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By Chris Mark

Fully-grouted roof bolts comprise more than 80% of the primary roof supports used in U.S. coal mines. However, nearly 1,500 MSHA reportable, non-injury roof falls occur each year, and most of these are attributable to failure of the roof bolt system. Anchorage failure is one failure mechanism for fully-grouted bolts. As roof deformation works its way upward, the bolts can become heavily loaded near their upper ends, If the applied load exceeds the anchorage, the bolts will simply pull out. Research dating back 30 yrs indicates that this type of anchorage failure is most likely when the roof rock is weak, just where roof support is most critical. In soft shale or coal roof, the small amount of data available in the literature indicates that 20-30 in of resin anchorage may be required to achieve the full capacity of the bolt. In other words, the "full resistance zone" of a 60 in bolt may actually be just 30-40 in. Despite its potential importance, there is no widely accepted anchorage test for fully grouted bolts. Standard pull tests have sometimes been employed, but they provide no information on the anchorage near the top of the bolt. An alternative, first described more than 25 yrs ago in the U.S., is the short-encapsulation pull test (SEPT). With this test, the bolt is installed with only a short (1 ft or less) tube of resin. In recent years variations of this test have become international standards. This paper describes recent studies using short encapsulation pull tests in the U.S. Tests were conducted in the National Institute for Occupational Safety and Health (NIOSH) Safety Research Coal Mine at Bruceton and at underground mines in Pennsylvania and West Virginia. The study found that the SEPT can be used to make a simple evaluation of resin bolt anchorage. Suggested procedures for conducting SEPT are included. The study also confirmed that poor anchorage can be an issue, particularly where the roof rock is very weak. Some simple techniques for improving anchorage, and thereby the effectiveness of fully grouted bolts, are discussed.

Jan 1, 2002

By Anthony T. Iannacchione

This study, funded by the Appalachian Research Initiative for Environmental Science (ARIES), examines how mine layout, mining method, and geology interacted to allow water from the Grove No.1 mine pool to discharge into local surface waters. Federal and State regulations prohibit the discharge of acid or iron-producing waters from underground coal mines. Mining companies design barriers around their operations to prevent mine waters from discharging to the surface. The Grove No.1 represents one mine where these designs proved inadequate. Unique geologic conditions at the Grove No.1 mine pool, combined with retreat room-and-pillar mining, contributed to a discharge that required passive water treatment.

Jan 1, 2013

By Yong-Ming Jiang

Longwall mining systems are extremely complex and expensive and require great caution for reliable and efficient operation. In many cases, ground control problems are the primary concern in longwall mining. For example, unexpected threats such as roof falls or shield damage can cause sudden interruption of coal production and safety problems, which can originate from many parameters such as shield capacity, setting load, shield type, conditions of immediate and main roofs, pillar stability, geological anomalies, etc. However, it is difficult to correlate ground-control problems at a longwall face with the related mining and geological factors. In this paper, a method of correlation analysis, called tree classification, which permits the use of a large and variable number of features, is introduced The method provides an effective approach to produce an accurate classifier or uncover a knowledge structure, which will be used to predict future consequences depending on past and current data sets. The application of the tree classification method in longwall stability analysis is discussed, and the major procedures of constructing a knowledge structure are included as an example utilizing monitored leg pressure data in a longwall face. Also, the importance and ranking of the variables related lo longwall shield stability is discussed This is a major step towards developing an artificial intelligence system for real-time interpretation of shield monitoring data and hazard prevention.

Jan 1, 1993

By Gheorghe Oncioiu

The phenomena developed on the influence of thick coal seam no.3 mining, by horizontal slices and rocks caving roof control, in the Jiu Valley coal basin, are very complex. By analyzing the influence volume of underground faces and defining the stress and strain zones, around the faces, it is possible to evaluate the suitable powered support performances. Also, Jiu Valley thick coal seams involve about 50% of reserve balance in the main safety pillars of social, industrial and natural objectives, situated on the mining surface perimeters. For a correct design and drawing of main safety pillars complex research was done with the aim of establishing the subsidence parameters. Upper face rocks medium could be assimilated as a granular medium, which is in a permanent movement. For describing the rocks movement, respectively the ground surface displacement, was taken into account the stochastic mechanics principles. The theoretical solutions are in accordance within situ measurements.

