipment that can be used. The findings of a survey of roof span stability issues and design practices are presented, and the observed factors contributing to roof instability are discussed. Identified stability issues are evaluated, such as maximum roof span, horizontal stress, and roof beam stability. The results of field observations, roof monitoring, and numerical analyses are used in the evaluation. The importance of the first roof beam thickness and its impact on roof stability and support requirements is presented. An evaluation of the current experience with stone mine roof spans relative to international experience is presented. A step-by-step design procedure is suggested which emphazises the need to collect adequate geotechnical data, understand the immediate roof stability, select a mining direction, and meet support requirements.INTRODUCTIONIn room-and-pillar mines, the roof between the pillars is required to remain stable during mining operations for haulage as well as access to the working areas. Roof and rib falls account for about 15% of all reportable injuries in underground stone mines (Mine Safety and Health Administration, 2009). Therefore, the stability of the roof is an important safety consideration. The size of the rooms in underground stone mines is largely dictated by the size of the mining equipment. Large mining equipment is used, requiring openings that are on average 13.5 m (44 ft) wide by approximately 7.5 m (25 ft) high to operate effectively (Esterhuizen et al., 2007). The dimensions of the roof spans are largely predetermined by the equipment requirements, and design is focused on optimizing stability under the prevailing rock conditions. If the rock mass conditions are such that the desired stable spans cannot be achieved cost effectively, it is unlikely that underground mining will proceed. In addition, large roof falls can extend across the full width of a room and may extend more than 5 m (16 ft) above the roof line. These large falls are a safety hazard and can have a significant impact on mine access and production. The National Institute for Occupational Safety and Health (NIOSH) research into stone mine roof stability has focused on identifying both the causes of roof instability and techniques to optimize stability through design."

Jan 1, 2010

By Uwe Wyink, Guenther Thiele, Isabel Gollnick, Klaus Thyrock

"The application of a thin spray on liner was implemented in longwall T-junction sections in 2009 for the first time. The development of the material and the application is based on a technical cooperation between RAG Deutsche Steinkohle AG and BASF Corporation. The report gives an overview of the first trials. The underground test occurred in a seam with a height of 1.5 m (5 ft) in a depth of 1,100 m (3,609 ft). The roadway is dual-use entry and the roof strata is composed of shale and sandy clay. In some sections geological faults impact the surrounding strata. The poor conditions induce separating and loosening of roof rock. Several remedial actions have to be implemented for decreasing deformations at the face entrance.The application of thin spray on liner (TSL) provides simplification and cost reduction of supporting work at longwall T-junction. The installation of TSL to avoid roof and coal fall operations for lagging are omitted.INTRODUCTIONFor safe operation of face retreat in dual-use roadways, the processes of securing and stabilizing the face end area are very important in order to minimize deformation and stress. Avoiding roof and coal fall are the main concerns for safe operation. For that reason different methods of reinforcement are tested in underground operation. In this paper the application of TSL is presented.The specifications of TSL are described, followed by the case study results. The underground study is located in one mine of RAG Deutsche Steinkohle. The extraction of the hard coal in this mine is characterized by great depth and high stress. There are poor geological and geotechnical conditions. The roadway used for the underground application was headed with the support system combination type A, which will be described later in the paper. After the first face retreat, deformation of steel arch support and bagging destroyed the roadway, so for the second face retreat remedial actions had to be undertaken. For this a new system of reinforcement, simplifying the supporting work at the longwall T-junction was tested. The reinforcement consisted of different parts that are described in the following report. Furthermore two different versions were applied. One method is the operation with TSL and the other one is the use of concrete for sealing. Both methods are compared in the text."

