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By Fangting Dong

The broken rock zone (BRZ) and the tunnel support theory are investigated. The relationship between the BRZ and the related factors, the interaction between the BRZ and supports, and the function of supports are analyzed in detail based on the authors' underground observations and model taste. A summary of the application of the BRZ theory in the past is also presented.

Jan 1, 1988

By Ross Seedsman

Discussion on the design of roof support in tailgates has often been conducted without a clear statement of the stress and failure conditions acting. There is general agreement that in the tailgate the vertical stresses above the chain pillar increase. Within the stress 'bulb' associated with this vertical stress increase there will be an increase in horizontal stress. This does not mean that the horizontal stresses acting across the roof line increase. In retreat longwalling, the proximity of the tailgate to the adjacent goof results in a substantial reduction in the horizontal stress field acting in the immediate roof. Furthermore, at the face/tailgate corner the onset of significant compression of the chain pillar. if it is designed to yield, will result in a further reduction in the horizontal stresses in the immediate roof. It is suggested that this latter mechanism can cause the horizontal stresses in the immediate roof to go to zero. Whatever, the horizontal stress acting across the roof in the face/tailgate corner will be significantly lower than in the face/maingate corner. The implication of the model is that support design in tailgates should be based on the assumption of zero horizontal (in fact tensile) stresses. How the roof behaves in a tensile stress environment is a function of joint spacing and orientation compared to excavation width and direction. It is speculated that the empirical relationship between support density and roof rating results from the relationship between joint spacing and bedding spacing. The onset of stress reductions has major implications to the design of cable and bolt anchorages in the tailgate environment.

Jan 1, 2001

By Anthony Iannacchionc, Jessica Benner, Harold Miller, Megan Witkowski

"Longwall bituminous coal mining has been practiced in Pennsylvania since the 1970""s. Hundreds of longwall panels have been mined during the last 40 years but only 24 have actually undermined Pennsylvania's Interstate Highway System. The impact to interstate highways are summarized in this paper and the factors that influence the magnitude of these impacts are examined.INTRODUCTIONOver the last 30 years, interstate highways in Pennsylvania have been undermined by portions of 21 bituminous coal mine longwall panels. From 2003 to 2008, three of the panel were extracted by the Emerald Mine and six by the Cumberland Mine both in Greene County and both owned by Alpha Resources. Previous assessments of longwall mining under Pennsylvania’s interstates have identified impacts to the highways but no danger to public safety. In general the magnitude and extent of impacts during the recent undermining of Interstate 79 or 179 appears to be similar as those identified in previous studies. The nine longwall panels are among some of the widest in the U.S. They are supercritical in character with overburdens ranging from greater than 152 m (500 ft) to less than 305 m (1,000 ft).The Pennsylvania Department of Environmental Protection (PA DEP) requested that the University of Pittsburgh summarize observations collected by their staff during the recent undermining of 179. Observations and photographs of impacts to the highway were made by PA DEP staff and entered into reports prepared by California District Mining Office (COMO) staff and statements written into the Bituminous Underground Mining Information System (BUMIS) database. In addition to the observations by the PA DEP, the Pennsylvania Department of Transportation (PennDOT) conducted a series of detailed land surveys along portions of the highway undermined from 2003 to 2008. To characterize the subsidence basin fanned by longwall mining. PennDOT information has been analyzed by another research group within the university (Guticrre7, et al., 20 I 0)."