Jan 1, 1999

By Boqiang Wu, Ganggang Xu, Xiaodong Wang, Hai Wang, Shibin Zhu

"Using a paste-like slurry as a filling from the ground surface to the goaf, artificial pillars can be built in the cavity goaf to support the goaf roof to prevent roof collapse and ground subsidence. If the accumulation laws of paste-like slurry are known well, the filling technology can be optimized. The laws can establish the foundation for finding out the best methods for building artificial pillars and filling technology. In order to understand the accumulation laws of paste-like slurry in the cavity goaf, Aeolian sand with wide distribution in northwestern China was selected as aggregate, and Portland cement was used as cementitious material to prepare a paste-like slurry. A test model of the cavity goaf was built to simulate the filling process by pump from the ground surface to the goaf to study accumulation laws of the paste-like slurry. The results show that the whole process of expansion and accumulation of the paste-like slurry could be divided into stages: free expansion, fill tightening expansion, and stable expansion. The angle of repose of the accumulation body of the paste-like slurry is constant in the stage of stable expansion, and it is the final angle of repose of the accumulation body. The angle of repose of the accumulation body is highly relevant to the slump of the paste-like slurry, and it is unrelated to the distance between the floor and roof. Under certain conditions of the cavity goaf, reducing the slump of paste-like slurry is one of the effective ways to obtain cost-effective pillars.INTRODUCTIONA large number of cavity goafs of coal mines were left in northwestern China; they were formed in the mining process by room-pillar or strip mining. Safety coal pillars were used to support the roof to keep long-term stability in the goaf, so the ground subsidence was not obvious (Fu, Deng, and Zhang, 2011; Xie, 2014). With gradual weathering, destruction of coal pillars, or additional load from ground surface buildings and structures, there could be a risk of instantaneous collapse of the roof in the goaf. This would affect the safety of existing structures on the ground and adjacent mines (Qu et al., 2017; Li, Xiang, and Jia, 2011; Tan, Guo, and Zhao, 2016; Tan et al., 2017). For disaster prevention of the falling cavity goaf, the current common practice is full filling (Hu et al., 2008; Liu, Zhu, and Huang, 2015; Pan and Wu, 2010). However, for buildings and structures with lower deformation requirements, full filling is excessive. For example, it is acceptable for goaf treatment under general roads and industrial sites to control the maximum deformation and keep the deformation continuous (Liu, 2011; Liu, 2012)."

Jan 1, 2018

By Jun Lu

This paper presents the evaluation, enhancement and comparison of two hydraulic fracturing techniques which were employed to mitigate the impact of a massive sandstone channel on two successive longwall faces. The hydraulic fracturing technique and Longwall Visual Analysis (LVA) software were initially employed to mitigate the impact of thick, massive sandstoneroof during mining of longwall Panel 21. Improved longwall face conditions and productivity were observed as Panel 21 mined through the sandstone channel. However, moderate sandstone fractures and falls at the face were still present in areas where hydraulic fracture influence was not present. Such adverse conditions caused two to three weeks of production delays due to large face cavities, jammed face conveyor and subsequent face gluing. Based on the mining experience in Panel 21, and site-specific roof geology in Panel 22, a modified and more effective hydraulic fracturing program was employed. Fewer well-placed holes were drilled, and larger volume of water was pumped into each hole in order to propagate the individual horizontal pancake fractures over a larger area, thereby enhancing caving and relieving face pressure caused by overhang of the massive sandstone. Longwall Visual Analysis (LVA) software was also employed at Panel 22 to predict the cavity risk index and guide the longwall face crew in double-pulling shields behind the shearer drum. This helped to reduce the tip-to-face distance at the critical spots along the longwall face, which helped mitigate the formation of the roof cavities and failure of massive sandstone onto the face conveyor. Underground observations, face advance rate and face rock delay analysis confirmed that longwall face conditions and advance rate in Panel 22were significantly improved compared to those of Panel 21. Consequently, safety and productivity were improved.

