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By Bret E. Leisemann

The geomechanical environment in which highwall mining must operate represents a hybrid between surface and underground operations. Mine design must take into consideration the complexities associated with unsupported spans, pillars with extreme aspect ratios and the instability related to the interaction between coal extraction and the highwall stability. The rock mass adjacent to the highwall, and for a significant distance in, is affected by lateral stress relief and blast damage. This paper describes a rigorous site investigation program aimed at providing some of the input parameters for the design of highwall mining layouts at the Moura Mine and focuses upon defining the nature and extent of the relaxed zone behind the highwall. The combination of borehole imaging and C factor analysis allows delineation of blast damage and relaxation of the rock mass immediately adjacent to the highwall excavation, as well as the more subtle relaxation effects deeper into the rock mass. This has led to the formulation of a proposed geomechanical model for the relaxed zone for direct input into highwall instability analysis and mine design.

Jan 1, 1993

By Erdal Akyol

The rockbolting method of support is employed widely in several countries in different of mines. Hard rock mines use this method extensively. Soft rock mines, such as coal mines, have employed standing supports (steel sets) as the main method, although some countries such as the U.S.A. and Australia have preferred rockbolting. There is increasing interest in extending the rockbolting method in coal mines. The application of rockbolting techniques in an industry which traditionally has not applied such techniques, has required the development of a systematic design approach. This design approach uses extensive monitoring, to asses the viability of the support system in differing conditions. Currently at a number of new and existing mine developments in the U.K., rockbolting is being successfully applied. Although, inevitably some localised geotechnical problems have been encountered and have needed to be taken into account. The effect of faulting on roadway conditions is discussed in this paper and a design approach has been applied using a numerical analysis method, namely the Finite Element Method. The results presented in the paper serve as a useful guide to roadway support designers, and in particular the role played by different bolts in a range of conditions.

Jan 1, 1992

By David H. Tang

Similar massive pillar failures were observed in two underground coal mines with different configurations of mine workings and overburden depths. Finite element models were used to analyze the causes and mechanisms of the pillar failure. The results of the analyses confirms the underground observation. For Case No. 1 the failure of coal pillar is mainly caused by a combined effect of smaller coal pillars under a large overburden and the interaction of multiple seam mining. For Case No. 2, the pillar failure was initiated by the weak roof rock but the ultimate pillar failure was caused by the splitting of pillars into very small size.

Jan 1, 1984

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

The challenge of planning to minimize interaction effects in multi-seam mining is compounded by the difficulties inherent in extrapolating essential geological and spatial data between sampling points. To predict the effects of ultra-close mining on lower seam entry stability, simulated mine openings were studied using finite element analysis and body loaded photoelastic models. Both symmetrical and unsymmetrical loading was used to represent stress fields generated under multi-seam conditions. Results were correlated with field data to determine stress levels at which failure could be expected to occur. Stability predictions require precise definitions of geologic and spatial data, which could not be accurately defined by interpolation between sampling points. To overcome this problem an expanded roof stability model, SESAME2, was developed to incorporate reliability concepts based on Monte Carlo sampling. The model was demonstrated in terms of variations of innerburden thickness using a case study.

Jan 1, 1994

By Yundong Ma

In this paper the dynamic process of rock deformation and failure in the roadways of the #9 coal seam, Wuhushan Coal Mine by was simulated using the software "Rock Failure Process Analysis System (REPA)" [I]. Based on the simulation results, this paper analyzes the failure process, characteristics and weighting of the overburden as the longwall retreat mining proceeds. The roadway's behavior including deformation characteristics, failure mechanisms, stress distribution in the surrounding rocks and the effects of various support systems on roadway stability were analyzed and studied. While simulating the rock failure process with REPA, the model stresses can be obtained through two approaches: (1) the stress is reflected by the grayscale of pictures in the model failure process and (2) the model stress can be obtained by muti-element information, which is one of the commands in RFPA system. Data files of stress can automatically be created, and curves of stress and displacement of each element can be displayed in each step of computation. Keywords: Rock Failure Process Analysis System (RFPA); roof control; roof to floor convergence; combined support

Jan 1, 2005

By Lorraine Kent

Comparison of in-situ stress overcore measurement results with those determined from laboratory evaluation of Acoustic Emission (AE) testing on oriented sub-core samples has been carried out for a variety of mining environments. These include UK and German Coal mines at moderate depth, shallow-level South African Coal mines and deep-level South African gold mines. Results show that the `Kaiser Effect' or 'Stress Memory-Damage is readily determinable throughout the brittle deformation process, although particular care is required in interpretation of results during the 'bedding-in' phase of testing and the region associated with unstable crack growth prior to failure. The results suggest that AE provides a useful complimentary method for comparison with existing methods of stress determination. The success of the method can, however, be dependent on both rock type and the likely magnitude of in-situ stress in comparison to the strength of the rock. Keywords: in-situ stress, acoustic emission, Kaiser Effect

