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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 Eric Westman

Ground control problems caused by multiple-seam interaction frequently sterilize coal reserves. Interaction mechanisms differ for overmining and undermining and are controlled by mining, spatial. and geologic conditions at each site. A multiple-seam interaction analysis software program, UGLY (Uppersearn Gateroad Longwall Stability), has been used to evaluate sterilization of coal reserves due to overmining at three locations. Four case histories were reviewed to determine the effectiveness of improved ground control in mitigating interaction effects.

Jan 1, 1997

By Steve P. Signer

A study of rock support interaction mechanics was conducted by the U.S. Bureau of Mines at the Jim Walter Resources No. 7 Mine, Brookwood, AL. Bolt loads and entry deformation were monitored at four sites in a tailgate roadway. Two sites were areas supported by fully grouted resin bolts, and two sites were supported by tensioned bolts that used resin-assisted mechanical anchors. The yield capacity of the tensioned bolts was 37 pct higher than that of the grouted bolts. To monitor loads, instruments were mounted on 16 bolts, which were installed during entry development in the standard bolt pattern. Additional instrumented bolts were installed as supplemental support just prior to the headgate pass of the first longwall. Instruments to measure roof movements included three-point roof extensometers and convergence stations. Readings were taken during section development and as each longwall panel passed the test area. The fully grouted bolts showed localized loads that were higher than loads on the tensioned bolts; however, overall loading magnitudes were similar. Many sections of the fully grouted bolts passed the yield point of the bolt steel (98 kN or 22,000 lb). However, the maximum strain measured at any loca¬tion was 50 pct of the ultimate strain, and there was adequate bolt length to transfer the high loads into the rock. Significant bending was detected at many points along the length of the bolts and seemed to be the most significant mechanism in support-rock interaction. Loads on the tensioned bolts did not exceed the yield point of 160 kN (35,000 lb). Entry deforma¬tion, supplemental bolt loading, and general entry degradation appeared to be slightly less at the tensioned bolt sites. Local geology and in situ horizontal stress were also observed to affect con¬ditions at the test sites.

Jan 1, 1993

By I. L. Follington

A breaker line support (BLS) monitoring system, BLSmon, has been designed and constructed by the CSIRO and installed and commissioned at Laleham No.1 Colliery, Queensland. This system was designed to record hydraulic leg pressure and canopy position measurements and relay these, in real time, to the surface. The system has shown itself to be capable of operating on the BLSs without any adverse affect on production, Analysis of the results of the monitoring exercise has shown that rate of change of leg pressures on the BLSs could be used as an aid in the prediction of adverse mining conditions. Under favourable mining conditions the supports were observed to initially unload from the set pressure, this unloading reduced with time and by the end of a typical lift cycle had either ceased or reversed slightly. This behaviour was considered to be due to one of four possible causes, these being; the compaction of floor debris, and/or compaction of the floor, and/or crushing of the roof and /or support creep. The actual cause can only be determined through detailed monitoring of both roof and support convergence. However observational data would suggest that compaction of debris and softened floor material beneath the supports was the most likely cause. The few positive loading rates recorded under these conditions were of short duration and occurred towards the end of individual lifts. Where mining conditions were unfavourable the negative rates of change of pressure, unloading of the support, were reduced throughout a lift, in both duration and magnitude, when compared with favourable mining conditions. Over the same period the positive rates of change were observed to be greater. The rate of change of loading and convergence can be utilised to identify the onset of instability in the lift area. However, before this facility can be utilised by the mine for the reliable prediction of instability more monitored data and sophisticated processing techniques are required. A mechanism has been postulated for the deformation behaviour of the roof and pillars in a mechanised pillar extraction operation from the continuously monitored data, observational data and results of previous studies. The 7 m wide fenders were observed to yield under abutment stresses before lifting commenced. The stability of these fenders throughout the lifting sequence was considered to be dependant upon the magnitude of their post peak load carrying capability. This is influenced by the post peak stiffness and the extent of roof lowering in the vicinity of the fender. Difficulties with fender extraction occurred where remnant coal left standing in the goaf caused uncaved goat to bridge to the mined fender. Difficulties were also experienced where sub-vertical jointing occurred in the roof of the split reducing it's ability to bridge effectively. Despite the relatively short monitoring period, the potential role and application of continuous monitoring in advancing understanding of the rock mass response to mechanised pillar extraction operations has been demonstrated. This understanding now requires enhancement and shows great potential for aiding the optimisation of mechanised pillar extraction operations.