Jan 1, 2010

By Jessica M. Wempen, Michael K. McCarter

"Differential Interferometric Synthetic Aperture Radar (DlnSAR), a satellite-based remote sensing technique, has potential application for measuring mine subsidence on a regional scale with high spatial and temporal resolutions. However, the characteristics of Synthetic Aperture Radar (SAR) data and the effectiveness of DlnSAR for subsidence monitoring depend on the radar band (wavelength). This study evaluates the effectiveness of DinSAR for monitoring subsidence due to longwall mining in central Utah using L-band (24 cm wavelength) SAR data from the Advanced Land Observing Satellite (ALOS) and X-band (3 cm wavelength) SAR data from the TerraSAR-X mission.In the Wasatch Plateau region of central Utah, which is characterized by steep terrain and variable ground cover conditions, areas affected by longwall mine subsidence are identifiable using both L-band and X-band DlnSAR. Generally, using L-band data, subsidence magnitudes are measurable. Compared to X-band, L-band data are less affected by signal saturation due to large deformation gradients and by temporal decorrelation due to changes in the surface conditions over time. The L-band data tend to be stable over relatively long periods (months). Short wavelength X-band data are strongly affected by signal saturation and temporal decorrelation, but regions of subsidence are typically identifiable over short periods (days). Additionally, though subsidence magnitudes are difficult to precisely measure in the central Utah region using X-band data, they can often be reasonably estimated.INTRODUCTIONDifferential Interferometric Synthetic Aperture Radar (DlnSAR) is a satellite-based remote technique that can be used to measure surface displacement over large regions with high spatial resolution. Under good conditions, displacements can be measured with centimeter to subcentimeter accuracy (Buckley, 2000; Massonnet and Feig!, 1998). DlnSAR also has high temporal resolution, and imaging periods typically range from 10 to 50 days (Rosenqvist, Shimada, and Watanabe, 2004; Roth, 2004). In the last two decades, the application of DlnSAR for mine subsidence monitoring has been demonstrated in coal basins in Europe, Australia, China, and the United States (Camec and Delacourt, 2000; Ge, Chang, and Rizos, 2007; Ge et al., 2008; Ismaya and Donovan, 2012; Ng et al., 2010; Perski, 2000; Perski and Jura, 2003; Popiolek and Krawczyk, 2006; Wegmuller et al., 2005; Wright and Stow, 1999; Zhao et al., 2014; Xinglin et al., 2014). Overall, these studies have demonstrated good data resolution, strong relationships between mine development and subsidence, and reasonable agreement between displacements measured by DinSAR and displacements measured by conventional surveys."

Jan 1, 2016

By Muhammad Zaka Emad, Rangzeb Goraya, Muhammad Usman Khan

"Numerical modeling has been used extensively to assess the stability of surface and underground mines. Tabular, steeply dipping ore bodies are generally mined out by open stoping methods with delayed backfill. Cemented rockfill (CRF) is employed as a backfill material for narrow vein deposits with low production rates. Sublevel stoping methods with delayed backfill are employed to mine out steeply dipping ore bodies. Backfill is exposed to blast-induced vibrations and dynamic loading due to loss of confinement, which could lead to backfill failures. Backfill failures could influence the operating cost of a mining stope and the productivity of the entire mining project. This work shows a detailed overview of the factors contributing to backfill failure. A detailed literature review is presented along with case studies of longhole stoping operations from the Canadian Shield region for this work. The backfilled stopes are analyzed with the finite difference code FLAC3D. The literature review shows that the factors primarily responsible for fill failure are the strength of backfill, stope geometry, mining sequence, and blast-induced vibrations. INTRODUCTION Sublevel stoping methods with delayed backfill are popular in Canadian hardrock and metal mines. Backfill facilitates ground control and greater recovery of ore with pillar mining. It reduces wall slough, thus decreasing ore dilution from adjacent walls. The production of cemented rockfill (CRF) is simple; rock aggregates are mixed together with cement slurry and then placed in a mined-out stope. CRF has a low water content, thus it requires no additional drainage unlike some other backfill. Low water-to-cement ratio enables faster curing rate, hence a faster mining cycle can be achieved (Annor, 1999; Hassani and Archibald, 1998; McKay and Duke, 1989). CRF is generally placed in stopes from the upper sill drift using a truck or a load haul dump machine. Some of the problems associated with CRF are the moderate to high operational cost, the difficulty to get a tight filling, and segregation. Backfill failure is a product of many factors, such as cementation, particle size distribution, placement method employed, water quality, segregation, and blast-induced vibrations. The problem of segregation in aggregates has been observed when aggregate is dumped from a backfill raise (Meech, 2012; Indiana DOT, 2013). Segregation becomes unavoidable when the aggregates are not properly sized (Yu and Counter, 1983; Stone, 2007; Yu, 1989). The exposed backfill zones will contribute to ore dilution due to backfill when subjected to dynamic loading due to fill failures. This paper presents a detailed literature review and numerical modeling results of a mine backfill, along with case studies from longhole stoping operations. The backfilled stopes are analyzed with the finite difference code FLAC3D (Itasca, 2017). The literature review shows that the primary factors responsible for fill failure are the strength of backfill, geometry of stope, sequence of mining, and the vibrations induced by blasting operations."

