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By William Kendall

INTRODUCTION Stillwater Mining Company, with its two underground, hard rock mines in South Central Montana, is the largest primary producer of palladium and platinum in the Western Hemisphere. Platinum Group Metals (PGMs) are mined from a specific zone within the Stillwater Complex, a layered mafic/ultramafic igneous intrusion, named as the J-M Reef. Over much of the J-M Reef, the economically minable, mineralized horizon is contained at typical horizontal widths of 6.5 ft to 8 ft wide. This feature of narrow mining widths has made it challenging to safely and effectively mine these areas of the J-M Reef using mechanized cut and fill and sublevel sill development mining methods. Mechanization of most or all of the steps of the mining process is important to Stillwater in the continuous pursuit of high standards in safety and health. Mechanized bolting within the J-M Reef ore production headings is an important piece of reaching these objectives. In general, metal/non-metal mining requires varied mining techniques, as compared to more traditional bedded mineral/resource deposits. Narrow vein mining is a selective mining technique that generates a minimal amount of waste material while obtaining the more valuable ore deposits. Drilling and bolting in narrow vein mining has traditionally been performed with air-powered, hand-held ?jackleg? equipment because mechanized equipment is, typically, too large for the selective narrow vein techniques. In addition, traditional hand-held pneumatic equipment provides significant flexibility for small-opening, narrow-vein developments, reducing waste rock handling and improving ore recovery management. The hand-held equipment is also relatively inexpensive to purchase. Due to its simplicity, it is also relatively easy to train operators to use hand-held equipment. Hand-held equipment has been routinely used to safely install ground support and to develop production faces in hard rock mines for many years. Although it is effective, the hand-held equipment requires operator exposure to certain inherent hazardous conditions in the active face/development mine areas. These include falling rock, tripping hazards, bending/twisting/lifting hazards, high noise levels, and chemical fumes from the hammer exhaust. Based on health and safety statistics from the Stillwater mines, operation of the rotary-percussive hand-held equipment adversely affects the long-term health of the operator?continual jackleg operation frequently results in long term damage to hearing, shoulders, arm and wrist joints, and back injuries. This reduces the quality of life for the operator and also results in higher personnel turnover rates for the mining company.

Jan 1, 2014

By Ry Stone

"There have been many design practices utilised within the coal mining industry to arrive at the minimum densities of primary ground support required during roadway development. This paper demonstrates the practical use of empirical databases, and focuses on the main drivers for ground support as demonstrated in conceptual models.Golder Associates’ empirical databases used for ground support include a Primary Roof Support Database and a Primary Rib Support Database. Both are based on successful ground support designs installed in mines in Australia, the US, the UK, South Africa, New Zealand, and Europe. The term “successful” refers to those designs that were used on a repeated basis for the purpose of roadway development.The Primary Roof Support Database indicates that the major factors influencing successful roof support designs are roof competency, expressed as the Coal Mine Roof Rating (CMRR), and in situ stress. As roof displacement is essentially a function of horizontal stress and roof competency, in order to provide some form of measure as to the likely magnitude of roof deformation during roadway development, the depth of cover is divided by the CMRR to arrive at a “Stress Strength Ratio” (SSR). The database indicates that the higher the SSR, the greater the likelihood that the roof will require higher densities of support.In regard to the Primary Rib Support Database, it is evident from the current database that the primary factors affecting the capacity of rib support required for a successful design are roadway height and depth of cover. In order to provide some form of measure as to the likely magnitude of rib deformation during roadway development, the depth of cover is multiplied by the height of the roadway to determine a “Rib Deformation Rating” (RDR). Similar to the roof, the database indicates that the higher the RDR, the greater the likelihood that the rib will require higher densities of support.These databases have been used to help determine the minimum primary ground support designs required at many mine sites in Australasia, Europe, and the US. This paper will demonstrate the effectiveness and practicality of these databases at two selected mines in Australia and the US.In order to improve the Primary Rib Support Database, this paper will also propose a new rib deformation rating based on the addition of site specific coal strength data for the Australian mines. The proposed rating attempts to capture the main variables that define the behaviour of a buckling column."