Jan 1, 1992

By K. Y. Haramy

Rock burst and coal mine bump research using static and dynamic rock mechanics instrumentation has been conducted for several decades. Research efforts typically have been conducted using static instrumentation such as pressure cells, convergence measurements, multipoint borehole extensometers, and shield loading pressure measurements; dynamic rock mechanics instrumentation, for example, microseismic monitoring, has been used to help provide precursors to those violent failures. Unfortunately, these different styles of study have never been used in tandem. We report in this study the unique opportunity to have both static and dynamic instrumentation available for monitoring a longwall panel in a deep western coal mine and show that the two methods used in tandem provide information to understand the causes of a major bump. Several static rock mechanics instrumentation sites were carefully chosen along the longwall panel in the coal seam, packwalls, roof, and floor to provide the stress history during mining, including the stress redistribution that occurred before and after the major failures. Monitoring of the shield pressure on selected shields of the longwall face revealed that the cyclical loading anticipated during normal caving processes was not occurring. In addition, by locating each microseismic event (rock noise), the areas that were not adjusting to the changing stress field were clearly delineated by a lack of microseismic activity. The area of the longwall panel that did not adjust to the changing conditions ultimately failed. The combination of static and dynamic rock mechanics instrumentation data analysis provided the necessary information to determine the interaction of the structural members and the cause of the major failures at this mine.

Jan 1, 1988

By P. A. Gray

Artificial ground support has developed into a relined science over the past 20 years. From reactive support systems such as timber props and steel supports, to active systems such as roof bolts and cable anchors, ground support technology has now developed efficient systems that work well in many instances. However this paper demonstrates that there is still a long way to go to develop optimum ground support systems. Some conventional rock bolts and cable anchors emphasise their tensile capacity only, without considering other factors such as their shear strength or shear modulus, or if the support member can transfer all of its support capacity to the surrounding rock. Examples are given of the use of ground anchors with high tensile and shear capacity in underground coal mining as well as examples of slope failures in open pits where the cable bolts have remained intact because they have been unable to transfer their full capacity to the surrounding weak rock mass. For underground coal mines, speed of installation is also of primary concern, but it must also have high tensile and shear capacity both in terms of strength and stiffness. Ground support systems of the future must therefore address these problems. An innovative new rock bait is discussed which goes some way towards solving these problems. This new rock bolt is screwed into the rock and thus has a direct mechanical interlock with the rock. Initial field results show that it has an extremely high anchor capacity and can be fully installed in approximately 30 seconds. It has considerable capacity to improve the productivity of the coal mining industry.

Jan 1, 1992

By Brian E. Travis

Recent Developments in the utilization of polymer grid structures for Longwall Shield Recovery operations are discussed. Specifically: various Longwall screen designs and the numerous methods employed during installation. Also discussed is the role of polymer grid structures in head and tailgate mf and rib control and their role in controlling soft bottom often associated in shield recovery operations.

Jan 1, 1993

By Yoshiaki Fujii

Displacement Discontinuity Method (DDM), which was originally developed by Crouch,S.L., is a powerful tool to calculate mechanical disturbances due to tabular excavations. Three topics on DDM applications to strata control problems in Japanese deep coal mines are described. First, is numerical simulation of microseismicity induced by longwall mining in Horonai Coal Mine. Microseismicity in the mine had been monitored by a mine-wide seismic array. Location of microseismic events and intensity of seismicity were well simulated. In the simulation procedure based on DDM, first, zones where a failure condition is satisfied by the advance of mining are detected. Then, the seismic moments, which indicate the intensity of seismicity, at the failure zones are evaluated. Second, is prediction of the shaft damage due to mining in Sumitomo Akabira Colliery. In this mine, the limit angle of the shaft pillar was changed from 600 to 800 to increase the amount of minable coal volume without increasing mining depth. Displacement of the shaft axis has been measured to assess the damage due to mining around the shaft pillar. In the calculation based on DDM, a fault intersecting the shaft and the ground surface are considered. Observed and predicted displacement of the shaft axis agreed well. Third, is the roadway deformation in Taiheiyo Coal Mine. Attainable maximum stress for a roadway was calculated considering three faults intersecting the roadway. It was realized that the roadway's convergence is related to the calculated stress by simple equations.