Jan 1, 2017

By M Nombe

Based on the analysis of a case study at POWELTON No.2 MINE, it was determined that mine stability problems had occurred due to floor softening. A fireclay (underclay) layer in the immediate floor was softening after coal extraction and failing under pillars load. The thickness of soft floor was considered as the governing factor in the determination of the magnitude of floor displacement. The effect of floor thickness on floor displacement was evaluated by creating seven models that were input into the LAMODEL program. The properties of the soft floor were determined from a bearing capacity test performed in the Mains close to the area where mine stability problems had occurred. Moisture increase was considered in the reduction of floor bearing strength. Various strain-softening properties were allocated to the materials representing the soft floor. The LOW GAP POWELTON No.2 MINE was modeled and LAMODEL program was used to determine stress and convergence on both the coal seam and the fireclay seam. The results of critical stress and convergence obtained from modeling were compared with the actual floor failure area in the mine.

Jan 1, 2002

By P. Oliver

Rockburst conditions were first encountered at Inco's Sudbury area operations in the mid 1930's when mining advanced to below the 610 m (2000 ft.) horizon. Now, mining is being carried out at depths down to 2130 m (7000 ft.) at Inco'a Creighton Mine. While rockburst potentials have increased with depth. improvements in mine planning, methods, support systems and developments in destressing have alleviated the condition to where it is less of a problem now than it was in the early years. Hang of the improvements were developed, of necessity. through observation and reasoning because these were the only tools available at the time. Now, we are making use of stress - measurements, numerical modelling and microseismic source location technology to - help further reduce rockburat hazards.

Jan 1, 1987

By Thierry Lavoie, Mark K. Larson

"Engineers at the National Institute for Occupational Safety and Health (NIOSH) and Itasca Consulting Group, Inc., conducted a study to determine the procedure for calibrating a caving model implemented in FLAC3D. The model is calibrated to measured surface subsidence. This sedimentary caving model was developed by Itasca for NIOSH for the purpose of more closely simulating mechanical processes during caving. When a model is considered calibrated to surface subsidence, the load between the extracted panel's roof and floor is a reasonable estimate of gob loading. Such information can also provide an estimate of the amount of overpanel weight that is transferred to the abutment. This load distribution is a parameter needed, for example, to estimate loading of pillars in longwall gate roads. For the current case, a calibrated model resulted in a calculated abutment angle of 30.6°considerably higher than the assumed value used in empirical methods, but reasonable considering the presence of massive stratigraphic members in the overburden. Because of additional information available after this study that the extracted seam height was actually greater in the first panel, a corrected range of abutment angle was estimated to be from 27.7° to 30.6°, still significantly higher than the empirical assumption. The expectation is that better estimates of abutment loading profiles will result in improved evaluations of mine layout design and, therefore, reduce the number of fatalities and injuries resulting from falls of ground.INTRODUCTIONThe safety of miners in underground coal mines depends on careful consideration of the mechanisms involved in failure and transfer of load. The successful design of pillars and panels in longwall and room-and-pillar coal mining requires thorough attention to any information and measurements available that help understand these mechanisms. Although empirical methods, such as Analysis of Pillar Stability (ALPS) Mark (1987); (Mark, 1992) or Analysis of Retreat Mining Pillar Stability (ARMPS) (Mark and Chase, 1997), can provide a sense of success or failure, complex geology and failure in the roof or floor material or complex material behavior of the coal can make evaluation by an appropriate numerical modeling tool necessary. The numerical tool must simulate important mechanisms and appropriate detail that underlie the mechanism. For example, Starfield and Cundall (1988) advised using ""the simplest model that will allow the important mechanisms to occur, and could serve as a laboratory for the experiments you have in mind."" To that end, data gathering about the geology, geometry, and material behavior is very valuable. Such data facilitate selection of an appropriate model and model inputs. Moreover, this data facilitate improved assessment of risk to miners and how to minimize that risk."