Jan 1, 2015

By Li-min Liu

The Semi-Analytical method can be used to solve a variety of problems in geomechanics and civil engineering. When it comes to layered rock strata and rock masses, it is superior to all other analytical methods. This paper presents the application of a 3-D semi- analytical method to 3-D surface subsidence and stress analysis caused by mining. In this semi-analytical method the layered elements and triangular elements ate used alternatively. Their displacement pattern, strain mahix, elastic matrix, stiffiness matrix, load matrix and stress matrix are presented. A numerical simulation program based on the semi-analytical method bas been developed using Fortran 90. This program is suitable for subsidence prediction and stress analysis of mom and pillar mining, strip longwall and mechanized longwall mining methods. It is applicable to all geological conditions and mining methods. Comparing with the finite method, this method reduces significantly the number of simultaneous equations to be solved and the amount of data preparation. It not only reduces the computer memory, but also runs faster and cost less. Key Words Semi-analytical method, surface subsidence, stress analysis, layered element, triangular prism element

Jan 1, 2002

By Ross Seedsman

There are many models for how roof bolts behave, one of which is prevention of shear along bedding. The basis of this model is that a roof beam will not delaminate if bedding parallel shear can be prevented. A key point in such a model is that bedding discontinuities are present naturally in the roof and that mining induced stresses do not generate new surfaces. The presence of these discontinuities means that the rock mass must be modelled as a transversely isotropic material if continuum codes are used. A limit-equilibrium type design method for specifying roof bolts is being developed using voussoir beam and excess shear force concepts. The method incorporates much of what operators know intrinsically about roof reinforcement, and allows mining engineers to focus on the key uncertainty - the frequency and location of bedding surfaces in the immediate roof. Areas where more research is required are highlighted.

Jan 1, 2000

By Kevin J. Ma, John Stankus, Dakota Faulkner

"Expandable rock bolts are widely used in hard rock mines as an efficient ground control product. However, capacity and service life can be significantly reduced if the metallic body is subjected to corrosion. In some hard rock mines in the U.S., highly corrosive ground conditions exist, and it has been reported that inflatable rock bolts have corroded within a few months after installation. To provide mining industry a cost-effective inflatable bolt and combat the corrosion issues, Jennmar Corporation, Inc., and its subsidiary Keystone Mining Services, LLC (KMS), analyzed corroded bolt samples, identified root causes, evaluated properties of various coating materials, and developed a new inflatable rock bolt, Python M3TM, that is protected with an innovative PyFlexU2TM coating. The new generation Python M3 features improved steel chemistry for reliable performance, modified profile for better inflation, and surface preparation and coating application. The PyFlexU2 is impervious to liquid and air, durable, and UV resistant. With a flexible adhesive, and highly corrosion-resistant undercoating, and a very hard sacrificial surface coating, the PyFlexU2 coating system provides the Python M3 superior protection against chemical corrosion and physical scratch damage. The under-coating has exceptional flexibility and adhesion to prevent coating micro-cracks or fractures after bolt inflation and possesses excellent corrosion resistance to acids (pH <3), alkalis (pH >11), fuels, and salt solvents. The corrosion and scratch resistant PyFlexU2 coating offers very effective bolt protection for extra longevity in highly corrosive environments. The Python M3 coated with PyFlexU2 has been tested in the most challenging conditions, including laboratory corrosion tests in extreme acidic and basic solvents, rock slurry, and borehole scratch insertion tests. With demonstrated corrosion and scratch resistance, the product has been greatly welcomed by hard rock mines in the West and is currently installed in large scale. This paper identifies the root causes of the bolt corrosion, discusses the analysis process, and details laboratory and underground tests carried out on the Python M3 coated with PyFlexU2. The Python M3TM and PyFlexU2TM are subjects covered by pending U.S. Patent Applications assigned to FCI Holdings Delaware, LLC."

Jan 1, 2017

By John Stankus

Accurate evaluation of stress and geological conditions is critical to ground control design in underground openings. For a mine slope entry, the problem becomes more complicated because a slope transverses through many different types of strata. Mine slope ground control has traditionally been practiced based on trial and error and prior experience. There is currently no well-accepted slope and support design methodology available for an economical and technically sound ground control plan. Over the last few years, Keystone Mining Services, LLC (KMS), the engineering affiliate company of Jennmar Cooperation, has developed a ground support design methodology called the ?Stress, Geologic, and Support design system? (SGSSM). This patent pending method enables accurate analysis of actual geo-technical conditions and stress; identifies strong, fair, and weak sections along the slope; formulates primary and supplemental roof bolting systems, designs optimal long-term steel structure support; and verifies the adequacy of the developed support. The method has been successfully utilized on various new slope projects and has been well received by state and federal regulatory agencies during the approval process. This paper details the concepts of the methodology using an actual case study of a slope project in the Midwest.