Jan 1, 2010

By Zbigniew Lubosik, Stanislaw Prusek

"The main elements of hard coal extraction systems in the Polish mining industry, including the characteristics of gateroads, are described, followed by several dozen underground measurements of gateroad roof sag performed in different geological and mining conditions in Upper Silesia Coal Basin.Based on the analysis of the roof sag measurements, an attempt to describe the gateroad roof sag process using a geometrical model was made. To build the model, it was assumed that the roof sag curve could be described by three straight lines. The first line describes the roof sag ahead of the longwall face; the second line, the sag at the longwall-gateroad junction zone; and the last line describes the zone behind the longwall face. Determining the intersection points and inclination angles is essential for the model. These model parameters depend on the geological and mining conditions in the gateroad.Using the statistical computations and analysis, the dependencies between the characteristic points of the gateroad roof sag model and the selected geological and mining parameters were defined. The STATISTICA computer program by StatSoft was used for computation.This paper presents a simple form of the geometrical description of roadway roof sag progression that is very useful during the gateroad support design process.INTRODUCTIONThe longwall system with caving is the dominant method in European mining in cases of hard coal seam extraction. For instance, in Poland, in 2009, 117 longwalls were driven, with total production of 77.5 million tonnes. In roughly 95% of the longwalls, mining operations were conducted using longwalls with roof caving; the remaining 5% used the hydraulic stowing system. The average mining depth was about 700 m (2,297 ft), and average daily output per face was approximately 2,600 tonnes per day. When using the longwall mining system, gateroads play an important role. These principal roadways transport materials, mine crews, and output. In addition, they impact ventilation of the extraction area, affecting, among other things, the climatic conditions or methane hazard. Unfortunately, difficulties often appear in maintaining an adequate size for the gateroads, due to the face line, which causes origination zones with stress concentration around the gateroad, resulting in rock mass movements in the form ofroofsag, floor heave, or horizontal convergence (Figure 1)."

Jan 1, 2010

By Robert B. Francini

Longwall mining and pipelines have coexisted since longwall mining was introduced into the United States. How the pipeline operator reacts to longwall mining under their line depends on many factors. This paper outlines the process the pipeline operator goes through to determine the appropriate response when their pipeline is being undermined. Typical mitigation procedures are reviewed. Finally, the paper looks at current design practices for new pipelines that are expected to be undermined during their lifetime.

Jan 1, 2011

By Andre Cezar Zingano

"The objective of this paper is to estimate the coal seam strength and the pillar safety factor for low coal seams using back analysis and rock mass characterization. The coal seam strength is very important to estimate both pillar capacity and mine section design. There are two approaches to estimating coal seam strength: (i) by geomechanics characterization of the boreholes cores; and (ii) by back analysis from pillars successes and failures, or some in-situ tests. For low coal seams (1.2 m, or 47 in.), the confinement of the center of the pillar is very important to defining global pillar strength because the low seam height increases the confinement of the pillar core, even in small pillars. We used in this case study one underground coal mine in the Parana State where the coal seam thickness is 0.8m the pillar height is 1.3m, the entry width is 5m and seam depth is 130m. A panel test was prepared to study the pillar behavior for use of the retreat mining method. The designed pillar size was 6x5m and 1.3m high with a predicted safety factor of 1.0 based on the Hoek and Brown failure criterion is 5.3 MPa. Once no pillar failure was observed, the pillar safety factor was assumed at 1.0 and the coal seam strength was backcalculated. The back analysis estimated the coal seam strength to be 6.2MPa. The confinement of the pillar center directly affects the coal pillar strength, and also the coal seam strength estimated by back calculation.IntroductionThe coal seam strength is very important to estimating pillar strength and mine section design. There are two approaches to estimating coal seam strength: (i) by geomechanics characterization of the coal seam using borehole cores; and (ii) by back analysis from pillar successes and failures, or some in-situ tests like panel test or ultimate pillar capacity (Salamon and Munro, 1967; Maleki, 1981; Peng and Dutta, 1992). For the first approach, lab tests (mainly triaxial tests) and borehole geomechanics descriptions (quality and quantity of the discontinuities) determine the coal seam strength by the Hoek and Brown failure criterion (Hoek and Brown, 1997). For the second approach, a panel test to study the pillar behavior can be built to define coal seam strength by back calculation. The safety factor for the pillars in the test is 1.0, based on the estimated coal strength from the geomechanics characterization. In this case, the probability of failure is 50%. If some pillar failure occurred, and the real pillar size was measured, the back analysis could be applied to define pillar strength and the coal seam strength."