Jan 1, 2014

By Behdeen Oraee

Constructing tunnels underground is generally described as a high-risk activity because conditions in the surrounding area tend to be unpredictable. This would and can create unique dangers in every situation. It is important to gain a deep understanding of the risks associated with tunneling in order to reduce the likelihood of their occurrence and severity of their consequences should they occur. For this to be achieved, different risks should first be evaluated and ranked according to their relative importance and criticality. The Amir Kabir tunnel is considered one of the most dangerous tunnels in Tehran, Iran, in terms of the risks involved in its excavation and their potential consequences. In particular, part T4 of the tunnel passes through a zone of different strata and also a compound of various soils. This paper studies part T4 of the Amir Kabir tunnel from a risk analysis and risk management point of view. In doing so, factors increasing total cost, causing delay in the project completion time, decreasing the operation rate and downgrading project quality were identified and ranked according to their importance using the aggregate primary index risk method. For this purpose, a comprehensive questionnaire was designed and completed by tunneling experts. By responding to the questionnaires, the experts determined the relative likelihood of occurrence of the studied factors and their potential consequences. As a result, the experts suggested several alternatives to reduce the risks involved in this particular study. Moreover, the experts completed new questionnaires whilst taking into account different alternatives. In the next step, the risks before and after applying the alternatives were compared. As a result of performing the said analyses, it was concluded first that the aggregate primary index risk is an effective tool in identifying the risks involved in such projects. Second, it was concluded that, by taking into account different alternatives when analyzing and ranking the risks, severity of the said risks can potentially be reduced. Also, by considering such matters when analyzing tunneling risks, the projects? efficiency can greatly increase as a result of reducing risks that threaten financial and human resources. The approach introduced in this paper, together with the methodology described, can be adopted by the mining design engineer in all similar situations. They will be of particular advantage in such methods as room and pillar mining, where the number of similar tunnels required is high.

Jan 1, 2014

By David W. Evans

"Induced hydraulic fracture of strata during roof bolt installation is a potentially prevalent, but masked phenomenon within the underground coal industry. Previously reported resin testing programs (McTyer et al., 2014) examined the relationship between resin mixing effectiveness and varying bore hole diameter. The methodology employed within this earlier test program facilitated a further critical area of research – the measurement of back pressures generated within the bore hole during standard rock bolt installation practices. Experimental data has indicated that fluid resin can be pressurised to levels where it exceeds the compressive strength of the strata, inducing hydraulic fracture within the immediate area of the bolting horizon. The routine cycle of roof bolting serves to propagate this effect, progressively fracturing and delaminating the roof during mine advancement. This masked phenomenon can lead to a perception of difficult ground conditions - mining efficiencies and costs are therefore affected, with increased needs for additional support subsequently required to restabilise the inadvertently damaged roof.Further analysis of the parameters associated with resin bolt installation has now been conducted, assisting in the development of an empirical relationship between bore hole pressure, bore hole diameter and bolt insertion times. This relationship has been analysed for 15:1 ratio resins and 2:1 ratioresins, within 28mm and 30mm boreholes. Further to this, load transfer performance has been comparatively assessed for both 28mm and 30mm boreholes, suggesting that for 2:1 resins, acceptable resin mixing and load transfer can be obtained within a 30mm bore hole. The combination of 2:1 resins, utilised within a 30mm bore hole, may well provide the optimal solution to reduce the risk of hydraulic fracture in weaker strata during resin bolt installation.INTRODUCTIONA growing number of industry papers have previously investigated areas of concern associated with the performance of cartridge style resins in roof bolting, predominantly focussing on the effects of plastic film gloving, inadequate resin mixing and the pressurisation of resin within the bore hole. These three effects have a critical influence on the load transfer of the steel bolt element, through the cured resin and into the surrounding strata.Gloving occurs due to the plastic film of the resin cartridge partially encasing or wrapping around the steel roof bolt element – a known phenomenon over many years (Pettibone, 1987). The plastic film creates regions of discontinuity between the bolt, cured resin and borehole, reducing effective load transfer into the strata. Experimentation conducted into the effects of bolt ends fully encased by plastic film (Pastars and MacGregor, 2005) revealed that load transfer can be reduced by 85 to 90% in worstcase occurrences. It is also known that the aggregate filler size used within the resin can assist in shredding and breaking up the plastic film – this effect was observed in a previous study on resin mixing effectiveness (McTyer et al., 2014)."

