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By Liane J. Terrill

Rockbolt failures at the Waste Isolation Pilot Plant have been recorded since 1990 and are categorized in terms of mode of failure. The failures are evaluated in terms of physical location of installation within the mine, local excavation geometry and stratigraphy, proximity to other excavations or shafts, and excavation age. The database of failures has revealed discrete areas of the mine containing relatively large numbers of failures. The results of metallurgical analyses and standard rockbolt load testing have generally been in agreement with the in situ evaluations.

Jan 1, 1995

By Andre Zingano

The coal seam strength and the failure criteria that represent the behavior of the coal seam are very important for mine section and pillar design. Additionally, mine section and pillar design take into account roof and floor layers, which form the roof-pillar-floor system. Therefore, underground ground control design begins, most importantly, with rock mass characterization. Another issue is the distribution or behavior of the coal seam strength for the whole mine area. Sometimes, the coal seam strength is assumed to be the same for an entire area, but this is not true because the coal seam is heterogenic and anisotropic, and also the coal strength is affected by faults and spacing between fractures. The objective of this paper is to demonstrate the use of the methodology for geomechanics characterization to determine the strength for the Bonito coal seam, and discuss the coal seam strength distribution around the entire mine area. The intact coal strength was determined by laboratory tests that were completed during the last 17 years, including triaxial and uniaxial strength tests. It was noted that there is a strong correlation between all laboratory strength tests with 90% confidence interval. The explanation for this coal seam strength variation is the presence of the discontinuities and their variation in quantity and quality at each location in the mine area, because the coal seam rock classification and its behavior will be affected by the quantity and quality of fractures.

Jan 1, 2014

By Srisharan Shreedharan, Gang Huang, Hongqing Song, Sijing Cai, Pinnaduwa H. S. W. Kulatilake

"Subsidence prediction due to ore extraction and backfilling during open stoping operations for an underground iron mine in China is reported in this paper. Geologic complexities including faults and interfaces between different lithologies, and the stoping and backfilling sequence adopted from the mine plans are incorporated in building a 3-D discontinuum numerical model. The stoping was simulated in two vertically stacked horizontal layers (Fig. 6), with a total of 16 stopes. Large displacements of up to 50 cm were predicted by the model along the roof of the stopes, and a maximum surface subsidence of 22.5 cm was predicted by the model . Backfilling was found to control subsequent deformation and subsidence. The extraction of the upper orebody led to relatively higher deformations in extracting the lower orebody. Finally, the predicted subsidences at all locations along a subsidence profile on the surface are shown. INTRODUCTIONThe Luohe iron (pyrite) mine, located in Hefei city, Anhui, is an underground mine in China, being mined through the open stoping and backfilling method. The orebody lies at a depth of about 400 m and has an average thickness of 250 m (Magang Steel Company, 2012a). Discovered in the 1960s, the location has a reserve estimate of 300 million tonnes, which have been mined since 2007. The orebody has been divided by the mining company into an east and a west section to facilitate planning for ore extraction.Figure 1 shows the surface above the underground mine, which is seen to contain many human settlements including villages, pastures and other infrastructure. These locations are sensitive in that they need to be protected from potential adversities due to the underground mining, such as surface subsidence. While it may be near impossible to completely eliminate the subsidence due to ore extraction, operations such as backfilling have been adopted by the mine in an effort to substantially reduce it. Hence, this paper focuses on describing the stability of stopes in the mine and estimates future subsidence in a part of the mining region through numerical modeling."