Jan 1, 2011

By Dale S. Preece

In 1975 Congress passed the Energy Conservation Act to establish a U. S. Strategic Petroleum Reserve (SPR) with a capacity of 750 million barrels of crude oil. The most economic storage medium was determined to be salt caverns leached in salt domes in Louisiana and Texas. Salt caverns existed at several sites when the reserve was created. These were obtained by the U. S. Department of Energy (DOE) and used to initiate SPR oil storage. In order to meet the storage capacity approved by Congress, new caverns also had to be leached. To support the resulting design effort, finite element computer program have been used to determine the creep closure and structural stability of salt caverns. Using site specific material properties including creep models, elastic moduli and fracture data, the finite element analyses have replaced earlier empirical approaches to cavern design. This report presents results of such finite element analyses to determine the best cavern roof shape and the minium pillar to dimeter ratio. , P/D. These numerical predictions indicate that the current cavern design is safe.

Jan 1, 1984

By Michael Jenkins

This paper reviews the history of rock burst research in the Coeur d&apos;Alene Mining District of northern Idaho over the last 30 years. Rock bursting is a problem because of a combination of geological factors in the region--complex structure, high horizontal stresses, and strong, brittle rock that readily stores strain energy. The Bureau of Mines has cooperated with industry to study rock burst phenomenon for many years. Early work focused on methods to mitigate bursting problems within the context of existing mining methods. In the 1970&apos;s, researchers began investigating methods for altering or controlling conditions that lead to rock bursts. These control measures become part of the stope development that precedes mining or are changes in the mining process itself. Although, researchers continue to develop microseismic techniques, the major research effort focuses on developing a new mining method to reduce the number and severity of rock bursts in order to improve safety and increase productivity in the Coeur d&apos;Alene Mining District.

Jan 1, 1987

By Richard N. Campbell

The world?s ever-expanding demand for energy and coking coals has resulted in the underground mining industry looking at taking on increasingly marginal and geotechnically challenging deposits. The drivers for this change include a reduced availability of open cut reserves and a lack of remaining ?good coal leases?. The lack of plum deposits is forcing some companies wanting to enter the coal supply market to purchase and operate in higher risk, lower margin areas. As geotechnical professionals, we are expected to develop strategies to facilitate safe and productive mining in ground that once would have been avoided. To be successful in these reserves, we are required to use every tool available, from exploration data to longwall automation, and invent or reinvent some new tools along the way. This paper presents geotechnical case studies of challenges in longwall mining and some of the methods developed and used to allow longwall mining to succeed where others said we would fail.

Jan 1, 2014

By William Monaghan

Over the period 2000 to 2003, roof falls have accounted for 4 to 14% of the fatalities in underground mining operations. The National Institute for Occupational Safety and Health (NIOSH) is conducting research to reduce the frequency, exposure, and risk of these events through an ongoing program of field and laboratory studies. One area of research involves the evaluation of polyurethane grouting technology that is commonly used to stabilize fractured mine roof strata. Concurrently, NIOSH is conducting work to evaluate the application of ground penetrating radar for advanced delineation of problematic mining areas. In this study, NIOSH partnered with Sub-Technical, Inc. who injected a polyurethane grout (also known as glue) into a roof area of NIOSH&apos;s Safety Research Coal Mine. The mine roof area was scanned using ground penetrating radar technology before and after grout injection in an attempt to determine the extent of grout infiltration. A comparison of the pre-and post-grouting radar records showed a significant change at a mine roof depth of about four to five feet. The interpreted radar records were then compared with drill core information, borescope evaluation of roof-bolt holes in the study area, and underground observations. At this site, the interpretations of the radar records correlated with data obtained from the core holes, borescope evaluations and underground observations. This study showed that ground penetrating radar technology can be a useful tool for detecting changes in mine roof due to the injection of the grout.

Jan 1, 2004

By Khaled Morsy

In this paper a new method to evaluate the stability of yield pillars was introduced. The evaluation of yield pillar stability was conducted in three steps: (1) Estimation of pillar loading - numerical modeling is a good tool to estimate the pillar loading for most of the geological, geometrical and mining conditions. In this study. finite element technique was used to estimate the state of stress, strain, elastic strain energy and plastic energy, etc. at every point within the pillar. (2) Defining the yield pillar loading zones¬. The proposed method considered the non-uniform stress distribution of yield pillars by dividing the area of the yield pillar into three loading zones, namely core, transition and rib zones. (3) Estimation of stability indicators - an appropriate stability measure was proposed for each loading zone. such as; Core Stability Factor. CSF, Pillar Bump Index, PBI and Rib Instability Index. RIF. A case study of successful and unsuccessful yield pillar application was used to assess the applicability of the proposed method.

