Search Documents

By Wei Wu

A new bench blast design method and its realization with the assistance of the computer is introduced. The basic design principle is to define a rhomb cell, called fully-fractured zone, through a series of full scale bench crater tests and apply this geometric pattern in multi-hole, multi-row blast design. To illustrate the design method, the optimization process and assessment of blast performance are presented in the context of a case study.

Jan 1, 1992

By Russell C. Frith

Risk-based roadway roof support design is now a critical part of the Australian Coal Industry. Safe and efficient mining demands that roof support be tailored to the prevailing geotechnical conditions and coal mining legislation in New South Wales, Queensland, and Tasmania is clear in requiring that formalized roof support design be undertaken. The assumption of dead-load suspension of an otherwise unstable roof has been used in roadway roof support design for many years. However longevity does not necessarily translate into either a best or even reasonable practice in the current industry, particularly when pro-active strata management systems are now routinely used and reinforcing support design methods are available. The paper discusses why the assumption of dead-load suspension is fundamentally incorrect in almost all instances when pro-active or reinforcing roof support is applied and how it can easily result in misleading and potentially under-designed roof support systems under various circumstances. The commonly used design methodology of balancing the installed axial capacity of long tendons with the assumed ?weight? of a future roof fall is summarized, and several fundamental flaws are identified. The critical area of selecting the design Factor of Safety is also discussed in detail. Accepting that under certain circumstances the roof of a roadway will require it to be ?suspended?, suggestions are given for how a robust support system can be developed, designed and applied.

Jan 1, 2011

By Terry Medhurst

Geophysical logging is routinely undertaken as part of most Australian coal mining exploration programs. Typically, the main application is the determination of coal seam depth and the qualitative estimation of coal quality, lithology and rock strength. For rock strength estimation, sonic transit times relate to the elastic properties of the rock and provide a general correlation with uniaxial compressive strength. However, experience has shown that site-specific relationships are necessary to account for different geological environments. In this paper we describe results from a series of recent studies that have provided a more comprehensive method of geotechnical strata characterisation using detailed lithological interpretation and rock property estimation from borehole logs. This has resulted in an improved method of estimating rock strength via the determination of the proportions of quartz and clay plus the porosity. We also propose a rock classification scheme based on geophysical log analysis. This builds on the existing acceptance of the sonic log as an indicator of rock strength and adds scores according to the moisture, cohesion, porosity and quartz content inferred from the geophysical logs. The new scheme can be related to existing mapping based schemes such as CMRR and the Q-system.

Jan 1, 2005

By Raj R. Kallu

Underground gold mines in Nevada are typically hosted in geologic zones with a wide range of geotechnical properties ranging from blocky competent rock to faulted and highly altered soil-like material. Geotechnical engineering design in weak rock utilizes rock mass classification systems as inputs to both empirical and numerical methods. The most common rock mass classification system utilized in the gold mines of Nevada is the easy-to-use Rock Mass Rating (RMR) system; however, in practice it is difficult to apply in weak to very weak rock masses. This paper attempts to improve the sensitivity of RMR in weak rock masses by utilizing W-RMR?a modified procedure for calculating RMR?and evaluates its performance against other methods of determining RMR for predicting in situ moduli. In-situ rock mass moduli were determined using a portable plate loading device (PPLD), and these data were used to quantify the improved sensitivity of the modified RMR versus existing classification systems. The PPLD was designed for performing rapid and economical in-situ deformability tests in a mining environment. Tests were performed at eight sites yielding thirteen rock mass moduli measurements. Statistical analyses indicate Detailed-Discontinuity RMR-89 and W-RMR provide the best correlations to the in-situ moduli compared to other methods of calculating RMR.