Jan 1, 2002

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 K. Itakura

Information about geostructure for effective rock bolting must be obtained before setting the bolt into rock in a coal mine roadway. Especially, rock type variation and distribution of discontinuities such as layer boundaries, and separation of strata and cracks all represent indispensable knowledge. To achieve in-situ evaluation of roof rock, we developed an MWD (measurement while drilling) system using mechanical data from a drilling machine for rockbolt. System hardware detects torque, thrust, revolution, and stroke of the machine, and software analyzes the mechanical data log and displays locations of discontinuities using Neural Network techniques. Also, this system is able to estimate roof rock 3-D geostructure for varied rock types and discontinuity distribution in cases of a drill hole array arrangement in roof rock. We conducted a feasibility study for it using hydraulic drilling at roadways of the Taiheiyo Coal Mine, Japan. Simultaneously, some core drillings were also performed to confirm system analysis data. For core samples, distribution of cracks was observed and the borehole wall was investigated by OFS (optical fiber scope). Must discontinuity locations estimated from roof logging were conformed with data obtained from core inspection. However, the core sample contained a number of cracks which were not found in the mechanical data log. This indicates that additional cracks were generated during and/or after core drilling by the influence of rotational force of the machine and water supply. Furthermore, in comparison with the OFS image, resolution of the discontinuity's log estimated by this system was clarified; that is, it is difficult to detect small size cracks (hairline cracks) from a mechanical data log and the core inspection by OFS. This paper outlines the new hardware system for hydraulic drilling machine and compares the discontinuity map obtained from the software system with drilling core inspection.

Jan 1, 2001

By Al Campoli

One of the critical variables affecting resin bolt performance is the size of the resin annulus, defined as the difference between the hole and bolt radius. The vast majority of the millions of resin roof bolts installed each year in the United States employ an annulus between 0.125 and 0.188 inch. As the annulus size decreases the resin flow around the bolt becomes more difficult, increasing the resin pressure induced by bolt insertion. As the resin pressure increases the loss of resin into bedding planes and vertical fractures in the rock increases. As the annulus size increases the mixing of the resin paste and mastic becomes less efficient. Thus, the lower and upper limits of the optimum annulus size is determined by insertion pressure and mixing considerations, respectively. Previous U.S. Bureau of Mines researchers have concluded that a 0 125 inch annulus is optimal. However, others have recently reported that a 0.095 inch annulus with special rolled deformation (SRO) bolts provide superior anchorage to standard rebar with a very large 0.250 inch annulus, attributing SRD resin loss in weak ground to oversized holes. The effect of annulus size on unit length pull strength and resin loss to fractured rock was investigated. Short encapsulation pull tests with resin annuluses of 0.125 and 0.095 inch resisted an average of 2.25 tons per grouted inch, with no significant variation. However, annulus size has shown a dramatic effect on the magnitude of resin lost to simulated rock fractures. No. 5 standard . No. 6 standard, and No 7 J-Bar rebar were inserted into two different viscosity resin cartridges designed for 4 foot boll encapsulation into a one inch inside diameter metal tube, To simulate a rock fracture, a 0.250 inch diameter orifice was located near the capped end of the tube Resin loses increased as annulus size decreased, with a minimum loss of 35% and a maximum loss of 60% associated with 0.125 and 0.095 inch annuluses, respectively.

Jan 1, 1999

By Daniel W. H. Su

"This paper presents results from a recent study conducted to evaluate the effects of longwall-induced stresses and deformations on the mechanical integrity of shale gas wells drilled over a longwall abutment pillar. The primary objective is to demonstrate that a properly constructed gas well in a standard longwall mining abutment pillar can maintain mechanical integrity during and after mining operations. Successful designs would allow for the safe recovery of both coal and gas reserves. Major elements of this comprehensive study include: (1) modeling of rock movements to predict impacts on gas wells, (2) measurements of actual rock movements at the surface and in the subsurface, (3) measurements of actual impacts on test wells designed similarly to real shale gas wells, ( 4) comparison of modeling results to actual rock movements to validate the model, and (5) employing the validated models to evaluate various casing and cementing designs to minimize casing deformations and absorb longwall-induced strains along the intermediate and production casings.A study site was selected over a southwestern Pennsylvania coal mine, which extracts 1,500-ft-wide longwall faces under about 600 feet of cover. Four test wells and four monitoring wells were drilled and installed over an abutment pillar with 125-ft by 275-ft centers. The four test wells were drilled to a depth of 642 ft, or 42 ft below the Pittsburgh seam, and were designed like a real shale gas well, except that only three casings (namely the surface, water protection and coal protection casings) were installed in each well. Casing steel grades, wellbore diameters and annular spaces were purposely varied in each of the four test wells to evaluate the effect of different well designs. The four monitoring wells - one vertical extensometer and three in-place inclinometers (IPI) - were drilled to different depths to measure the vertical settlement and horizontal displacements as a result of longwall mining on both sides of the abutment pillar. In addition to the test wells and monitoring wells, surface subsidence measurements and underground coal pillar pressure measurements were conducted as the 1,500-ft-wide longwall panels on the south and north sides of the abutment pillar mined by."