Jan 1, 2016

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 David Newman

The latest version of the Surface Deformation Prediction Software (SDPS) package is presented through its application to a number of case studies. The package now exploits all the benefits of the Microsoft Windows platform and has a direct interface to CAD software, thus facilitating input and output. Also, the package features advanced calibration routines to handle predictions over specific regions. The case studies presented highlight the features and capabilities of the new package.

Jan 1, 2001

By Xingkai Wang, Wenbing Xie

"To control the large deformation of soft rock around the roadway subjected to high stress, it is important to improve the stability of support structure, in addition to increasing support strength. In this study, the causes for the destabilization of U-steel frame and its loading characteristics were analyzed by theoretical analysis, laboratory experiments and field tests. The technology of constant support resistance, the principle of support structure compensation and the compensation technique are proposed. The results indicated that the main reasons for frame structure destabilization are caused by low strength of connector, irrational structure of connector, poor interaction between the surrounding rocks and support frame and structure, and unstable support structure. Thus, the corresponding technological measures, such as connector with high strength and rational structure are adopted. The bearing capacity of frame is unstable after slipping. In order to maintain the high support resistance when the frame is performing the support function, the frame connector should be retightened after slipping. Though the working resistance may be improved greatly by using the backfilling or back grouting of U-steel, the frame remains unstable because of the unstable frame structure. Therefore, according to the deformation characteristics of rocks surrounding the roadway and the structure instability of U- steel frame, the stable structure device is used to carry out the structural compensation of U-steel frame, and the whole bearing capacity of the frame and its structural stability are improved greatly.INTRODUCTIONThroughout the development of the roadway supports for coal mines in China, it can be seen that the supporting strength and supporting resistance increase continuously with increase in mining depth and supporting difficulties [1]. With respect to the ground control mechanism and supporting technology of rocks surrounding the soft rock roadways subjected to high stress, considerable research has been done both at home and abroad. The backfilling technology behind the high-resistance and yieldable U-steel wall, the new type of grout/bolt support technology and the bolting with high-strength mesh support technology were developed, and have been used widely in coal mines with notable economic benefits [2-7]. However, a large number of engineering practices show that in many cases the existing rock control technology of roadway may not control the large deformation of the soft rock roadways subjected to high stress. The reason that the large deformation of soft rock roadway occurs is mainly due to poor structure stability of existing frame bearing structure, not insufficient support strength."

Jan 1, 2015

By Tristan H. Jones

Rib-related accidents in underground coal mines continue to cause injuries and take lives at a rate of 1.3 fatalities per year for the last 17 years. Prevention of these accidents cannot simply be a matter of installing more rib support, though the eventual final solution will undoubtedly include that component. Ultimately, rib safety will be brought about through an understanding of the causes and mechanisms that cause rib fatalities, and by fine-tuning mining methods and procedures that are currently in place. This paper investigates the factors likely to be most influential in the occurrence of rib fatalities. A careful analysis has been conducted of the rib fatality reports documented by the Mine Safety and Health Administration between 1996 and 2013. The primary outcome is the determination of the relative impact of each type of rib fatality factor and a better understanding of the impact each factor has. Among the rib fatality factors considered, the influence of rib profileis considered for the first time. Mining height, rib stratigraphy, mining cycle, and the standing position of a miner next to a pillar are also considered. Questions are also raised about the current interpretation of rib fatalities due to mining depth, with the indication being that mines operating at the greatest depths may actually have the most respectable rib fatality record among all mining depths. Additionally, a new method of considering rib exposure is developed. Based on the results, recommendations are made for future work. By gaining a better understanding of how particular factors contribute to rib fatalities as opposed to rib failure or stability, it is hoped that future efforts can be more precisely directed towards the desired outcome of fatality reduction.