Jan 1, 2016

By David Bigby, Tim Ross, David Conover

"Extensive arrays of telltale roof monitoring instruments were recently installed in a Mexican copper mine and a large underground mined storage facility in the eastern U.S. The data were used to evaluate roof stability during development and retreat mining and after installation of supplemental supports. Both manual-reading and automated systems were employed and consisted of up to 64 instruments. This paper will discuss the practical issues involved with installation and monitoring, including reliability and maintenance. Examples will be given for telltale response in relation to known events affecting roof stability, including nearby pillar extraction roof caving, installation of supplemental cable bolts, and separation of the immediate roof layer. The strategy for processing the large quantity of data, presenting the data for review, monitoring the system remotely, and identifying and reporting critical events will be described.INTRODUCTIONTelltales have been used for many years with many varied applications. They are relatively inexpensive, easy to install and interpret, robust, and adaptable. The term ""telltale"" generally denotes a strata extensometer that incorporates a visual indication (and warning) of strata movement. The most common dual-height telltale (Figure 1) was developed by British Coal in the early 1990s as roof bolting was replacing steel arch support (Bigby, 2001). Dual-height telltales provide immediately visible measurements, distinguish between movements above and below the bolted height, and are an established means for providing preemptive warnings of roof falls.Many improvements and modifications on the basic design have been developed to adapt to different mining conditions. For example, triple-height telltales are commonly used where roof bolts of different lengths are installed. The standard manuallyread telltales are limited by the difficulty of obtaining frequent, consistent, and accurate readings, especially in high openings. The instruments may also be located in unstable or inaccessible areas. Consequently, an intrinsically safe electronic telltale system was developed that combines high accuracy and both local and remote reading options. The system supports a network of up to 100 dualheight telltales on each of four separate branches. The instruments are connected with a two-conductor cable and can be read both underground, using a portable readout, and on a surface computer (Figure 2). The automated system still provides the visual warning indication underground."

Jan 1, 2010

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 Christopher J. Brown

"Proper ground stability in gate roads has been a critical aspect of success in longwall mining. In particular, the tailgate has proven to be one of the more active mine areas that needs to be managed (Barczak, Dolinar, and Mucho, 1999). As longwall tailgate support technologies continue to be developed and improved, Strata Worldwide has worked to design and manufacture a revised Propsetter to provide additional support capacity and enable applications at greater mine heights. The support is targeted for tailgate applications; however, it can also be used for roof support in various types of mining environments. This paper describes the design changes for enhancing the support capacity of the Propsetter to more closely align its performance with other support technologies currently being employed in longwall tailgates. Balancing changes to each design aspect of the Propsetter was key to successfully improving the roof support load capacity while maintaining controlled deformation. Increasing the load capacity would allow a mine to improve support efficiency with a smaller, less intrusive support compared to support alternatives and to reduce the cost per foot of supported entry. The target support capacity was 80 to 100 tons at two inches of deformation to more closely match the performance of competing tailgate supports, such as the 24-inch-diameter CAN®. To achieve this goal, changes were made to all design aspects of the Propsetter: diameter, wedge cut depth and angle, pod size, and confinement rings. Test results from the NIOSH Mine Roof Simulator verified that these design changes were successful in increasing the capacity of the Propsetter to the 80- to 100-ton range at two inches of deformation. INTRODUCTION Roof support is one of the most fundamental elements of the underground mining process. As materials are removed from the earth, geological conditions will almost always require some form of roof support to keep the extraction area open and stable. Coal mining environments often experience difficult geological conditions that require both primary and secondary forms of roof support to keep the mining environment safe and economical. The secondary support system serves as additional support in the case the primary support system is inadequate. If designed properly, the secondary support system will also assist in maintaining integrity of the immediate roof beam (Barczak, 2000)."