Jan 1, 2014

By William R. Kohl, Craig B. Clemmens

"Since 2010, North Central Resources, Bridgeport, WV, has been using the sonic log to estimate the rockmechanics properties of the coal and adjacent strata in their exploration drilling program. In 68 out of the 85 boreholes they have drilled since 2010, they have conducted shear wave analysis on sonic logs run in these boreholes. By incorporating the rock mechanics logs into the geological model of the coal seam, the engineers can develop a better understanding of the support requirements for the mine.WellCAD Borehole processing software provides vastly improved efficiency in analyzing shear waves from borehole sonic data than previously available methods. The calculation of most rock mechanics parameters requires knowledge of the density and both the compressional wave velocity and the shear wave velocity. Density logs are well known and an industry standard. Sonic logging is equally well known, although less frequently used, as an accurate source of compressional wave velocity. Shear wave velocity is difficult to measure in a geophysical log, but can be estimated from digitally recorded wave form records. This is a time consuming manual procedure which requires the visual examination of thousands of individual wave forms.WellCAD allows the viewing all the wave forms on a VDL display. The observer can create a pair of false logs by manually tracing the interference points on the display for each receiver. He can check himself by observing wave forms in the traditional format. He can then create a shear wave transit time log or “Delta-S” log by subtracting the false log from the near receiver from that of the far receiver. This can be plugged into standard formulae to generate curves for most standard rock mechanic parameters, including bulk modulus, shear modulus, Young’s modulus, and Poisson’s ratio.While these parameters do not entirely substitute for destructive testing, they are extremely cost effective. An entire borehole can be analyzed for the cost of a single destructive test. This allows for a more judicious use of destructive testing while permitting the extension of test data into the rock mass with much greater confidence."

Jan 1, 2015

By D. Newman

An underground limestone quarry, developed in the 1940's, transitioned from a surface quarry to an underground room-and-pillar operation. The quarry expanded to a second underground level when the reserves on the first level reached the property boundaries. The historical basis for the sill pillar thickness between the first and second level, pillar dimensions, and mining height was lost during the course of several changes in ownership. When reserves on the second level reached a five-year life, and plans were initiated for a third level, quarry management embarked upon a thorough evaluation of i. current ground control practices for the support of the immediate roof, ii. roof support in the declines between the surface, Level 1, and Level II, iii. pillar centers and pillar safety factors, iv. pillar alignment between levels, v. mining height within a level, and vi. the locations of declines to access future reserves on level three. The evaluation began with detailed mapping of the location and orientation of calcite veins, joints, fractures, and ground conditions in the immediate roof of Level II. The orientation and location of calcite veins, joints, and fractures on Level II were compared with similar mapping on Level 1, completed in the 1970's to establish whether these features were continuous and would be expected to affect conditions on Level III. Rock strength, physical properties, and fracture spacing were obtained from two (2) diamond core holes drilled from Level I to Level II and two (2) holes drilled from Level II to define the vertical extent of future reserves on Level III. The effort to characterize geological conditions was expanded to a rock mechanics investigation of the: i. performance of the current roof support, ii. optimization of roof and rib scaling, iii. calculation of safety factors for 416 pillars on Level II, iv. evaluation of future pillar centers and mining height within a level based upon pillar strength and pillar geometry, v. development of beam safety factors for the immediate roof on Level II, vi. identification of the most suitable strata for the roof in the future level three, and vii. use of LAMODEL to develop a large numerical model of overburden, inter-level, and total stress on the pillars in levels one and Level II with a stress "footprint" on the proposed level three. The characterization of engineering parameters and geological conditions resulted in changes to the: ? active operations in Level II regarding scaling ? equipment and scaling practices, ? pillar centers, ? mining height, ? raising the immediate roof horizon to avoid ? laminated strata in the current roof, and ? a change in rock bolts from a friction based system ? to active tensioned rebar. The planning and design of level three is ongoing and relies heavily upon the geological and engineering investigations conducted during the past year and one-half.

Jan 1, 2001

By J. M. Girard

Ground control problems at surface mining operations can occur for a variety of reasons. Stress, gravity loading, rock strength, geology, pore pressure, weather effects, underground workings, and many other factors contribute to slope instabilities that range from small rock falls to massive slides of material. While some of these failures can he predicted or controlled by preventive measures, each year many completely unexpected failures occur. Current methods for monitoring generally involve measuring displacements at a few, selected points in and around the suspected area of instability. While most of the displacements along these points will be in a downslope direction, freeze-thaw cycles of water-filled joints, horizontal stresses or pressure, buoyancy in saturated soils, human measurement errors, or other situations can produce deformation in almost any direction, even without any instability in the slope. Determination of which, if any of the observed movements represents a potential hazard is essential.

Jan 1, 1998

By Axel Preusse

To detect improper deformations of a roof bolted roadway in order to reduce inner mine damages in the most early stage and to initiate suitable countermeasures if necessary, several different measurement systems were used for monitoring a roadway's stability on the hardcoal mine "Auguste Victoria" (DSK AG). In this paper the different convergence measurement procedures are described and their respective results are introduced. Subsequently the results of the single measurement methods and procedures are correlated with each other.