Jan 1, 2014

By S. K. Sharma

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

Jan 1, 1992

By 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 Chaojun Fan, Du Wenzhang, Sheng Li, Hong-Hong Shi

"Coal and gas outburst is an extremely complex dynamic disaster in coal mine production process which will damage casualties, equipment facilities and disorder the ventilation system by suddenly ejecting a great amount of coal and gas into roadway or working face. This paper analyzed the interaction among the three essential elements of coal and gas outburst dynamic system. A stress-seepage-damage coupling model was established which can be used to simulate the evolution of the dynamical system, and then the size scale of coal and gas outburst dynamical system was investigated. Results show that the dynamical system is consisted of three essential elements, coal-gas medium (material basis), geology dynamic environment (internal motivation) and mining disturbance (external motivation). On the case of C13 coal seam in Panyi mine, the dynamical system exists in the range of 8-12 m in front of advancing face. The size scale will be larger where there are large geologic structures. This research plays an important guiding role for developing measures of coal and gas outburst prediction and prevention. INTRODUCTIONCoal and gas outbursts are a complicated mine gas dynamic phenomenon that pose a serious threat to coal mine safety production. Since the first recorded coal and gas outburst occurring in Isaac coal mine of the Lule coal field in 1843, more than 40,000 outbursts have occurred in the world. China ranks first in outburst frequency and intensity. In recent 20 years, China on average increases 37 pairs of coal mines and more than 280 times outburst accidents every year (Yan, 2013). The coal and gas outburst occurred on Oct 20, 2004 in Daping Coal Mine, Zhengzhou Coal Group, caused serious gas explosions and ejected 1,894 tons of coal into the face, causing 148 fatalities. In recent years, despite implementation of numerous control measures, coal and gas outburst still occur frequently ash mining depth and mining intensity increase. The outburst mechanisms can be summarized in the following four aspects: the gas leading role hypothesis, the geostress leading role hypothesis, the chemical effect hypothesis, and combination hypothesis. The last hypothesis has won the acceptance of most scholars who consider that an outburst is the result of combined effect of stress, gas and coal (Yu, 1992).Numerous experimental measurements and numerical simulations were conducted to investigate the permeability and damage development process in coal seam. Sobczyk, et al (2011) carried out a laboratory research of the influence of sorption processes on gas stresses leading to coal and gas outbursts. Wang, et al (2013, 2015) conducted laboratory experiments to investigate the rapid decompression and desorption induced energetic failure in coal using a shock tube apparatus. Xu, et al (2006) developed a coupled gas flow and solid deformation numerical model and applied to simulate the coal and gas outbursts in underground coal mines using RFPA2D-Gasflow. An, et al (2013) and Xue (2014) analyzed major parameters and the effect of gas desorption on outburst initiation and established a model of gas migration and mechanical processes during excavation."

Jan 1, 2016

By C. Thomas Blevins

This paper is intended to be a hands on experience account about ground control in a Southern Illinois coal mine. Its aim is to show how a combination of real world mining practices and constraints along with practical engineering theory mesh together to formulate a ground control program in adverse conditions. These conditions - your typical localized slips, splits, rolls, clay intrusions, water, changing types and thicknesses of materials, etc. - being amplified by a very strong horizontal stress field. A review will be made about the formation of the overall ground control plan and to what degree it has been successful. A presentation of the various roof control systems and their component parts, how these systems were developed, and the theory behind them will be made. The pointing out of some of the flaws that we found in the components, installations, and/or theory will be part of this presentation. A brief look at some of the efforts in the following items will be made: quality control of the roof control products; using the computer to help keep track of roof control information and soliciting help from various universities, governmental agencies and consultants to give different perspectives to the problem.