Jan 1, 2003

By Behdeen Oraee-Mirzamani

Since the advent of New Austrian Tunneling Method (NATM), shotcrete as a primary means of support in tunnels has been widely applied. It?s most important features are durability, speed of application and cost effectiveness. This paper introduces some tables that provide guidelines for the thickness of shotcrete required in some common situations of mine roadways. In order to devise such tables, two different arch sections, together with three different overburden types, were considered. Geotechnical parameters such as apparent cohesion and angle of internal friction of surrounding rocks were chosen, based on the five-category classification of Bieniawski. Two K0 factors (the ratio of horizontal stress to vertical stress) and an average rock density were utilized. Using numerical methods, 60 models were then devised in this way. By applying interaction diagrams of axial force and the bending moment for different thicknesses of shotcrete, appropriate shotcrete thickness for these models were calculated. The results of this research, as well as the methodology applied, can be used in mining roadway support design and all types of civil engineering tunnels.

Jan 1, 2011

By Mark Colwell

This paper summarizes the results of a research project whose goal was to provide the Australian coal industry with a chain pillar design methodology readily usable by colliery staff. The project was primarily funded by the Australian Coal Association Research Program and further supported by several Australian longwall operations. The starting point or basis of the project was the Analysis of Longwall Pillar Stability (ALPS) methodology. ALPS was chosen because of its operational focus; it uses tailgate performance as the determining chain pillar design criterion rather than simply pillar stability. Furthermore, ALPS recognizes that several geotechnical and design factors, including (but not limited to) chain pillar stability, affect that performance. There are some geotechnical and mine layout differences between United States and Australian coalfields that required investigation and, therefore, calibration before the full benefits offered by the ALPS methodology could be realized in Australia. Ultimately, case history data were collected from 19 longwall mines representing approximately 60% of all Australian longwall operations. In addition, six monitoring sites incorporated an array of hydraulic stress cells to measure the change in vertical stress throughout the various phases of the longwall extraction cycle. The sites also incorporated extensometers to monitor roof and rib performance in response to the retreating longwall face. The study found strong relationships between the tailgate stability factor, the Coal Mine Roof Rating, and the installed level of primary support. The final outcome of the project is a chain pillar design methodology called Analysis of Longwall Tailgate Serviceability (ALTS). Guidelines for using ALTS are provided.

Jan 1, 1999

By Gabriel S. Esterhuizen

"Cable bolts are sometimes used in low-seam coal mines to provide support in difficult ground conditions. This paper describes cable bolting solutions at two low-seam coal mines in similar ground conditions. Both mines used support systems incorporating cable bolts as part of the primary support system. Two original cable bolt based support systems as well as two modified systems are evaluated to estimate their ability to prevent large roof falls. One of the support systems incorporated passive cable bolts, while the other used pre-tensioned cable bolts. The results and experience at the mines showed that the modified systems provided improved stability over the original support systems. The presence of the cable bolts is the most important contribution to stability against large roof falls, rather than the details of the support pattern. It was also found that a heavy steel channel can improve the safety of the system because of the ‘sling’ action it provides. Additionally, the analysis showed that fully-grouted rebar bolts load much earlier than the cable bolts, and pre-tensioning of the cable bolts can result in a more uniform distribution of loading in the roof.INTRODUCTIONCable bolting is sometimes used as primary support in coal mines experiencing difficult roof conditions. In low-seam mines the flexibility of the cable bolts allows greater length supports to be installed near the advancing face without the use of couplers. When used as primary support, the cables are typically installed in the same row as fully grouted bolts, replacing two or more of the bolts in each support row. A heavy steel channel may be used as a strap to spread the support load over a greater portion of the roof. Historically, the Mine Safety and Health Administration (MSHA) has not allowed widespread use of partially grouted un-tensioned bolts (e.g., passive cable bolts) for primary support; however, pretensioned cable bolts have been accepted.Various solutions using cable bolts as primary support were attempted at two low-seam coal mines in Western Pennsylvania that were experiencing difficult roof conditions. Both mines originally used fully grouted rebar bolts as primary support and cable bolts as supplementary support. It was found that when a large roof fall occurred, the cable bolts may be contained within the dome of fallen rock. As problematic roof conditions continued to exist, both mines decided to use cable bolts as part of the primary support system. The cable bolts were located near the ribs of the entry, to increase the likelihood that they would be anchored outside the dome of potentially unstable roof. The cable and rebar bolts were installed on a heavy steel channel that acts as a ‘sling’ to distribute the load across the width of the entry. At the first mine, Mine A, pre-tensioned cable bolts were used while at Mine B, un-tensioned cables were used. The two support systems consist of essentially the same support components installed in different patterns and with varying degrees of pre-tension."