Jan 1, 2000

By Gift Makusha

Pre-feasibility work at Goedehoop Colliery indicated that stooping of the undersized pillars could be undertaken safely and economically. In light of the findings a decision was taken in may 2003 to undertake the project in the 600 area where three panels (628, 630 and 632) were identified for pilot test sites. The findings of the pre-feasibility studies were presented in a paper to the 22nd International Conference on Ground Control in Mining held at Morgantown in 2003 (1). This paper is a progress report for the project 12 months down the line. Work commenced to prepare the sites for mining including, acquisition of equipment, equipping the panels, selecting and training people to undertake the project. Three monitoring sites were set up distributed evenly along panel 628. Each site comprised of a surface extensometer station in the overburden strata adjacent to a pillar underground containing hard inclusion vibrating wire extensometer cells. This combination would be used to record pillar loading ahead of stooping as well as the depth of caving and strata deformation in the goal zone behind the stooping face. In July 2003, support work began to secure the roof and rib sides. The support was required to be at least three-pillar center distances ahead of the mining front to ensure that the effective stooping zone is secured at all times. The NEVID method of pillar extraction developed at Sasol and described in detail by Ismet and Moolman (2) in a Coaltech 20-20 report was adopted for the pilot project. As part of training the selected stooping team was send for a week to Sasol's Bosmanskrans Colliery to familiarize with the NEVID method of stooping. Further training sessions were carried out on site especially on the use of Mobile Roof supports. The first cut was taken on 20 August 2003 by implementing a form of chequerbord mining taking every second row of pillars. Three rows were mined this way which period was considered training to get workers familiar with the new equipment and mining method. On the fourth row the full design span equivalent to mining 4 pillars in a row was implemented taking out all the pillars in successive rows. Production figures gradually increased from 300 to more than 1,000 tons run of mine coal per shift. Panel 628 was mined out at the end of December 2003 and the second of three panels (630) is being mined. Substantial data from the monitoring set up has been recorded since the start of the project. Monitoring will continue to the end of the pilot project and the data will be analysed and fed back into the design tools to improve the design in future undertakings. The pilot project has been a success so far and other old workings are being evaluated for stooping.

Jan 1, 2004

By C. J. H. Brest van Kempen

The techniques developed should provide a useful tool, not only during the initial formulation of a suitable bolting plan for a new section, but also for periodic check¬ing on the utilization factor of hardware installed in the roof. Indications are that with present practice, utilization seldom exceeds 507.. Retro-fittable bolter equipment now available (FCBS?) allows doubling of the utilization factor in many cases. We are also introducing a computer pro¬gram based on the material outlined in this paper. This program can save a great deal of time in examining the real effect of implementing alternate bolting plans or using different installation techniques. It allows fine-tuning of bolt lengths, spacing and installation tension to mini¬mize roof support cost, while maximizing safe life of the mine opening. We have developed a procedure which allows easy, site-specific quantification of the need for reinforcement of a mine roof. The concept is based on the forma¬tion of a direct link between theory and empirical data to custom-fit each mine section. In its simplest form, the proce¬dure requires only recording of roof bolt torques in a sample working place (16 to 25 bolts) several times during the first week or so after initial bolting. The way in which the torque pattern changes during this time can reveal whether or not the bolts carry enough tension to maintain stability in that specific roof. From the same data, the nature of the anchorage horizon is derived. Specifically, shale or limestone can be distinguished from sandstone. If the procedure is repeated a few times (within the same mine section) with different bolt installation tensions, we can numerically characterize the roof in terms of one single value: an "effective angle of internal friction". This number can then serve to calculate trade-offs in bolting plan design, such as bolt length, bolt spacing and optimum bolt tension. Additional refinement is possible if a sagmeter is installed in the sample work¬ing place being monitored and if bolt anchorage capacity is determined. The lat¬ter can be done automatically on each bolt as it is being installed, using a system we developed and patented. Alternately, a number of pull tests could be performed. Using the sagmeter, bolt torque and anchor¬age data, it is possible to project the safe life of the supported opening. The details of the procedures described apply specifically to tensioned roof bolting, but a brief discussion paints the way for a similar analysis of full-column resin bolting also.