Jan 1, 1992

By Stephen P. Signer

This paper presents the results of a case study conducted in a two-entry gateroad in a coal mine where excessive roof deformation and bolt loading resulted in failure of many roof supports. The instruments consisted of 16 fully grouted, strain gaged resin bolts, load cells on both full and partially grouted cable bolts, and vertical deflection multipoint extensometers. These instruments were installed on-cycle to measure loading on bolts in an existing support pattern and to determine if cable bolts could be used to improve roof stability. Results showed significant amounts of movement within the bolted zone. In sonic cases, these movements were enough to load the supports past their ultimate strength. Geology appeared to be a significant factor in localized roof degradation; that is, a very weak rock layer in the immediate roof overlain by a very strong rock layer contributed to the development of shear planes. Only one instrumented bolt had a maximum strain of less than 2,000 microstrain, which was the yield point of the steel. The average maximum strain on all bolts was 20.000 microstrain. However, electrical continuity to many gages was lost, and a pattern was noted that would indicate possible bolt failure. Wire mesh and concrete cans installed as secondary support performed very effectively.

Jan 1, 1998

By E. Bane Kroeger

For many years, researchers around the world have been investigating coal pillar stability. Many have focused on trying to optimize the size of the pillars by examining stable and failed pillars in underground coal mines. For mines that are already in operation or companies that are opening a new mine adjacent to an existing mine where the pillar dimensions and stability can easily be determined, this is the preferred technique. However, when a new mine is planned for an area where there has been relatively little mining, then samples of the coal need to be drilled and tested for strength. Because of the fractured nature of coal, this lab testing is often problematic and may often yield poor results because of the limited size and number of samples that can be obtained from the core. To remedy this problem, research was conducted in the labs at SIUC to determine if improvements could be made when optimizing the size of pillars for underground coal mining in Illinois. Through a greater understanding of the relationship between dimensions and strengths of coal samples tested in the lab, it was hoped the in situ strength of coal pillars could be determined with greater accuracy. This could allow the extraction ratio to be increased and more of the coal resources in Illinois recovered without impacting the safety of the mine workers. Many different agencies around the world have published recommended minimum numbers of samples for strength determination based on the size of the coal samples. For two inch cubes, uniaxial testing of at least seven samples is recommended and this minimum number decreases to four when testing four inch cubes. However, the laboratory testing of coal from two seams thus far has suggested this number should be almost doubled to accurately determine the coal strength. Testing also showed there is a pronounced decrease in the variation in the strength of the samples tested as the dimensions of the sample were increased. From the testing results thus far, it is recommended that a sample size versus strength curve be constructed for every coal seam for Illinois. Creating such a curve will better determine both the minimum number of samples and minimum sample dimensions for that seam or region of a particular scam. A second set of tests were conducted in the laboratory to determine the increase in strength with reduction in height for samples with the same width and length. In Illinois, typical pillar dimensions are 40 to 80 feet (12.2 to 24.4 m) wide and long, with coal thickness varying from about 4 to 9 feet (1.22 to 2.74 m). This provides width/height ratios ranging from 4.5 to 17.5. Coal samples having width/height ratios ranging from 0.5 to 13 have been tested thus far. As with most pillar design formulae, the strength of the Murphysboro coal versus width/height ratio was linear and closely matched shape factors published by other researchers. The laboratory testing thus far has identified three factors that affect the strength of laboratory samples, the number of discontinuities, the angles of critical discontinuities, and the shape of the samples. The factor that probably had the most pronounced effect on the strength of the coal samples was the angle of critical discontinuities. This testing has confirmed that there is still a gap that needs to be bridged between the strength obtained from laboratory testing and the in situ coal strength predicted.