Jan 1, 2015

By Stephen C. Tadolini, Anand Bhagwat

"Intrinsic supports installed in mines with limited seam or mining heights have always proven to be difficult when the required support length is longer. To date, three solutions have been used to ensure that the correct length bolt is installed to achieve the required anchorage or to maintain immediate roof material: combination bolts, milled-notch bolts, and hot-notch bolts. The combination bolts use threaded sections and couples, which can be screwed together to form the required length. The primary bolt of choice, to minimize costs and achieve the desired support, has been milled- or hot-notch rebar bolts. These notches can be placed at strategic (required) locations along the bolt axis, so the bolt can be successfully bent and straightened during the installation process. However, these notches can reduce the strength of the bolting system; special reductions are detailed in ASTM F432-13 (2013). Technical advances in steel rolling and forged heading processes have led to the development of a deformed rebar bolt termed, the BORE bolt, which stands for Bendable Oval REbar. The BORE bolt can be bent at any location and straightened during the bolt installation cycle. This support exceeds the minimum designed capacities described in ASTM F432-13 and meets the yield and tensile strength requirements of a full diameter no. 6 rebar bolt. The profile and shape also reduces the resin annulus during the installation process to enhance mixing and shreds the cartridge film to help eliminate glove fingering.INTRODUCTIONThe height or depth of competent strata material, in the immediate mine roof, can be greater than the developed mine openings heights. This is especially true in coal mine openings where 50% of the minable seams in the US are less than 50 inches (132 cm). When this scenario occurs, it can be difficult to insert steel bolts in a solid single length. Figure 1 illustrates what difficulties can be experienced when a roof bolter operator is working and supporting a thin seam coal mine. There are three viable options currently available. Lengths of the bolt can be threaded and a steel coupler can be used between the sections to reach the required depth of anchorage. These systems, called “coupled bolts,” usually require a larger diameter hole to facilitate the larger coupler diameter and can be anchored with an expansion shell, resin, or a combination of both. To ensure that the coupler will develop the tensile capacity of the bolting system, required by ASTM 432-13 Standard Specifications for Roof and Rock Bolts and Accessories (2013), the threaded bars should be completely rotated and engaged into the coupler. Premature failures can occur when the threads “strip out” if not inserted completely."

Jan 1, 2018

By William D. Gallant

This paper presents the results of a joint cooperative research project between the CANMET Cape Breton Coal Research Laboratory and the Cape Breton Development Corporation to observe the subsidence profile over the stopline of 2 West panel, Phalen Mine. A novel technique was designed and Implemented to determine vertical displacements of the overlying strata from within a near horizontal borehole drilled over the 2 West panel. Restrictions on data collection techniques due to the submarine nature of the coalfield are discussed. Development and implementation of the borehole technique using a piezometer and inclinometer Is described. Performance of the instruments and profiles obtained by them are presented and compared. Recommendations for design improvements are made to Improve the efficiency of the technique.

Jan 1, 1990

By Anthony (Sam) J. S. Spearing

Backfilling using coal waste is not a novel means for waste disposal throughout the world; yet, backfilling using a high density (paste) backfill is relatively new, except in Germany and Poland. One of the main benefits of a high density backfill, as opposed to a low density (slurry) backfill, is that, if placed at a low permeability and low moisture content, the risk of water table contamination can be relatively low. In addition, if paste backfill can be placed cost effectively, environmental benefits are huge because the surface impoundment footprint can be greatly reduced or even eliminated. This is especially true in room-and-pillar applications. In its current state, backfilling is considerably more costly than surface disposal. This is mainly because paste, as opposed to slurry, significantly increases backfill placement costs; however, this could be offset by increasing the achievable extraction rate. Overall, such a system could potentially improve the public?s image of coal mining in the US for a myriad of reasons. This paper discusses a proposed backfill system for a typical room-and-pillar coal mine in the Illinois Basin of the US and clearly shows the technical advantages, although costs are still higher with fill. The cost differential will likely reduce over time, in favor of the backfill option, as costs of traditional surface disposal increase, due mainly to ever-restrictive disposal regulations.

Jan 1, 2013

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 Soheil Bahrampour

Void detection and identification of bed separation while drilling holes in rocks for roof bolting can be a great source of information for the evaluation of ground conditions and adjustment of ground control or support measures. This can be accomplished through the monitoring and analysis of drilling parameters by the roof bolter. This paper reviews work currently under way in a joint study between Penn State University and J. H. Fletcher & Co. The detection of voids, along with identification of rock types, can lead to development of 3D roof maps, which can subsequently be used to evaluate ground support and its suitability for given ground conditions in tunneling and mining operations. Existing algorithms for void detection utilize the sensory information of the drill unit, such as feed and rotary pressure, revolutions per minute (RPM), and penetration rate, to detect voids or joints on the roof/ ribs. While the void detection algorithms are being revised to allow analysis of available data from the existing sensory systems of the roofbolters with higher accuracy and lower number of false detections, additional sensors have been added to the drill to improve its ability for accurate detection of the voids and joints. This includes installation of vibration and acoustic sensors on a drilling unit at J. H. Fletcher & Co., and pertinent data analysis which are discussed in this paper. Initial results indicate that the data obtained by vibration and acoustic sensors at high sampling rates can be used to improve the performance of the void detection algorithms independently, and when combined with the modified algorithms using existing sensors, it can substantially improve the capabilities of current void detection system.