Jan 1, 2017

By Kazem Orace, Navid Hosseini, Mehran Gholinejad

"Selection of the optimum support system for underground openings such as tunnels is a complex process. In this paper a new approach, based on a combination of the Analytical Hierarchy Process (AHP) the Technique for Order Preference by Similarity to Ideal Solution (TOPSIS) and the Preference Ranking Organization METHod for Enrichment Evaluations (PROMETHEE), is introduced. For this purpose, the selection process is assumed to be a multi criteria decision-making problem. First different support systems using FLAC3D numerical code, based on technical safety and stability parameters of the tunnel are specified. Then taking economic and performance parameters as the decision criteria by using the combination of Al-IP, TOPSIS, and PROMETHEE the optimum support system is selected. As a real mine case study this approach is used in the main access entry to the C, coal scam of Tabas collieries. Results clearly demonstrate that the proposed support system selection method is more advantageous than other alternatives.INTROD CTIONGenerally, the support system for a tunnel is selected primarily with the aid of design engineers experience. In other words personal judgment instead of intellectual and scientific criteria is mostly the main basis (Oracc. 2005). However since various support systems can be applied in any particular situation, accurate selection of the optimum support method depends on the integration of many technical, economical and performance parameters, and also on the analysis of the intensity of influence by each of the criteria.In this paper an applicable approach for se lection of tunnel support systems based on Multi Attribute Decision Making (MADM) techniques including: Al IP (Orace et al.. 2009a; Saaty, 1980), TOPSIS (Hwang and Yoon, 1981: Yoon and Hwang, 1995), and PROMETHEE (Brans and Vincke. 1985; Brans and Mareschal. 1992) is introduced. As a field study, this approach was applied to tunnel C1 as one of the main entries in the Tabas coal mine which is the major collieries in the central pan of Iran (Hosseini, 2008).Due to the coal seam conditions coal is mined by mechanized longwall mining (Hosseini 2008). This method requires the 58 excavation of several entries some of which will be used for many years and even for the entire life of the mine (Peng, 2006). Therefore the selection of the support system in these entries is·very important in mining design."

Jan 1, 2010

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 Rudi Podjadtke

In a short chronological review the development concepts and improvements in steel roadway support are shown. The successful standardization concept of RAG Deutsche Steinkohle and currently used support systems are important topics. The history of optimizing the material properties for the steel arch support with different alloys up to the tempering used today is discussed. The present-day application of the steel support is additionally presented. Today, it represents a system component of modern roadway support also for highly stressed roadways. In addition to the yielding roadway arch support, this overall system - the combination support - consists of the additional components of a rockbolting system and concrete backfilling. The comparison of the application areas for each individual partial system with the overall system illustrates its high performance capability. Actual practical experience with the state of the art typed steel arches are pointed out exemplarily. Understanding the complete system permits the selection of the economically optimal support for all conditions.

Jan 1, 2008

By T. P. Mucho

Strata behavior and support performance were evaluated in a longwall tailgate test area at the Eagle Nest Mine near Van, WV. The mine operates in the Eagle coalbed which ranges in height from 4 to 6 ft (1.2 to 1.8 m) on the mine property. Rugged topography results in rapid changes in overburden which ranges from 200 to 1200 ft (61 to 366 m). The immediate roof at this mine transitions from sandstone to shale. In some areas the sandstone appears to be massive while in many locations it is highly laminated, fossilized, and interspersed with coal streaks. Horizontal stress levels appear to be sufficient to create instabilities in the roof in some locations, especially where the roof is thinly laminated. Traditionally, hardwood cribs have been used at the mine to provide secondary support for longwall tailgates; Strata Products' Hercules cribs are installed on 8 ft (2.4 m) centers staggered left and right in the tailgate entry. In a 250 ft (75 m) tailgate test area, however, cable bolts were installed in lieu of cribs. The bolts were 12 ft (3.6 m) long tensionable cables anchored with 5 ft (1.5 m) of resin. Four cables were installed across the entry and rows were spaced on 6 ft (1.8 m) centers. Primary roof support was maintained in the vicinity of the test area using 42 in (1.1 m), grade 60, No. 6 (19 mm diameter) resin bolts installed on 4 ft (1.2 m) centers through T-2 channels. The cable test area was located under a stream valley in an area of relatively shallow overburden. Strata instabilities consistently had been associated with the stream valley in both the Eagle and the overlying Powellton coalbeds. In an effort to further expose the cables to the most difficult stress conditions available at the time of the test, the specific site chosen was directly beneath a barrier pillar remaining in the upper Powellton mine (approximately 150 ft (45 m) of interburden separates the two coalbeds).