Jan 1, 2015

By Phil R. Ames, Anil K. Ray, Murali M. Gadde, John A. Rusnak

"Many factors can be responsible for coal mine roof falls including, mining geometry, stress conditions and local geology. In the Illinois Basin coal mines, roof instability is commonly attributed to the low strength of the immediate roof, which is also moisture sensitive. Recently, in order to understand the mechanisms of roof instability, numerical modeling has often been the preferred ground control analysis tool. Since roof materials can exhibit significant spatial variability, such modeling results may not represent every section of the mine. At the same time, it is not feasible to carry out numerical modeling for the full range of geological and mining conditions possible at a mine. Operationally, it is more convenient and highly practical to have a simple map that depicts the expected roof conditions, determined through either complex numerical models or simple empirical analysis throughout the reserve area. In this paper, a case history is presented that shows how a combination of simple empirical analyses of geologic data and past ground control experiences at a mine can be synthesized into a roof stability map, which can then be used to forecast future ground conditions. The usefulness of such a roof stability map will be examined in terms of the predicted and actual ground conditions encountered over a period of more than a year after the map was created.INTRODUCTIONA roof fall, which can result in fatalities and lost work time, is considered to be one of the most severe ground control hazards in underground coal mines. In general, there are many factors responsible for roof falls, including mining geometry, stress conditions and local geology. In the Illinois Basin, roof falls are commonly attributed to the weak nature of immediate roof rocks, and some superficial falls may be initiated due to moisture sensitivity. Some unstable areas also occur when the immediate roof is highly laminated.Numerical modeling is a useful tool used to understand the mechanism of a roof fall by incorporating site-specific parameters. In the past, numerical modeling has been used to explain the mechanism of roof falls as well as simulate· roof falls and cutter behavior, but such work had been strictly site-specific (Gadde and Peng, 2005; Ray, 2009). Because of the restrictions imposed by the input parameters, numerical modeling results cannot represent the entire mine. In the real world, mining conditions vary from place to place, and even in a single panel, modeling results valid at one site may not apply at another nearby site. At the same time, due to the lack of input data pertaining to rock characteristics and in-situ stress variations, it is not practical to carry out numerical modeling for each local area of the entire mine site. In addition to these difficulties, mine operators prefer simple tools to understand the ground conditions at their mine sites. One simple and commonsense based tool is the roof stability map. A stability map not only synthesizes the geologic and stress analysis data empirically, but it can also be used to predict future ground control issues at that mine. Because of this practical appeal, stability maps are more likely to gain mine operators' acceptance and be used in their planning operations. In this paper, a case study is presented on the development of a 'Roof Stability Map' that was principally used for roof fall prediction for a mine in the Illinois Basin."

Jan 1, 2010

By Tex M. Kubacki

On August 6, 2007, the Crandall Canyon Mine experienced a collapse which trapped six mine-workers. An MW 4.1 seismic event was recorded in conjunction with this collapse. An additional 54 seismic events were detected and located in the vicinity via routine processing techniques during the month of the collapse. Waveformbased cross-correlation detection techniques have been used to uncover 1,328 small, previously undetected seismic events which occurred near the collapse area between July 26 and August 9, 2007. This work has resulted in lowering the magnitude threshold of completeness from ML 1.6 to ML -0.5, a 2.1 order-of-magnitude improvement over standard processing techniques. Event locations have been computed using a double-difference inversion. The spatial pattern of events ultimately helps to delineate the extent of the collapsed area. It is shown that many of the large catalogued events (>ML1.6) tend to be closely followed by numerous (often 20 or more) smaller aftershocks. Cluster analysis of the known catalog reinforces prior work indicating the existence of seismic clustering on the east and west sides of the collapse. Research now demonstrates that it is possible to differentiate events related to the main collapse from those related to the rescue attempt following the collapse. The cross-correlation techniques described herein greatly improve the ability to detect small mining-induced seismic events. These methods may lead to improvements in assessing seismic hazards for conditions similar to those observed at Crandall Canyon preceding the collapse.