Jan 1, 2013

By Elia Elias, Alan Crosky, Paul Hagan, Damon Vandermaat, Serkan Saydam, Peter H. Craig

"The University of New South Wales (UNSW Australia) had been involved in the study of premature failure of rock bolts in Australian coal mines from the initial identification of the problem in 1999 (Crosky et al 2002). Rock bolt steel changes over the last decade appear to have not reduced the incidence of failures. A broadened UNSW research project funded by the Australian Research Council (ARC) and Industry has targeted finding the environmental causes through extensive field and laboratory experiments. This paper describes the field studies conducted in underground coal mines, in particular attempts to measure the contribution to corrosion from groundwater, mineralogy and microbial activity. Various underground survey techniques were used to determine the extent of broken bolts, with the presence of both stress corrosion cracking (SCC) and localized deep pitting making no single technique suitable on their own. Groundwater found dripping from bolts across various coalfields in Australia were found to be not aggressive and known groundwater corrosivity classification systems did not correlate to where broken bolts were found. In-hole coupon bolts placed in roof strata containing claystone bands confirmed the clay as being a major contributor to corrosion. Microbes capable of contributing to steel corrosion were found to be present in groundwater, and culturing of the microbes taken from in-situ coupon bolts proved that the bacteria was present on the bolt surface. An ‘in-hole bolt corrosion coupon’ development by the project may have multiple benefits of 1) helping quantify newly developed corrosivity classification systems, 2) providing an in-situ ground support corrosion monitoring tool, and 3) for testing possible corrosion protection solutions.introductionThe problem of premature rock bolt failure in Australian coal mines was published in 2002 and 2004 by UNSW Australia from Australian Coal Association Research Program (ACARP) funded projects. The majority of the broken bolts examined had steel Charpy impact toughness values of 4 – 7 Joules, and the failure mechanism was most often stress corrosion cracking (SCC). Steel fracture mechanics predicts that an increase in impact toughness will increase the length of the crack before sudden brittle failure. In the final report of 2004, anecdotal evidence from one coal mine indicated that the problem may be eliminated in some environments by a change to steel grades with higher Charpy impact values of ~16. (Crosky et al, 2002 and 2004)"

Jan 1, 2015

By Jamal Rostami, Asok Ray, Samer S. Saab Jr., Wenpeng Liu

Ground instability, such as roof or rib failures, is one of the most serious and frequent safety hazards that occur in underground mining, tunneling, and underground construction. The ground-related parameters and features of interest, in assessing roof or rib failures, are location, frequency, and orientation of the joints, voids, as well as rock-strength. These parameters are the main components of rock mass characterization that offers effective strategies for ground support, which, in turn, allow one to mitigate the risks related to ground instability. The main objective of this study is to determine the location of joints and classify rock-strengths by analyzing the drilling data recorded from an instrumented roof bolter while drilling for rock bolt installation during the operational cycle. For this purpose, computer programs based on updated pattern recognition algorithms were developed for joint-detection and classification of rock types to offer an estimated strength. This paper briefly introduces ongoing research on joint detection, especially for joints with an aperture less than 0.125in (3.175mm), by employing an instrumented roof bolter in controlled environments. To improve the capability and precision of joint detection programs in detecting the joints from the drilling data and reducing the number of false alarms, many laboratory tests with various simulated joints and rock-strengths were carried out at a J.H. Fletcher & Co. facility. This paper reviews testing procedures, data analysis, updated algorithms used for joint detection, and discusses the latest round of testing in samples with simulated joints at various angles along the borehole.