Jan 1, 1996

By Mark K. Larson

Calibration of LaModel for deep coal mines begins with adjustment of overburden stiffness for load transfer distance. This calibration determines whether mining induced stresses are concentrated in the immediate abutment of the gob, or spread over a wider area. In the absence of site information, an empirically-based estimate is often used. Where data is available, a recommended calibration procedure can be followed, as is done in this case study at a deep western longwall mine, with minor modifications and extensions. In this case, data sources included borehole pressure cells, flatjacks installed in series with support Cans, and observed changes in ground conditions. Calibration was based on the first detection of mining-induced stress from the advancing longwall. Detection was found to depend on the type and sensitivity of measurement. Borehole pressure cell data with a 20-psi-increase detection threshold were chosen as the standard, in accord with data used in deriving empirical relationships. Application of this standard produced a load transfer distance estimate roughly four times the empirical relationship developed by Peng and Chiang (1984). This is a significant deviation with important implications for panel design at this site. Data from flatjacks, ground condition observations, and borehole pressure cells with higher threshold values, were tightly clustered and show load transfer distances from two to four times empirical guidance. Ratios between various measurements can be used to adjust measurements to the standard (although they may depend on mine geology). Once established, ratios can be used to adjust estimates, say from support flatjacks or ground observations, to the standard, where borehole pressure cell data is not available.`

Jan 1, 2012

By Khaled M. Mohamed, Michael M. Murphy, Ted M. Klemetti, Heather E. Lawson

"The design of rib support systems in U.S. coal mines is based primarily on local practices and experience. A better understanding of current rib support practices in U.S. coal mines is crucial for developing a sound engineering rib support design tool.The objective of this paper is to analyze the current practices of rib control in US coal mines. Twenty (20) underground coal mines were studied representing various coal basins, coal seams, geology, loading conditions, and rib control strategies.The key findings are: (1) any rib design guideline or tool should take into account external rib support as well as internal bolting; (2) rib bolts on their own cannot contain rib spall, especially in soft ribs subjected to significant load—external rib control devices such as mesh are required in such cases to contain rib sloughing; (3) the majority of the studied mines follow the overburden depth and entry height thresholds recommended by the Program Information Bulletin (PIB) 11-29; (4) potential rib instability occurred when certain geological features prevailed—these include draw slate and/or bone coal near the rib/roof line, claystone partings, and soft coal bench overlain by rock strata; (5) 47% of the studied rib spall was classified as blocky—this could indicate a high potential of rib hazards; and (6) rib injury rates of the studied mines for the last three years emphasize the need for more rib control management for mines operating at overburden depths between 500 ft and 1,000 ft.IntroductionEfforts to improve the stability of underground mine ribs have continued for decades. These efforts by mining professionals around the world persisted throughout the 1990s and 2000s as well, though much of this work focused on rib support methodology and selection of proper support for a specific location. Despite these efforts, it is unfortunate that there has been a continual occurrence of rib-related fatalities in underground coal mines. While overall improvements in rib safety have been achieved, the average fatality rate is still about 1.3 fatalities per year over the last 18 years (Jones, 2014).To manage and control the rib hazard, rib support is used in U.S. coal mines. The current rib support methods for U.S. mines fall into two main categories: rib control based on intrinsic support systems and those based on external support systems. Intrinsic or rib-bolt support requires a roof bolting machine that can install a bolt fixture into the rib. Furthermore, the appropriate level of support needed to control the ribs with an intrinsic-based support system as well as the design parameters for those support systems for various mine conditions have not been established in U.S. coal mines. Furthermore, surface control systems are often an integral part of these support systems and must be considered in the system design."

Jan 1, 2015

By M. K. Larson

The U.S. Bureau of Mines and Cyprus Shoshone Coal Corp. conducted a study of deformation mechanisms in strata around a longwall gate road system at two sites in an underground coal mine near Ham, WY. Of particular interest was timedependent (on the order of several months) roof deformation. Strata above and below the coal seam are very weak, carbonaceous mudstones that have cohesionless planes of weakness oriented along bedding, which dips approximately 8° to 16° at N 49° E. An extensive array of instrumented roof bolts, roof extensometers, and biaxial stressmeters were installed during development mining at these two sites under variable depth of cover [I83 to 335 m (600 to 1,100 ft)], seam dip (10° to 16°), and top coal thickness [estimated to be 0.30 to 0.61 m (1 to 2 ft) on the downdip side of the entry]. Bolt load and roof deformation histories during the entry development periods were compared. Bolt loads at site 2 were less than those at site 1, which is consistent with the amount of overburden. Deformation magnitudes at site 2 were similar to those at site 1, but deformation rates were approximately 44% greater and nearly constant over time, whereas the deformation rates at site 1 decreased exponentially over time. Also, most deformation occurred above the bolt horizon at site two, but extensometer measurements at site 1 showed that a significant amount of deformation occurred within the bolt horizon. The most significant factor that influenced bolt tension and timedependent strata deformation was strength and Young's modulus of the strata. Finite-difference models support this conclusion.