Jan 1, 1984

By Enguang Zhu, Kangning Zhang, Yang Li

"Due to the use of outdated mining technology or room and pillar mining process in small coal mines, the coal recovery ratio is only 10% - 25%. In many regions of China: the damage area caused by the small coal mines amounted to nearly one hundred square kilometers. Therefore, special mining techniques must be taken to reclaim the wasted resource in disturbed coal areas. This paper focuses on the different mining methods by analyzing the longwall panel layout and abandoned gateroad (AG) distribution in in the abandoned area of Cui Jiazhai Coal Mine in northwestern China. On the basis of three-dimensional geological model, FLAC- 3D numerical simulation was employed. The abutment pressure distribution was simulated when the panel face passed through the disturbed areas. The proper angle of the inclined face was analyzed when the panel face passed through the abandoned gateroads. The results show that the head end of the face should be 13-20m ahead of the tail end. The pillars on both sides of abandoned gateroads had not been damaged at the same time, and no large-area stress concentration occurred above the main roof. Therefore, the coal reserves of disturbed areas can be successfully recovered by using underground longwall mining. INTRODUCTIONDuring the 70-90's, there were a lot of small coal mines in China. Due to their disordered mining, most of the small coal mines left a great many abandoned gateroads (defined as AG in this paper) and a large area of damage coal seam. In these type of coalfields, there exist many mining risk issues, including nonuniform stress in the overburden, flooding and hazardous gases accumulated in the gob. So how to lay out a longwall panel and successfully mining under this type of abandoned small mines is a big issue nowadays in China.As mentioned above, longwall panel is being performed in the damaged coal seams in order to increase recovery of coal reserves (Wang, 2009). However, the abandoned geteroads intersect to each other all around the area. So a series of production issues arose while the panel face is passing through these AG. Gang and Gong (2015) analyzed the roof stability and abutment pressure distribution during the face was passing those AG that were parallel and inclined to the faceline. Xie, et al. (2015) proposed the ground control mechanism by passing the AG in top-coal caving longwall mining. He recommended to stop the face advance until the roof is over, using high strength bolts and cable bolts, and grouting consolidation. This paper studies the different mining methods by analyzing the longwall panel layout and AG distribution in damage coal seams. Theoretical analysis and field observation were conducted to determine the overburden movement when the face was advancing under the AG. The pillar between face and parallel AG changes from elastic to plastic and eventually to the residual supporting status. All these three status was combined with the roof structure to establish the theoretical model to calculate the shield support capacity. Numerical modeling was employed to simulate the pillar stability and determine the abutment pressure distribution as the face inclined to the AG during longwall mining."

Jan 1, 2016

By Lou Gaozhong, Guo Wenbing, Zhao Gaobo, Wang Shuren

"The height of a fractured zone due to high-intensity mining (HIM) plays a vital role in the safety analysis of coal mining under bodies of water. This paper described the characteristics of HIM. The processes of overburden failure transfer (OFT) are analyzed. The state of destruction transfers from a certain stratum to the upper stratum as face advance distance increases, and the processes of OFT are divided into two stages: transmission development stage and transmission termination stage. Through theoretical analysis, the limited suspension-distance and limited overhanging distance of each stratum are proposed to judge the damage of each layer. Based on this, the OFT criteria of fractured zone development are obtained. The mechanical model of strata suspended integrity and overhanging stability are established. Finally, a new method to predict the height of fractured zone due to HIM is put forward based on the processes of OFT, the limited suspension-distance, and the limited overhanging distance. Further, a HIM panel in the Datong coal mine is taken as an example. Theoretical methods are proposed, and engineering analogy is used to predict the height of the fractured zone, which is compared with in-situ measurements reported in the literature. The results show that the in-situ measurements for the height of the fractured zone are in close agreement with the theoretical calculation result. Therefore, the rationality and reliability of the theoretical method proposed to predict the height of fractured zone is verified.INTRODUCTIONThe height of the fractured zone (HFZ) due to the exploitation of coal is a key parameter for the safety analysis of coal mining under a body of water (Wenbing, Youfeng, and Hou, 2012). Accordingly, it is necessary to predict HFZ under the corresponding mining conditions before coal mining."