Jan 1, 2004

By William B. Beerbower

When full length resin grouted bolts were introduced into U. S. coal mines some 11 years ago, it was assumed that safety would be the primary benefit. There is ample testimony to substantiate the safety factor as an important reason for using resin roof bolts. More recently, it has been determined that many coal operators not only concur with the safety consideration but also benefit economically from the safer working conditions resulting from a reduction in roof falls. IMPROVED SAFETY Some typical roof fall reduction experiences with resin bolting follow: ? Western Kentucky Mine - Resin bolting reduced roof falls by 80%. ? Southern West Virginia Mine - Experienced 141 roof falls between 1971 and 1972. In 1973 with resin bolting, roof falls were reduced 85%. ? Western Pennsylvania Mine - In 1973 had 59 roof falls with mechanical bolts. During the same period of 1974 with resin, only 19 falls were experienced - a reduction of 66%. The overall experience with resin bolted roof support has been an average 50% reduction in roof falls. Long term, this can translate into improved worker morale. As a matter of fact, a general mine foreman at the same Southern West Virginia mine mentioned above commented that improved miner morale was noticeable following the introduction of resin bolting. HOW GREATER SAFETY CAN BE COMBINED WITH INCREASED PRODUCTIVITY IN COAL MINING OPERATIONS Experience has also shown that a reduction of roof falls with resin bolting has resulted in increased production for many coal operators. Some typical examples with mines using resin bolting follow: ? Western Kentucky Mine - Resin bolting contributed to increasing production from 7000 to 11,000 tons per day, due to a 80% reduction in roof falls. This was accomplished in part by a reduction of 50% in timbering requirements and the reassignment of three roof fall cleanup crews to coal production. ? Pennsylvania Mine - Following resin bolting introduction the section showed a 50% increase in production. ? Central West Virginia Mine - Continuous miner production averaged 6 cuts per unit shift prior to resin bolting. Reduction of cleanup time, resulting from improved roof support with resin bolting, increased production to an average of 9 cuts per unit shift. PROPERTIES OF RESIN BOLTING SYSTEMS Several of the important properties of resin bolting systems which contribute to its improved performance are listed below: ? Full length anchored resin bolts and many resin anchored, point anchored bolts provide effective resistance to changing rock strata stresses throughout their length, thereby improving stabilization of rock strata over long periods of time. ? Individual layers of rock are anchored into a single high strength beam providing excellent resistance to vertical and lateral strata movement, the most common cause of roof failure. ? Moisture cannot corrode the fully grouted bolt. ? Has twice the pull strength of mechanical bolts (See Figure 1)

Jan 1, 1982

By Z. T. Bieniawski

A major study was performed on determination of long term strength of coal strata directed to the design of coal pillars to ensure their stability beyond the life-time of a mine. Based on field investigations, numerical modeling and reliable data from 12 case histories, three aspects of time effect in the strength of coal strata were studied; (1) strength reduction based on the concept of volumetric strain estimate from in-situ tests, (2) strength reduction in terms of three failure criteria applicable to coal strata: original Hoek-Brown criterion, modified Hoek-Brown criterion, and Generalized criterion, and (3) long-term strength estimation as a function of coal strata quality RMR based on case studies. From these investigations, an integrated approach to long-term pillar stability is proposed, based on numerical modeling and empirical pillar strength formulation. The values of factors of safety for coal pillars are analyzed in terms of probability of failure and practical meaning for mine design purposes.

Jan 1, 1994

By Yang Li

For the study of overburden movement in longwall mining of shallow coal seams, most scholars merely consider the thickness of bedrock and top soils. In this paper, the overburden movement in longwall mining of shallow coal seams with thin top soils and key strata group was studied. By using numerical simulation, the development of fractures in the overburden was analyzed for different ratios of panel width to depth (W/D) and depth to mining height. The results show that the overburden movement changed with the variation of mining height, panel width and depth under the specific geological conditions. It also explain how different overburden disturbance structures, such as ?three zones? (caved, fractured, and continuous deformation zones), ?two zones?(caved and fractured zones) or ?steps-like collapse? are formed under different mining conditions. The results broaden the vision for research on overburden movement in longwall mining of shallow coal seams.