Jan 1, 1996

By Keith A. Heasley

Numerous different instruments and techniques have been developed for the purpose of measuring rock stress and stress changes around excavations. The interpretation of the output from many of these instruments has often been problematic. This paper covers the research performed in attempting to analyze one particular stress measuring instrument called the Borehole Platened Flatjack (BPF) . The approach used in this analysis was to combine the analytical solution for the displacements around a borehole and the analytical solution for the bore- hole-platen interaction with the empirical stiffness behavior of the flatjack. The resulting model of the BPF response was then verified using measured laboratory data with very good results. In addition, techniques for applying the method to the field are briefly discussed. The major accomplishments of this comprehensive modeling and testing program are; 1) the establishment of the critical installation and site parameters of the hydraulic pressure cell and 2) the development of a systematic procedure for calibrating the cell based on the given values of these critical parameters.

Jan 1, 1989

By Kazem Oraee

Falls of ground have historically been the main cause for fatalities in underground mines. Although recent advances in technology have reduced the number of such accidents, when failures occur they usually result in severe consequences. Risks of rock falls, use of heavy machinery and electrical apparatus, entry to confined spaces, working in noisy and dusty environments, and working on unstable platforms are some of the hazards in ground control operations. Managing these risks requires a management regime involving strict adherence to operational codes of practice and an enshrined culture of safety. These should be subsequently reinforced through active participation of management, systematic training, and stringent internal and third party auditing. Accident reports show that the major contributing factor in most rock fall accidents is the failure to adequately manage known risks due to the lack of a systematic process. OHSAS 18001 (Occupational Health and Safety Assessment Series) is an international occupational health and safety management standard specification to develop occupational health and safety at the workplace. This standard is intended to help mine operators control occupational health and safety risks. Addressing the requirements of OHSAS 18001 can be a complex and demanding task. A comprehensive guide for efficient and accurate implementation of this standard is provided in this paper. The discipline can be used to establish an accountable management system foreground control activities in underground coal mines. The paper also provides guidelines for preparing necessary documents, devising safety policies, procedures, performing risk assessment, and handling instructions. Finally, the paper concludes by providing a sound analytical basis in terms of the creation of a robust safety management system foreground control operations in underground mines. Full benefits of implementing an effective and systematic operational health and safety management system are illustrated. The procedure adopted and prescribed in this paper can be used in all underground coal mines where lack of appropriate ground control practices can create deficiencies in both safety and productivity.

Jan 1, 2012

By Stephen Tadolini

The annulus thickness, for bolts anchored with resin, effects the systems final performance. Recent developments in the manufacturing process of these support systems, coupled with improved resin characteristics, have resulted in design changes to minimize hole drilling requirements and reduce resin usage. The amount of resin to properly anchor a 7/8-SRD (Special Rolled Dimensions) bolt in a 25.4 mm (1-in) diameter hole has been reduced to about one-third of the resin required to anchor a No. 7 rebar bolt (22 mm (7/8-in) diameter) traditionally installed in a 35 mm (1-3/8-in) diameter borehole. More importantly, laboratory and field tests, indicate that higher bond strengths and anchorage capacities can be realized with reduced annulus systems. These factors have a direct bearing on reinforcement efficiency and subsequent roof support performance.