Jan 1, 2018

By Joe Wickline, Mark A. Van Dyke, Wen H. Su

"Geological factors and mine design contribute to mining-induced seismic activity, and this is especially true in longwall mining. Massive rock units are a key cause of seismic activity due to catastrophic failures that release large amounts of energy. Other factors, such as overburden depth, panel design, pillar design, rock strength, and proximity of the massive rock unit to the coal seam, all have a role in the potential and size of a seismic event. A recent seismic event was recorded by a deep longwall mine in Virginia at 3.7ML on the local magnitude scale and 3.4 MMS by the United States Geological Survey (USGS) in July 2016, which had no impact on the mining operations. Further investigations by NIOSH and Coronado Coal researchers have shown that this event was associated with geological features that have also been associated with other, similar seismic events in Virginia. Detailed mapping and geological exploration in the mining area has made it possible to forecast possible locations for future seismic activity. In order to use the geology as a forecaster of mining-induced seismic events and the energy potential, two primary components are needed. The first component is a long history of recorded seismic events with accurately plotted locations. The second component is a high density of geologic data within the mining area. In this case, 181 events of 1.0ML or greater were recorded by the mine’s seismic network between January, 2009, and October, 2016. Within the mining area, 897 geophysical logs were analyzed from gas wells, 224 core holes were drilled and logged, and 1,031 fiberscope holes were examined by mine geologists. From this information, it was found that overburden thickness, sandstone thickness, and sandstone quality contributed greatly to seismic locations. After analyzing the data, a pattern became apparent indicating that the majority of seismic events occurred under specific conditions. Three maps were created using MineScape geological mapping software. MineScape deploys an interpolator known as FEM (finite element method) and is based on a series of gridded triangles to forecast the probability and magnitude of an event if a particular panel were to be mined. The forecast maps have shown accuracy of within 74%–89% when compared to the recorded 181 events that were 1.0 ML or greater when considering three major geological criteria of overburden thickness of 1900 feet or greater, 20 to 40 feet of sandstone within 50 feet of the Pocahontas number 3 seam, and a longwall caving height of 15 feet or less."

Jan 1, 2017

By Anthony T. Iannacchione

The Solar No. 7 and Solar No. 10 mines were designed to prevent mine waters from directly discharging to the surface. In January of 2004, a discharge occurred as the Solar No. 7 mine was in its final stage of flooding through a barrier averaging 456-ft in width over a 1,600-ft length (No. 1). Another barrier (No. 2) was designed to protect four municipal water wells operating 1,800- ft away and producing from within or slightly below the Upper Kittanning coalbed from the adjacent Solar No. 7 and No. 10 mine pools. These two barriers were selected for this study because of varied mine layouts, mining methods, and geology factors affecting their performance. Analysis was aided by the collection and georeferencing of 6-month mining maps from the Pennsylvania Department of Environmental Protection (PA DEP). A structure contour map of the Upper Kittanning coalbed was constructed from mine map elevation points. These data and 2-ft surface contour maps were used to measure the precise location of the coal outcrop. The actual discharge point was as little as 425-ft away from a thinpillar section and 475-ft away from a pillar recovery section. The discharge point was found to be located slightly above the coal, indicating that the drainage path was at least partially through the strata directly above the Upper Kittanning coalbed. Hydraulic conductivities and potential discharge pathways are analyzed and discussed. The possible factors influencing the performance of the down-dip barriers at Solar No. 7 and Solar No. 10 are: a) the width of the barrier, b) the applied hydraulic gradient, c) the type of mining adjacent to the barrier, d) the overburden, e) the pattern and characteristics of fractures, and f) the quality of waters contained within the mine pools. This study, part of a larger effort funded by the Appalachian Research Initiative for Environmental Sciences (ARIES), is tasked to determine the key factors responsible for the successful design of down-dip coal barriers.

Jan 1, 2013