Jan 1, 1986

By Eddie Martin

Jim Walter Resources, Mining Division, employs the retreat longwall mining system as the primary method of coal production. The strata control aspects of developing and maintaining roadways for retreat longwall mining have continued to improve and advance at Jim Walter Resources. New and innovative concepts in pillar design, roof and rib support, as well as longwall face support, have been combined to provide better strata control. Control of immediate roof breakup in the headgate and tailgate entries has been a problem at Jim Walter Resources' mines. The application of roof trusses as supplemental roof support failed to provide the desired strata stability. The immediate roof deterioration has been controlled through the installation of increased density of roof bolts in the affected roadways. Control of the rib sides has been a documented problem at Jim Walter Resources' No. 4 Mine since its start-up. The installation of 3/4 inch diameter roof bolts at 2 ft. spacing failed to control rib rolls. The development and application of a yieldable rib support system provided the desired results. Rib rolls have been virtually eliminated where the yieldable system of rib support has been installed. An evaluation of longwall panel development using the total yield pillar mining plan has been completed. Results of the strata stability and other operational data have been compiled. Coal production and gate entry stability were not adversely affected using the total yield pillar mining plan for longwall panel development. Strata Control on the longwall face is provided by the face support system. Excessive convergence and roof breakup along the longwall face have been experienced at several of the Jim Walter mines. In an effort to eliminate this problem, an evaluation of higher setting pressures for longwall face supports is being conducted. Results have been inconclusive to date. Each of the above strata control projects will be presented as case studies. Since Jim Walter Resources business is coal production, emphasis will be directed toward the application rather than the theoretical aspect of the projects.

Jan 1, 1988

By Syd Peng

It is necessary to fully understand roof geological features in order to design a proper roof support system for underground coal mines. These geological features include: rock type, rock strength. rock layer interfaces, voids, cracks, and bed separations. Traditionally, engineers have relied heavily on the use of surface borehole log data to identify geological features. Since surface boreholes very expensive to be drilled in close spacing, log data are often insufficient to provide detailed information on the roof geology. Recent advances in control technology have made a large amount of roof bolter drilling data available to mine personnel. These data can be obtained during normal roof bolt installation cycle with little or no interruption in mining operations. This research project has developed several methods to identify geological features in a mine roof from measured drilling parameters. Power trend line analysis can be used to locate discontinuities and interfaces in mine roof. Power averaging can be applied to determine the relative strength of roof strata. Thrust trend line analysis can be used to locate fractures in mine roof. The results of these methods are combined to produce a mapped roof geology In a production environment, manually interpreting and managing collected drilling parameters is not practical since hundreds of holes will be drilled during a production day. A new software package, MRGIS, has been developed to allow mine engineers to make use of the large amount of roof drilling parameters for roof support design. An actual underground test site is presented to demonstrate the analysis and display capabilities of MRGIS.

Jan 1, 2003

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 John C. Stankus

At the 19th International Conference on Ground Control in Mining, a paper was presented detailing new types of Jennmar INSTáL tensionable cable bolts (Stankus and Taylor, 2000). Since that time, significant testing and trials of the various types of tensionable cable bolts have been conducted. Several MSHA districts and state regulatory agencies have now approved these systems as primary support This paper will detail the use of tensionable cable bolts as primary support at two longwall mines in Virginia (MSHA District 5).

Jan 1, 2001

By Thomas Barczak

A new generation of hydraulic mine support prestressing devices has been developed. These thin-walled steel shells are machine-welded and can be inflated with water or any liquid to provide prestressing for a wide variety of roof support products ranging from Can supports to props and cable bolts. With an expansion capability of several inches, depending on the geometry and size of the unit, they can also establish roof contact eliminating the need for wooden wedges or crib blocks that are commonly employed with several standing roof support products. Capable of withstanding pressures from 1,000 to 6,000 psi, these cells are able to transfer roof loading into the support structure without rupturing. A recent development has been the addition of an inexpensive yield valve that provides control of the maximum pressure and load development on the support. This paper examines the performance capabilities of these inflatable prestressing units and the impact they have on the performance of various support systems, including an evaluation of the overall stiffness of the support system and the load control during yielding of the prestressing unit. Recommendations are made regarding the design requirements of the prestressing unit to optimize the support performance. A summary of mine applications using this technology is provided with general comments regarding their impact on ground control.