Jan 1, 2016

Since the initial development and trials of FLEXIBOLT flexible roof bolts, further comparative testing and product development has been completed as part of the changeover to this roof support system at two Australian longwall operations, Ellalong and Angus Place Collieries. Details of the practical implementation of FLEXIBOLTS at both collieries are outlined together with their impact on roadway development rate productivity. At Ellalong where fully encapsulated 2.1 m long bolts are used, the reduction in bolt density is discussed with respect to both support system design and work force acceptance. The impact of shear deformation on bolt loading is described. Bolts longer than the working height have been required at Angus Place to achieve a suitable bolt anchorage horizon. Details of the long hole drilling and the benefits of a two stage grouting procedure are outlined. Improvements in roof conditions which have resulted from the inclusion of FLEXIBOLTS into the overall roof support strategy at Angus Place are described. Their importance to the viability of the Angus Place operation is also highlighted.

Jan 1, 1994

By Charlie Li

A cut-and-fill mine stope, located at a depth of about 1000 m in a lead-and-zinc mine, was closed down because of large roof convergence and fractures appearing on the hanging wall. The instability problem was owing to the relatively low strength of the rock subjected to high in-situ stresses. Extra reinforcement was required to secure the stope so that the remaining ores in the stope could be mined. Assessment of the fracture mode in the country rock surrounding the stope was carried out and then the reinforcement de- sign was worked out. It was proposed that the unstable section of the stope would be reinforced by bolt-shotcrete ribs as well as selective long bolts. The philosophy of the design was to hold up the fractured country rock by bolts and shotcrete so that a load-bearing arch could be formed in the roof and in the hanging wall. Six shot- Crete ribs were set up in the unstable section of the stope. Displacement measurements showed that the roof convergence reduced from about 2 &day to the ordinary creeping level, about 0.25 &day, immediately after the reinforcement operation. Six month later, the hanging wall collapsed into the stope in a place just outside the reinforced section, which indirectly proved the effectiveness of the bolt-shotcrete ribs in reinforcing the fractured country rock. This case showed that an appropriate rock mechanics assessment would form a solid base for reinforcement/support design. It also provided an evidence for the philosophy of the reinforcement by bolt-shotcrete ribs - establish a load-bearing arch in the rock.

Jan 1, 2005

By M. G. Schuerger

The United States Steel Mining Company, Inc., (USM) Lynch District, operated the No. 37 Mine in the Harlan coal seam, Harlan County, Kentucky, from 19'2 until its sale to Arch of Kentucky in 1984. Overburden depth varied from 400 to 2000 feet and mining height varied from 8 to 12 feet. In 1981, because of deep overburden, high seam height, and previous pillar stability problems, U. S. Steel Research and the Lynch Mining District entered into a joint project to measure longwall pillar stress. Eight gateroad pillars were instrumented with a total of 86 vibrating-wire stressmeters under overburden depths varying from 500 to 1550 feet. Stress-increase measurements were taken during: (1) pillar development, (2) retreat of longwall panel No. 1, and (3) retreat of longwall panel No. 2. Total average pillar stress was approximately three times the pre-mining stress with total stress at an individual meter reading 5.5 times the pre-mining stress. When maximum, measured average pillar stress was compared to pillar strength predicted by several popular pillar design methods, the pillars were found to be stronger than predicted. This paper will summarize the pillar instrumentation program and present the results. Stress distributions, average measured pillar stress, and comparison of measured pillar stress to predicted pillar strength by several design methods will be given.