Jan 1, 2014

By Eric Keller, Jamal Rostami, Wenpeng Liu

"Roof and rib instability is an important issue in underground mining. To optimize ground support design, enhance ground stability, and reduce the possibility of roof or rib failure with minimal use of artificial ground support, it is essential to have an accurate understanding of ground conditions. This includes the location of voids, cracks, and discontinuities, as well as information about the different strata in the immediate roof. This paper briefly introduces ongoing research on void detection by using the roof bolter feed and rotation pressure. The goal of this project is to improve the sensitivity of detection programs to locate smaller joints and reduce the number of false alarms. This paper has presented a brief review of the testing procedures, data analysis, logic, and algorithms used for void detection. In addition, this paper has discussed the results of preliminary laboratory tests and statistical analysis of the data from these two drilling parameters used for void detection.INTRODUCTIONRoof falls are one of the most serious and frequent accidents in underground mining. Each year, personnel are injured or killed because of roof falls. Equipment can also be damaged when this occurs. According to Mine Safety and Health Administration statistics collected from 1999 to 2008, ground falls were the largest cause of fatal accidents in underground coal mining, causing around 40% of all fatalities (NIOSH, 2015a). Out of 8 to 10 fatalities and over 800 injuries are recorded each year in underground mines, there are nearly 2,000 reportable non-injury ground falls every year (NIOSH, 2015b).Certain features in the ground, such as voids, cracks, and discontinuities, are significant factors that cause roof support failures and roof fall accidents. Currently, several manufacturers provide smart roof bolters with limited capabilities for void detection. While successful in many instances, in laboratory experiments, the existing void detection systems have limited success in detecting void openings that are less than 1/8 inch This limited success was observed in testing concrete blocks with gaps in between, simulating a rock medium with voids. Detection programs have missed some voids, while, in some cases, they have shown false detection. These algorithms need to be more accurate in sensing voids to improve the capabilities of these programs."

Jan 1, 2016

By Axel Preusse

Stowing or waste fill is common in the German mining industry since the beginning of mining activities, With increasing mechanization, flat and gently inclined seams (0 -- 18) were mined. These mines were located within the densely populated areas close to important infrastructures such as industrial complexes, roads, highways, railroads, rivers and canals. These buildings and infra¬structures were rarely protected against subsidence influences and other ground movements and dump capacities close to the mines were exhausted. For subsidence reasons and because of the low dump capacities pneumatic stowing methods were continuously developed and used in German longwalls until 1997. Due to techni¬cal standards and because of some restrictions pneumatic stowing in Germany is only used at a minimum seam thickness of 2.00 m and average costs of this method are $20 - 30 per ton. A significant reduction of the disturbance potential and by that a decrease in expenses for subsidence regulations due to reduced subsidence and ground movements could not be achieved. In this paper the experiences with pneumatic stowing in the Ger¬man mining industry are described. Possible subsidence reduction and economic aspects of pneumatic stowing are taken into account.

Jan 1, 2002

By Daniel W. Su

This paper presents the implementation and evaluation of the hydraulic fracturing technique and Longwall Visual Analysis (LVA) software to mitigate the impact of a 1,000-ft (305-meter) widemassive sandstone channel on a 1,500-ft-wide (457-m-wide) longwall face. Based on a underground roof geology reconnaissance program, four frac holes were drilled and fraced along the center axis of the sandstone channel. To further provide detailed monitoring of the longwall face, the Longwall Visual Analysis (LVA) software was installed to track the face pressure and cavity formation index. In mid-December 2011, the longwall face approached and entered into the western edge of the sandstone channel. The longwall face mined through the center axis of the sandstone channel in mid-January 2012 and advanced outby the eastern edge of the sandstone channel in early February 2012. Results from daily underground observations and production delay analysis confirmed the effectiveness of the hydraulic fracturing program and the validity of the LVA Software.

Jan 1, 2012

By Karl H. Brandt

In the Auguste Victoria mine a bolted rectangular roadway has been kept open for the first time in the seam D/C after retreat. The target of the project was to gain a basis for planning and experience for future mining layouts in the same scam in view of the reuse of boated rectangular roadways. For this the roadway was additionally secured in sections with longbolts before the influence of the retreating panel had taken place. On retreat a concrete sidepack was concurrently put in place on the face side. The installed monitoring systems consisted of endoscopic devices, multiple extensometers (Tell-Tales), shear-tales, extensometers and convergence measurements. The experience that was gained was carried over into a physical and numerical model.