Jan 1, 2013

By B. Macdonald

In 1991, roofbolts were introduced as a primary support for inseam retreat longwall entries at CBDC's Phalen Colliery. Increasing mine depth (+550 m) necessitated replacing conventional inseam steel sets in these entries because they were being severely damaged during wall retreat and subsequently causing gate end problems. Roofbolts had been used in the Sydney coalfield between 1949 and 1963 in room and pillar working, however, some concerns arose with using roofbolts as a primary support in long (2.2 - 3.4 km) single entry roadways. In order to alleviate these concerns, a phased approach was adopted whereby conventional steel densities remained along with roofbolts. The roofbolt design was such that the roofbolts would be the primary support and the steel sets would experience little or no loading. A detailed geotechnical monitoring program was developed by both CBDC and the Cape Breton Coal Research Laboratories technical staffs. The results of this monitoring were used to build-up confidence in roofbolts and subsequently reduce the steel set densities until eventually only roofbolts were used as a primary support. This paper will discuss briefly the phased introduction of rootbolts at Phalen Colliery and more specifically the development of geotechnical data collection and interpretation.

Jan 1, 1994

By Andrew Lewis, Amarit Nitaramorn, Alexander Garcia, Peter Altounyan

"Indonesia is a major coal producer, the vast majority being sourced from opencast operations. Underground production to date has been constrained by unfavorable ground conditions and limited local experience. However, an underground trial in East Kalimantan has now shown that it is possible to successfully extract coal and mine safely, despite the technical challenges.A trial drift mine has been worked to a current maximum depth of 160 m (525 ft) in a dipping coal seam. The 3 m (10 ft) thick coal seam is overlain by very weak mudstone and sandstones. The mudstone forming the immediate seam roof strata proved a particular challenge given its adverse reaction to water and overall friability.Rock bolts were used as roadway support, the system being carefully selected to optimize effectiveness. An innovative method of partial extraction was trialed, which demonstrated the potential for productive mining despite the ground conditions. Roof falls have occurred within the working area, and roadways demonstrate reasonable stability over time.Throughout the course of the trial, strata behavior monitoring has been an important aspect, and a comprehensive ground control monitoring system is in place. This has enhanced the ability to increase the level of support where required and to react to deteriorating conditions. In addition, monitoring information has allowed a quantitative assessment to be made of the technical feasibility of future full-scale mining.INTRODUCTIONWith current production at over 200 million tons, much of this destined for overseas markets, Indonesia is the second largest exporter of coal (Ritschel and Schiffer, 2007). The majority of the coal is sourced from the Kalimantan coalfields. Mining is predominately opencast, yet in recent years there have been attempts made to introduce underground operations. This move would permit continued access to good quality seams that are beyond the reach of viable opencast mining."

Jan 1, 2010

By Stephen C. Tadolini, Anthony T. Iannacchione

"Coal burst represented a major hazard for some U.S. mining operations. This paper provides an historical review of the coal burst hazards, identifies the fundamental geological factors associated with these events, and discusses mechanisms that can be used to avoid their occurrences. Coal burst are not common in most underground mines. Their occurrence almost always has such dramatic consequences to a mining operation that changes in practice are required. Fundamental factors influencing coal burst events include strong strata, abnormal strata caving, elevated stresses, critical size pillars and the lack of sufficiently sized barrier pillars during extraction. These factors interact to produce excessive stress, seismic shock and loss of confinement mechanisms. Over the 90 years of dealing with these hazards, many novel prevention controls have been developed including novel mine designs and extraction sequences, most of which are site specific in their application. Without an accurate assessment of the fundamental factors that influence coal burst and knowledge of their mechanisms of occurrence, control techniques may be misapplied and risk inadequately mitigated. INTRODUCTIONCoal burst1are violent failures of ribs, roof or floor in underground mines. They are known to occur in complex ways and often under unique sets of conditions. This has made them extremely difficult to anticipate and control. Assessing coal burst risk requires engineers, managers, and safety professionals to recognize the fundamental factors responsible for these events. This helps practitioners recognize the hazard potential and understand the context in which various controls and barriers should be used. Only then can coal burst events be anticipated andappropriate mitigation techniques deployed.Over the last 90 years, a few coal burst events have been anticipated but most have eluded prediction. The authors examined the historical record (Appendix A) and determined that an inadequate understanding of the initiating geologic factors and their associated mechanisms of occurrence have contributed to the poor predictive capability. The historic record also indicates that control techniques have typically been site specific in their application."