Jan 1, 1998

By Peter Zhang

Pumpable cribs are being increasingly used in longwall tailgate entries and bleeders for roof support under abutment pressure. Their high peak capacity and confinement-controlled yield characteristics from reinforced-bags make them relatively safe and reliable standing supports for highly productive longwall panels. The pumpability of crib material over a long distance from the surface greatly reduces material handling and makes the underground installation more efficient and flexible. However, the performance of pumpable cribs is dependent upon crib properties, crib pattern, quality of installation, and geological and mining conditions. For proper crib selection and safe support design, it is critical to understand the basic technical requirements and performance of pumpable cribs in longwall tailgate entries and bleeders. Based on more than 10 years of experience using pumpable cribs for longwall standing support, the requirements and performance of pumpable cribs in longwall tailgate entries and bleeders under different mining conditions are presented. Experience shows that the load-displacement curve from full-scale lab testing is a good indicator of pumpable crib suitability for certain applications. Crib peak capacity, stiffness, and residual characteristics define its applicability to certain gate entries and bleeders under different cover depth. Crib pattern and spacing can be determined by abutment pressure level, roof, floor, and rib conditions, local roof stability, and ventilation requirements. To achieve proper in-situ performance, quality control of crib installation is also important.

Jan 1, 2012

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 Doug Tesarik

Researchers from the National Institute for Occupational Safety and Health1 (NIOSH), together with employees from The Doe Run Company, are monitoring and evaluating the stability of a pillar retreat section at the Fletcher Mine, a lead/zinc, room-and-pillar mine near Bunker, MO. Cemented rockfill is being used to provide roof support as high-grade, remnant ore pillars are mined. Earth pressure cells and deformation meters placed in backfill panel pillars monitor stress change and vertical deformation, respectively, as nearby ore-bearing pillars are extracted. Similarly, biaxial stressmeters and borehole extensometers are used to monitor mining-induced stress change and deformation, respectively, in long-term support pillars. Data from these instruments indicate that when ore-bearing pillars were mined, stress was redistributed from support pillars to a backfill panel pillar. The data also indicate that the backfill and support pillars have remained stable for two years after pillar extraction was completed in the area. This research demonstrates the importance of using reliable instruments to document the geomechanical behavior of pillars and backfill. Gaining a better understanding of roof support and stress redistribution mechanisms is a fundamental part of NIOSH's goal to reduce and/or predict rockfalls, which are still among the leading causes of injuries in metal/nonmetal underground mines.

Jan 1, 2006

By Christopher Newman, Zach Agioutantis, Gabriel S. Esterhuizen

"As conventional underground deposits are continuing to be depleted, mining operations have been forced to produce at greater depths and in more geologically and geometrically challenging conditions. Over the past decade, there has been a global increase in the application of cemented paste backfill (CPB) material as a means of ground support in open stope mining operations. Although CPB has been extensively used within the industry, there are many fundamental factors affecting the design of safe and economical fill structures that are still not well understood. A critical design issue with respect to backfilled stopes is determining the stress state within the fill material as well as the surrounding rock.This paper details the development of a reliable numerical model for the simulation of stress distributions around the excavated stope area, as well as through the backfill material. Through discussions on modeling parameters and output results, one is provided with insights into the mechanisms of stress redistribution allowing for further accuracies in simulating the behavior of the CPB material as well as its interaction with the surrounding rockmass.INTRODUCTION AND BACKGROUNDModern technological innovations with respect to material processing, cementitious composition and chemical additives, and material transportation have provided the mining industry with a ground support material that greatly reduces mine waste costs, while, when used appropriately, increases reserve production and mine stability. As mining operations continue to produce in more complex geologic and geometric conditions, cemented paste backfill (CPB) has become commonly employed within primary-secondary excavation-support sequencing allowing for total extraction of the mining reserve. As shown in Figure 1, primary stopes are initially excavated and then backfilled in a two-staged pour: plug and final pours. The plug (initial pour), is initially placed at the bottom of the stope with a CPB material of heightened cementitious material. While the plug provides a strong foundation on which to build the rest of the CPB structure, it also provides protection against barricade breakthroughs (Li and Aubertin, 2009), as well as providing operations a means of underhand stope-and-fill mining through the use of an artificial sill pillar (Hughes, et al., 2011). Following placement, the plug is allowed to cure to a designated design strength. Upon completion of the plug, a less cementitious “final” pour material is used to backfill the rest of the excavated area."

Jan 1, 2018