Jan 1, 2001

By Hanjie Chen

A mine panel collapse may occur when pillar sizes are too small or the surrounding rock strata (roof or floor) yields. After a small pillar fails, its loading is rapidly transferred to adjacent pillars that in turn fail. If the surrounding rock strata yield, the loading on pillars will redistribute. which may result in some pillars under high stress. This paper analyzes the causes of two cases of mine panels failures in two room and pillar mines. One case of massive pillar failure occurred in a phosphate mine and the other, mine panels failure, is in a coal mine. In the phosphate mine, both the roof and floor strata were a thick competent shale stratum. The finite element analysis shows that the major reason for the massive pillar failure is that the pillar sizes were too small in a panel. In addition, the high cliff topography has some influence on the failure. In the coal mine, the failed area was about 1,000 ft long by 760 ft wide (covering more than 3 panels) involving 390 pillars. The harrier pillars between the panels were punched on retreat into small chain pillars and the extraction ratio in this failed area was more than 80%. The roof remained intact, and pillars were crushed and punched into the floor. Based on the stress analysis, it is found that multiple panel collapse occurred because the barrier pillars were punched into small pillars and the loading on panel edges increased due to floor failure.

Jan 1, 1999

By Salah Badr

Longwall mine layouts with their entries, chain pillars, face support systems, advancing mining faces and compacting gobs represent a geomechanically complex mining operation. Geomechanical aspects are further complicated if mining occurs at depths exceeding 350 m, as stresses induced adjacent to the excavations exceed the unconfined strength of the coal and result in fracturing of the sidewalls of entries and chain pillars during development. Traditional methods of geomechanical analysis, whether empirical or numerical, are unable to represent such behavior satisfactorily. Achieving some meaningful results requires explicit representation of the three-dimensional boundary conditions and implementation of appropriate constitutive models for the rock mass. This paper describes a numerical modeling methodology for deep longwall mines operation using the FLAC3D finite difference de, which incorporates a strain softening constitutive law for modeling failure of coal. An important aspect of the modeling method includes accounting for stress relief provided to the pillars and entries due to the increasing load applied to the compacting gob.

Jan 1, 2003

Effective strata control, utilising fully encapsulated roof bolts, is dependent on the installed quality of the reinforcement elements. One mechanism by which fully encapsulated roof bolts may become less than fully effective is by gloving (glove fingering) and un-mixing of the resin. Following some small roof failures containing gloved bolts, investigations have been undertaken in some New Zealand coal mines to determine the extent of, and mechanisms involved in, the gloving of fully encapsulated roof bolts for a range of roof types and installation methods. Gloving, in this context, refers to the resin cartridge partially or completely enveloping a length of the bolt, typically with unmixed resin filler and catalyst/hardener remaining within the cartridge. Results have shown that gloving and un-mixing is a systematic and widespread phenomena, occurring across a range of resin and/or bolt manufacturers, and across a variety of roof types. Gloving was found in bolts installed using: hand held pneumatic and continuous miner mounted hydraulic methods, under run of mine (ROM) conditions by mine workers, and under controlled "best practice" conditions. The roofs bolts investigated are all 1800mm (6ft) long, and have shown that, on average, 485mm (-19") of the bolt length is gloved and/or unmixed. This equates to a 27 percent reduction in effective bolt length. The effected length of bolt is sufficient to impact on the load transfer/reinforcement capabilities of the bolt system, and hence impacts on the design assumptions and subsequent stability assessments. Furthermore, there is a direct relationship between the mechanism for the gloving of bolts and an increased occurrence of resin loss and reduction in encapsulations length in soft (less than 10 MPa) and/or highly fractured rock. Investigations have concentrated on developing an understanding of the mechanisms involved. This has been undertaken using both field installations and recovered bolts, as well as test bench trials. The mechanisms involved have been confirmed as being: the development of a pressure front as the bolt encounters the resin and is spun up the hole, which, in turn, leads to over-pressurisation and radial expansion of the resin cartridge, resulting in an increase in diameter of the plastic resin cartridge, which allows the bolt to be spun up into the cartridge without making sufficient contact to shred the cartridge or the hardener tube, resulting in an unmixed portion of resin.

Jan 1, 2003