Jan 1, 2004

By S. M. Matthews

The magnitude and orientation of the in situ stressfield has been determined as a key factor in controlling the stability of openings for both coal mine development and extraction. Monitoring of roadway and pillar behaviour in conjunction with numerical analysis has identified specific stress control measures that have been successfully applied to improve productivity in underground coal mining operations. The influence of the major horizontal stressfield on the stability of openings has been identified at a wide range of sites including mines in Australia, Britain, the USA, New Zealand and Japan. Stress mapping, stress measurement and stress monitoring techniques have been applied to define the in situ stress regime and, together with the correlation of measurements of rock deformation and rock reinforcement behaviour, has enabled appropriate stress control measures to be developed. Stress control measures for development headings have required the selection of roadway orientation, roadway sequencing, reinforcement density and pillar size. The monitoring of behaviour at a range of trial sites has confirmed significant improvements in development stability. Stress control measures for longwall extraction have required rational selection of longwall orientation and extraction direction. Where a mine layout design incorporating the optimum longwall layout design has not been possible, the use of sacrificial roadways has been successfully applied to protect key roadways from stress concentration effects. Comprehensive in situ monitoring of behaviour has been carried out to confirm the mechanisms involved.

Jan 1, 1992

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 Marc T. Filigenzi

Two of the major ground control safety issues confronting underground mine operations today are shaft pillar stability and the failure of rock around active mine openings. Failure of a mine shaft can lead to the entrapment of workers. Failure of rock around active underground mine openings can lead to roof falls, which in turn can result in worker injuries and fatalities. Finite-element analysis has proven to he a reliable method for predicting stresses and displacements around underground mine openings. However, this is a complex and time consuming technique and is not used as often as it could be in mine planning. The purpose of this paper is to demonstrate one technique developed at the Spokane Research Center that allows the user to create a finite-element model of a two-dimensional section of an underground mine in a relatively straightforward manner. This model is then used to calculate stresses, displacements, and safety factors around mine openings. With this information, mine planners can evaluate the stability of mine openings as well as the stability of pillars and mine shafts. This analysis will help develop mining sequences and layouts that minimize stresses in that section of the mine. This, in turn, should minimize the occurrence of shaft failure, roof falls, and other hazards associated with underground mining.

Jan 1, 1997

By Jeffrey K. Whyatt

Estimation of the initial vertical stress carried in a coal seam is an important first step in virtually all methods of evaluating the required size of pillars in coal mines. Such estimates are a trivial exercise in coal fields overlain by gently undulating topography. The LaModel program is easily and routinely applied in such cases to efficiently estimate mining-induced stresses throughout the coal seam. However, application of LaModel to coal mines operating under rugged topography and/or in tectonically disturbed regions is more difficult. LaModel estimates stresses in a twostep process. Topographic and multi-seam mining effects on insitu stresses are estimated in an initial step, and then mininginduced stresses are estimated in an independent second step. A LaModel program function has recently been introduced that decouples these two steps, allowing in-situ stresses to be directly entered from a file rather than calculated. One application of this function is to streamline sensitivity studies by avoiding repeated recalculation of in-situ stress. The function also allows importation of in-situ stress estimates developed by other means. There are no constraints on the imported stress field, which may include significant stress perturbations due to topography, faults, etc. Hence, the range of geologic environments that can be examined by LaModel is significantly increased. This study uses the new decoupling function in an exploration of in-situ stress estimation in mountainous terrain. One option is to use the topographic stress estimation feature included within the LaModel program, and the influence of important parameters in this routine is explored. Another option explored is importation of results from a detailed volume element model. Results of this study suggest that the method used to estimate in-situ stress, and the details of its execution, can have a significant impact for mines in mountainous terrain with strong overburden. More generally, decoupling is an efficient and flexible function that should be considered for most design studies.

Jan 1, 2011

By Douglas Morrison

Hardrock mines in Ontario generally operate at between 3,000 and 7,500 ft below surface and generally experience significant rockburst activity below 5,000 ft. Since the introduction of bulk mining techniques around 1980, the safety record of Canadian hardrock mines has improved dramatically. However, the process of improvement was not a smooth one and, in the early years, the number of accidents and fatalities actually increased. One incident in 1984 that resulted in 4 fatalities in Ontario initiated a complete re-examination of mining practice and by the mid 1980s many changes had been introduced, driven by new non-prescriptive mining legislation. These changes affected not only the technical discipline of mining engineering but also the management of operational procedures and, together, they resulted in the significant improvements that have continued to this day. The analysis of the rock-fall related fatalities showed 5 major causes (loose, scaling, installing support, collapse and rockburst) and by the mid 1990s fatalities in all of these categories, except rockbursting, had been eliminated and even then, all these rockburst fatalities occurred in one mine which was closed shortly afterward. This paper discusses both the technical and social reasons for the initial failure and the ultimate success in improving the safety of these mines, with implications for mines in other major mining jurisdictions.

Jan 1, 2003