Jan 1, 2015

By Kazem Oraee, Mehran Gholincjad, Navid Hosseini

"Chain coal pillars are pans of the structure of longwall mining system that play a significant role in the stability of the entries. With mechanization and developments in the various aspects of the method, higher efficiency in optimization of the design of chain coal pillars seems appropriate. In this paper the three main methods of chain pillar design, namely the empirical analytical and numerical methods arc compared. Real data from the Tabaslran coal mines have been used in order to make the comparison process reliable. It is concluded that the most apparent advantage of the empirical method is the reliability of the results while the use of numerical methods enjoys the advantage of flexibility. On the other hand the analytical methods are complex unless simplifying assumptions arc made that can substantially decrease the accuracy of result which is thought to be the main advantage of the design method. A new method is therefore introduced here that combines all of the three presently used methods and by doing so the new method has the advantages of the three while minimizing complexities and inaccuracies associated with the use of the individual methods.INTRODUCTIONLongwall mining is a high production method that can use mechanization to its full extent. Its initial development can be traced to 17th century European collieries but it was not until the second half of the 20th century with the development of self advancing support systems that the method gained popularity in the USA (Oraec. 2001). During the last decade, the production of longwall faces has increased significantly (Peng. 2006). The ability for mechanization along with good safety condition is the main reasons for such development. In the longwall mining method a good design for panel and entries on both sides, are two essential parameters. In this method, the thickness of overburden (depth of mine) nom1ally ranges from 60 to 20 m (197 to 2.7 9 ft) (Hartman, 2000) and with increasing mine depth designing for stabilizing the entries on both sides of the panel become more difficult. Thus, for safety and stability the size of chain pillar must increase. This causes the overall coal recovery to decrease and increase mining costs per ton. As such wide pillars hinder the air flow and hence ventilation costs increases too. Therefore with accurate identification of loading conditions and stress analysis. pillar that are a minimum size and sufficient safety factor are carefully designed."

Jan 1, 2010

By Mark Van Dyke, Gamal Rashed, Khaled Mohamed, Timothy Barton, Morgan Sears

"The unconfined compressive strength (UCS) of coal is an important parameter that controls the stability of coal ribs in underground mines. Estimation of the UCS of coal units composing the rib is required for conducting the rib stability assessment technique. This technique requires conducting large numbers of conventional UCS tests in a laboratory, which are expensive and time consuming. The research work outlined in this paper examines the use of indirect strength estimation methods—the point load test (PLT) and the in-situ Schmidt hammer test (ISHT)—for evaluating the intact strength of coal. The conventional UCS test was conducted on 80 coal specimens from 10 different coal seams for eastern coals. For each coal seam, the PLT and the ISHT were compared with the UCS test results to establish a correlation between the strength indices from the PLT or ISHT and the UCS. Moreover, to simplify the estimation of the UCS, a correlation between the megascopic coal lithotype and UCS was established.INTRODUCTIONBased on Mine Safety and Health Administration (MSHA) reports, rib failures have resulted in 17 fatalities, representing 52% of the ground-fall fatalities in underground coal mines in the United States over the past decade (MSHA, 2018). In an attempt to control rib failures in underground coal mines, the National Institute for Occupational Safety and Health (NIOSH) is working on developing a coal rib stability rating technique to provide a quantitative means to characterize the stability of coal ribs and to provide guidelines for early selections of rib support. The unconfined compressive strength (UCS) of coal units composing the ribs is one of the key input parameters required to estimate coal rib stability. The UCS of coal is an important parameter that controls the resistance of coal ribs to fracturing and slabbing. Direct determination of the UCS is time-consuming and expensive. Moreover, the UCS may be limited to testing only the strong coal samples since the weak or highly cleated samples might not survive the transportation and the preparation process (cutting and grinding). Therefore, alternative methods have been tested to indirectly find the intact strength of coal, such as the point load test (PLT) and the in-situ Schmidt hammer test (ISHT). The acronym ISHT refers to Schmidt hammer tests that were conducted in-situ on coal pillar ribs, while the acronym SHT refers to tests that were conducted on small scale blocks of coal."