Jan 1, 2017

By Ashok Kumar, Rajendra Singh, Arun Kr. Singh

"A simple analysis of site conditions and capability of the mechanised depillaring (MD) operation found that this approach is suitable to address the strata control issues in coal pillar extraction under the weak and laminated roof strata of Pinoura mine. A continuous miner (CM) with a ram car and quad-bolter are used for single pass pillar extraction (first time in the country) to achieve a relatively faster rate of depillaring. A total of eleven panels are depillared under weak and laminated roof strata with a fair-rated rock mass rating (CMRI-RMR). It is observed that there is less influence of roof caving over the surrounding mining structures, especially those located out-bye the side of the extraction line. An attempt is made to quantify the goaf influence by field studies, which show that the maximum value of mining-induced vertical stress is 2.83 MPa, and the range of influence remains limited to around 15 m from the line of extraction. However, the problem of roof instability under the influenced zone was experienced due to the inherent nature of the weak roof strata. The chances of trapping of CM under weak and laminated roof strata of the panels is reduced by deriving a threshold value of the roof dilation, on the basis of a simple statistical analysis of the obtained field results. However, few incidences of roof collapses are observed over the CM due to the unpredictable behaviour of the geological structures below the freshly exposed strata, which also caused an obstruction in the speed of the extraction. This paper presents results of the field and laboratory studies, conducted for a successful application of the single-pass pillar extraction method, along with details of some CM trapping cases below the weak and laminated roof strata.INTRODUCTIONCoal is the major natural resource to meet the growing energy demand of India and this trend is likely to continue for the next 20 years in the country. A major part of the Indian coal belongs to Gondwana period, and only a small part belongs to Tertiary period. The estimated coal reserve in India per the exploration work done up to a maximum depth of 1200 m is 315.149 billion tonnes, including the proven coal reserve of 143.058 billion tonnes (CMPDIL, 2017). Even in presence of these huge resources of coal in India, there is a considerable gap between production and the users’ demand of coal. Most of the mines are operated by conventional semi-mechanised Bord and Pillar (B&P) methods; thus, the mining industry has gained a considerable amount of operational experience (Singh, Singh, and Dhar, 1996) and knowledge about strata behaviour (Singh and Singh, 1999; Singh et al., 2008) using this method. However, recently, the Indian coal mining industry has introduced some mechanised depillaring (MD) operations, where the level of production is relatively high. This approach of depillaring improves safety along with the production and productivity of a depillaring operation (Ram et al., 2017). The nature of the overlying strata of these depillaring operations vary greatly: from highly laminated, weak, and easily caveable to massive, strong, and difficult to cave."

Jan 1, 2018

By Chase K. Barnard, Rahul Thareja, Sean Warren, Rajagopala Kallu

"Inflatable rock bolts are commonly utilized for ground support in Nevada underground mines, however, there is limited information regarding what factors influence the bond strength of these bolts. Bond strength is an important parameter in friction bolt support design, and this information is usually not available from manufacturers as a standard. Previous research has investigated the bond strength of friction bolts; however, these studies focused primarily on split set bolts and ground conditions more favorable compared to those commonly encountered in Nevada underground mines. Without site-specific bolt pull-out tests, ground control engineers are required to make assumptions regarding the in-situ bond strength of rock bolts. This is especially a concern for projects in the feasibility and initial development stages before site-specific experience is acquired.The purpose of this paper is to establish confidence in anticipated minimum bond strength for inflatable rock bolts by comparing the bond strength to variable geotechnical conditions using the Rock Mass Rating (RMR) system. To investigate a correlation between these parameters, the minimum bond strength of pull-out tested inflatable rock bolts was compared to the RMR of the rock in which these bolts were placed. Bond strength vs. RMR plots indicate that expected minimum bond strength is positively correlated with RMR, however the correlation is not strong. Cumulative percent graphs indicate that 97% of pull-out tests result in a minimum bond strength of 1 ton/ft and ½ ton/ft in RMR =45 and <45, respectively.Although lower bond strengths are more commonly encountered in low RMR ground, high bond strengths are possible as well, yielding higher variability in bond strengths in low RMR ground. Bond strength of friction bolts relies on contact between the rock bolt and drill hole. Experience in Nevada indicates that RMR is known to affect both the quality and consistency of drill holes which likely affects bond strength. Drilling and bolting in low RMR ground is more sensitive to drilling and bolting practices, and strategies for maximizing bond strength in these conditions are discussed."