Jan 1, 1984

By Robert Kimutis, Ke Yang, Yi Luo, Jianwei Cheng

"Surface subsidence processes, especially those associated with a longwall mining operations, could cause operational and/or safety problems to railroads. For railroads with normal amounts of traffic, both the limited time window between traffic and the limited space available make the efforts to return the railroad back to its original state nearly impossible. In order to ensure the safe operation of a subsiding railway, the partial lifting method has been developed and applied. This paper describes the principle and application of the method. A case of applying this method to mitigate subsidence effects on a long section of railway has been demonstrated.INTRODUCTIONMaintaining an operational railway as it is experiencing a longwall subsidence process is not an easy task. The subsidence process may induce sufficient strain and curvature to cause problems to the railroad structures. Localized subsidence-induced slope often is significantly larger than the permissible ruling gradient of 0.7% for a locomotive to haul a loaded train reliably. In order to ensure safe operation, it would be ideal to instantly lift the railroad track back to its original level as it is subsiding. However, it is often impractical to do so due to the limited time windows available between railway traffic and the available space on the railroad base.A partial lifting method has been developed to keep the railroad operational under such limitations. The subsiding railroad track is raised by only a determined, partial amount of the occurred subsidence at a given time. By making such a determined adjustment, the railroad track can be maintained on a smooth and operational profile, and the high strain and curvature on the railroad structures can be reduced as well. To define a smooth railroad track profile in a subsided section the following two conditions should be met: (1) the maximum railroad gradient should be less than the permissible ruling gradient, and (2) the recommended amount of adjustment should be able to be made by the repair crew within the available time windows between scheduled railway traffic.The subsidence effects on a long section of railroad over two longwall panels have been successfully mitigated using the partial lifting method. The railroad experienced a maximum subsidence 89 of about 5 .0 ft (1.5 m) with the average subsidence along the entire length being about 3.45 ft (1.05 m). However, the average adjustment due to lifting that was made along the entire length of the subsidence-influenced railway was only 0.5 ft (152 mm)."

Jan 1, 2010

By Zach G. Agioutantis

The prediction of ground movements due to underground mining using the influence function method is a mature technology, widely used by researchers and planning engineers around the world. Surface strain, either tensile or compressive, is an important indicator of potential impacts on structures such as buildings, bodies of water, pipelines, railway and power lines, and tailings dams. Calculation of static or dynamic surface strains, for varying surface terrains and slopes, should be accurately predicted and be independent of the methodology applied for the calculation. This paper discusses and compares the development of horizontal and ground strains on the surface due to underground mining for different surface topologies (sloping terrain, monitoring lines, etc.) and presents the steps needed to ensure accurate calculations. Examples are given to demonstrate how ground strains can be calculated for one- and two-dimensional gridded surfaces.

Jan 1, 2013

By Qiang Fu

ABSTRACT This paper aims to identify the main issues concerning top-coal loss and the optimum drawing interval of LTCC (Longwall Top-coal Caving) mining by underground tests and DEM (Distinct Element Method) modeling. By extracting a new roadway in the goaf of the mined LTCC Panel No.4318 in Wang Zhuang coal mine of Lu-an CMA (Coal Mining Administration), and underground observation of the distribution of loss coal and broken rock in different sections of the roadway, this paper analyses the composition and presents a 3-D distribution map of the loss coal in the exploration roadway. This study indicates that the DEM modeling of broken top-coal is reasonable and scientific as it can simulate the movement of broken coal and waste rocks and determine the optimum drawing interval of top coal.

Jan 1, 2002

By Joe Zelanko

In order to prevent roof fall accidents and injuries, U.S. underground coal mine operators must develop and implement effective roof control strategies. Various tools are available to assist engineers in developing designs. For example, criteria to address global stability issues (e.g., pillar stability) are well established in the U.S.; effective tools (e.g., empirical and numerical) are available to diagnose potential problems and avoid catastrophic failures. However, most miner injuries occur as a result of local instability (e.g., a roof fall in a single heading or a rib roll from a pillar corner). Hazard recognition and the availability of support plans suitable to those conditions are key to maintaining stability and safety on the working section.

Jan 1, 2011

By Zhijie Zhu

In order to determine the overburden failure characteristics of extra-thick seam mining, EH4 electromagnetic image system and borehole televiewer survey were applied to panel 8100, Tongxin coal mine. EH4 electromagnetic imaging system was used to map the overburden strata before and after mining. Analysis of the distribution of the electrical conductivity showed that the fractured zone height was 492-558 ft (150-170m); Observation through the borehole televiewer showed that the fractured zone height of water conductivity (or water flowing fractured zone height) was 110m-142m. The results of the two methods showed that the fractured zone height of water conductivity was 492-558 ft (150- 170m) or 10-11.3 times the mining height.

Jan 1, 2014

By Vasundhara

Underground coal mining in soft floor environments has often led to strata control problems ranging from small scale issues such as floor heave which disrupts traffic and ventilation, rib spalling and induced fractures in pillars, through to large scale instability and long term subsidence. Past investigations of the problems encountered in Newcastle Coalfields of New South Wales, Australia, have revealed that the soft floor unit comprises of claystone which is very sensitive to moisture. It is prone to swelling, reduction in stiffness and loss of strength in the presence of moisture. Long term stability of mine panels in such an environment is governed by the extent of deterioration of the floor strata. The paper reports the field investigations conducted at Cooranbong Colliery to assess the extent of softened floor zone in the roadway and under a pillar.

Jan 1, 2001