Jan 1, 2018

By Satyendra K. Singh

If the extraction is beneath massive roof strata, the roof tends to 'bridge' across large spans without failing, the caving behaviour is complex and irregular. Understandably, these give rise to many geotechnical problems with safety implications and involving sometime huge direct and indirect costs. Only two important problems, 'feather-edging' in pillar mining and `severe periodic weighting' in longwalling, are discussed in this paper in the background of geomechanics investigations in Australia at Blue Mountain Colliery and in South Bulga Colliery respectively. `Feather edging'- a major source of injury and fatal accidents in Australia and South Africa in such situations, refers to a phenomenon where the roof, at the goat edge, falls as a thin wafer of rock often suddenly, over-riding conventional timber breakerlines. The results of the first comprehensive, monitored trial in an Australian pillar extraction coal mine, suggests how the roof falls into feather-edging caving (generally) because of multiple shear-planes (interpreted) and how it may be alleviated. Because of complex fracturing behaviour of massive roof strata, periodic weighting at a longwall face is certain to manifest itself as 'yielding supports', heavy conditions', rapid convergence and in the worst case, as a cause of face stoppage. After briefly discussing the programme of instrumentation (in a chain-pillar and over the would-be goal) and results, this paper attempts to improve mechanistic understanding of the related geotechnical issues associated with the weighting behaviour and aims at establishing a 'threshold' value of weighting severity index (suggesting intensity of periodic weighting) with an objective to optimise operational controls before mining through potential (future) weighting areas.

Jan 1, 2000

By Yun Zhang, Haijun Jiang, Shenggen Cao, Chen Wang

The distribution of front abutment pressure is closely related to the force, deformation and energy distribution of hard roof before periodic weighting. So it is necessary to carry out research on the relationship between them. According to front abutment pressure distribution feature, using the location of peak front abutment pressure as the dividing point, coal seam is divided into two parts along the direction of mining: the yield zone that is the area between the faceline and the point where the maximum front abutment pressure occurs, and the elastic zone that is the part before the point of peak front abutment pressure. The proposed mechanical model of unit width hard roof at the panel center periodic weighting consists of five parts including the yield zone. All parameters of the deflection equations for each of the five parts that satisfy the continuity conditions and natural boundary conditions are obtained by using the Matlab software. The continuous curves of front abutment pressure, deflection, bending moment and bending strain energy density distribution of hard roof are obtained by iterative approximation method. The following conclusions can be drawn through analyzing those curves: (1) The position of peak front abutment pressure is located at the transition of coal seam foundation from yield zone to elastic zone (elastoyield critical point), the distance of it to the faceline is equal to the width of yield zone. The peak front abutment pressure increases with yield zone with. (2) The peak bending moment in front of the faceline, and the distance of it to the faceline increase with the yield zone width. The deflection of hard roof at the face and midspan also increase with yield zone width. (3) As the yield zone width increases, the accumulated elastic energy of hard roof in front of the faceline increases, and the core zone of energy moves forward away from the face. (4) When the yield zone width is zero, the whole coal foundation is regarded as elastic foundation, and the peak front abutment pressure occurs at the faceline.