Jan 1, 2015

By David F. Gearhart, Mark Van Dyke, Heather Dougherty, Ted Klemetti, Gabriel S. Esterhuizen

"Coal mine longwall gateroads are subject to changing loading conditions induced by the advancing longwall face. The ground response and support requirements are closely related to the magnitude and orientation of the stress changes, as well as the local geology. This paper presents the monitoring results of gateroad response and support performance at two longwall mines at a 600-ft and 2,000-ft depth of cover. At the first mine, a three-entry gateroad layout was used. The second mine used a four-entry, yield-abutment-yield gateroad pillar system. Local ground deformation and support response were monitored at both sites. The monitoring period started during the development stage and continued during first panel retreat and up to second panel retreat. The two data sets were used to compare the response of the entries in two very different geotechnical settings and different gateroad layouts. The monitoring results were used to validate numerical models that simulate the loading conditions and entry response for these widely differing conditions. The validated models were used to compare the load path and ground response at the two mines. This paper demonstrates the potential for numerical models to assist mine engineers in optimizing longwall layouts and gateroad support systems.INTRODUCTIONThe need for roof stability in coal mine longwall gateroads places significant demands on the gateroad ground support system. The variable loading conditions associated with the approaching longwall face can cause excessive ground deformations that need to be accommodated by the support system. In coal mines in the United States several different types of supports are used to maintain the stability of gateroads. The supports may include grouted roof bolts, cable bolts, and roof trusses as intrinsic support. Standing supports include timber cribs, engineered timber props, and various types of cement-based cribs. The intensity of support is governed by the expected geologic conditions and the expected stress changes that will occur during longwall mining. At present, the design of ground support for gateroads is largely based on local experience. In-mine trials are usually conducted when new support technologies become available or ground conditions dictate a change in the support system. The research reported in this paper has the objective to assist mining engineers in evaluating alternative gateroad supports and to design suitable support systems for the local ground conditions."

Jan 1, 2018

By William J. Schloemer

INTRODUCTION The history of coal mining the Ostrava Region of the Czech Republic goes back nearly 250 years. Both coking coal for steelmaking and thermal coal for the production of energy are mined here; however, it was the coking capabilities that subsequently turned the region into a major steel-producing area in the early 1800?s, something that continues to this day. Ostrava lies in the northeast corner of the Czech Republic, near the southern border of Poland, and falls within the Upper Silesian Coal Basin. The region contains over 400 seams of coal of which about 150 have been mined over its history and continue to be mined. Eighty percent of the Upper Silesian coalfieldlies in Poland and is presently mined by about 35 coal mines; the remaining 20% falls in the Czech Republic, which presently has four mines operated by OKD, a.s., a subsidiary of Amsterdam-based New World Resources. Traditionally, all the mines in this region have employed the longwall method of mining, of which subsidence is an unavoidable consequence. In fact, due to the long history of longwall mining here, there are areas where the overall subsidence has exceeded 37 meters (121 feet). The region is fairly densely populated with residences and industry, and a considerable amount of coal lies under some of these areas where subsidence is not allowed or has to be minimized. It is this reason that has driven OKD to consider alternate methods of mining that will not create subsidence. The decision was made to consider the room and pillar method of mining in order to leave stable pillars of coal to support the surface. After considering several locations, OKD chose the shaft safety pillar at the CSM mine as a test mining location. To say the least, this is a very challenging location in which to conduct a test. As with the entire region, there are many faults from tectonic activity, the seams are pitching at 12 to 20 degrees, the area is affected by high horizontal stress, and the depth of 750 to 900 meters is perhaps the deepest room and pillar coal mining ever attempted.

Jan 1, 2014

By M. Ashraf Mahtab

A feature of interest in stability of excavations in domal salt is the phenomenon of gas outbursts which has occurred in five of the six mines in Louisiana salt domes. Gas outbursts are sudden erruptions of salt and gas from the face or roof of a mine opening. They are associated with anomalous zones in salt, depth of mining, and the method of excavation. Potential risks from gas outbursts include an unplanned connection to a reservoir of water or hydrocarbons and, possibly, a loss of the mine. It is proposed that a gas outburst is initiated by disking in biaxial compression and propagates by spalling of the cavity walls and further disking. The outburst is terminated by a combination of dilatancy hardening, increase in stress along the cavity axis, and enlargement of the cavity to the boundaries of the pressure pocket. Identification of burst- prone areas will require research into in situ measurement of material characteristics and further under- standing of the mechanism of gas outbursts. Shock blasting in advance of mining appears to be a promising technique for controlling gas outbursts.

Jan 1, 1981