Jan 1, 2015

By Murali Gadde

Among different types of ground control problems associated with underground coal mining, those related to in situ stresses are the most common ones affecting the safety and economy of a mining operation. As a result, this has been the focus of many ground control research works done in the last three decades. This paper summarizes a recent work done on in situ stress related ground control issues of underground coal mine development workings. The approach used in this study was three¬-dimensional finite element modeling by an implicit code, ABAQUS/Standard. In the models, roof stability evaluations were made using Hoek-Brown rock mass failure criterion. Influence of the in situ maximum horizontal stress angle on the stability of both entry and intersection was examined to help determine the best development layout orientation. Also, the effect of in situ stress ratios on the stability of development openings was studied. To make the work more general, both 'low' and 'high' in situ stress fields were considered In general, the modeling results indicated that entries oriented in the direction of maximum horizontal stress were in the best condition while those oriented at 90° had the least stability. For intersections, the maximum and the minimum stability were seen at 0°/90° and 45°, respectively. However, under some combinations of input conditions, the best or the worst conditions were noticed at same other orientations as well. Based on these findings, layout design charts were prepared for both entries and intersections. It was also found that the change in ground conditions with change in layout orientation was more significant for 'low' in situ stress fields than for the higher ones. Change in the average safety factor with change in the ratio of in situ maximum horizontal to vertical stress resembled a lognormal distribution curve. However, the effect of the ratio of in situ maximum to minimum horizontal stress lacked such a clear trend for different input combinations considered.

Jan 1, 2004

By Chris Kelly

The Monroe County Quarry is an open pit granite quarry located near the town of Bolingbroke. Monroe County. Georgia. The quarry has been in continuous operation for approximately 10 years. The quarry covers approximately 27 acres and has an annual production rate of 1.0 million tons. On the morning of April 26. 2000 a sudden rapid highwall failure occurred at the west end of the north wall of the quarry pit. The failure occurred approximately one hour after a three inch-wide crack was observed in an access road located at the crest of the highwall. No indications of failure had been observed prior to the development of the crack. The active mining area had been at the base of the failed area. The pit was successfully evacuated prior to the failure and there were no losses of either personnel or equipment. The amount of failed material was estimated to be in excess of two million tons. In order to ensure a safe working environment at the quarry a slope monitoring program was established in the area of the failure. The program consisted of using an EDM survey instrument to periodically measure the location of survey prisms installed at carious points on the north wall. The locations of the prisms were recorded over several months and changes in location were analyzed in order to identify and characterize the movement of the failed mass. The results of the analysis provided a basis for recommendations regarding access to various areas of the quarry. This paper discusses the factors contributing to the failure, the method of post-failure slope monitoring, anti the interpretation of the monitoring data.

Jan 1, 2002

By Monte R. Hieb

Study of moisture-induced swelling strain in Pennsylvanian Period rocks of the Appalachian Plateau in West Virginia shows a 2:1 biaxial horizontal anisotropy which trends NW-SE. Oriented in-situ slabs of coal mine roof rock were collected across West Virginia, dry-sawed to size, air-dried to stability, and then monitored for moisture changes and deformations in the horizontal plane through one wetting cycle and drying cycle. Biaxial swelling strains were exhibited in nearly all rocks tested, with maximum horizontal deformations occurring normally-incident to the strike of regional geologic fold structures. This relationship persists as fold strike varies across the study area. A connection with Alleghanian fold development and/or post-Alleghanian fold relaxation is suggested. Deformations were largest in shales, but occurred in sandstones and siltstones as well. During testing, moisture-induced fractures (MIF) developed parallel to the regional fold structure and ranged in size from hairline cracks to sample-destroying cracks that were orthogonal to the observed biaxial swelling strain direction. Preferential wetting and swelling of clay minerals along NE-trending fractures, related to a fold-parallel fracture fabric of regional extent across West Virginia, is the most likely reason for the observed systematic NW-oriented biaxial swelling strains. The testing methods used in this study are simple and inexpensive and produce robust results that offer insight into a regional, anisotropic mechanical weakness that imparts a dynamic, asymmetrical horizontal stress to moisture-sensitive coal mine roof rocks over time. Because a similar mechanical fabric has been reported in Devonian-age shales, these findings may have implications for shale gas developers in the Appalachian Basin challenged by rapid aperture closure rates along horizontallydrilled induced fracture networks in Devonian rocks like the Marcellus Shale. The paper concludes with an example of how the concept of a preferred swelling strain direction might be applied to improve the effectiveness of roof trusses.

Jan 1, 2013