Jan 1, 2015

By Kevin MacPherson

After 34km of roof bolted drivage the first fall of ground occurred at the Phalen Colliery in the maingate of the 8th East longwall panel. The fall occurred over the stage loader as the wall approached an area of poor roof geology. This area had developed a severe cutter since initial drivage and the longwall had yet to mine through 500m of even worse conditions. This paper describes the cause of the poor roof conditions, the use of polyurethane resin injection and cable bolting to re-stabilize the roof outby the fall, and the measurement of roof movement and performance subsequent to the continuation of mining under the rehabilitated roof.

Jan 1, 1999

By Wenjun Ju

It&apos;s very important to determine the limitation of bed separation in entry roof in order to predict roofs stability and take some necessary precautions. Based on numerical modeling and monitoring in the field, the new concept of Limitation of Roof Layer Separation (LRLS) and its method of determination LRLS is given in this paper

Jan 1, 2001

By Michael Kelly

Current research in longwall mining in CSIRO Exploration & Mining is focused within five major projects and several minor projects. The major projects are: 1 3D Aspects of Longwall Geomechanics. This was reported to the conference last year and a final report will be available by mid-2001. 2 Mine Gas Control. This just completed project centred on Dartbrook Colliery which has a longwall gas make of more than 8 cubic metres per second. 3 Rapid Roadway Development This includes the construction of an autonomous conveying and bolting module (ACBM) that sits between the miner and haulage system. 4 Top Coal Caving In conjunction with the YanKuang group in China, CSIRO and the University of NSW are conducting a prefeasibility study of the Top Coal Caving method in Australian conditions. 5 Longwall Automation. In partnership with the Cooperative Research Centre for Mining Technology and Equipment (CMTE), the CSIRO has conducted a scoping study on this issue and will commence an initial three-year demonstration project in July this year. This paper describes these protects in more detail with particular focus on the last three in the above list.

Jan 1, 2001

By Houquan Zhang, Yu Wu, Yanlong Chen, Zhenfu Luo, Aihong Lu

"A non-destructive detection scheme for bolt installation quality was designed by adopting the method of non-destructive detection on bolt installation quality. Based on the field measurements, a random non-destructive detection method for determining the characteristic length including anchoring length, free segment length and initial axial load of bolt was discussed, and the characteristics of increasing bolt axial load in the initial stable stage was obtained. The results show that the bolt axial load and deformation of surrounding rocks correlate each other. Compared with the traditional detection technique for bolt installation quality, the non-destructive detection method has obvious advantages such as random, accuracy, and economic. The method provides a reliable basis for bolt support optimization and stability control of surrounding rocks.INTRODUCTIONBolt support in roadway can make the surrounding rocks and bolt form a structure in the early stage of roadway development through the bolt pre-tension, which can maximize the self-bearing capacity of surrounding rocks. Thus it is a form of active support (Kang, 2007). In order to make the bolt play an active supporting role, it needs to ensure the installation quality of bolting. Thus the bolts can supply a high initial anchoring force and control the deformation of surrounding rocks (Shan, 2009; Wei, 2012). Moreover, bolt supporting is a concealed work. Once installation quality problems appear; it will be difficult for effective detection and screening (Xie, 2008).THE PROBLEMS OF BOLT INSTALLATION QUALITY IN COAL MINE AND ITS DETECTIONThe Problems of the In Situ Bolt InstallationDue to the problems in installation environment, machine, and technique, and workers proficiency (Li, 2003), there are many problems exist in bolt installation. If the borehole depth is too long, the anchoring agent will accumulate at the bottom of the bore, which reduces the anchoring length. Furthermore, if the borehole depth is short due to the poor geological condition, the exposed part of the bolt will be too Jong and the pallet cannot cling to the rock face, which prevent the pre-tension from being applied. The pre-tension is very important for bolt installation. If pre-tension is insufficient the rock bolt cannot control the deformation of surrounding rock efficiently and the bolt anchoring force will decrease considerably. The bolt length is another important parameter for anchoring quality, which will affect the control area of bolt supporting. In addition, the cleanness of drill hole has a great influence on the effective anchorage length of bolt. If the drill hole cleaning is not enough, it will directly affect the anchoring agent and the contact area of surrounding rocks, which will decrease the effective anchorage length and initial anchor-force."

Jan 1, 2016

By Ken Barish

Jane Mine of the Keystone Coal Mining Corporation was started in 1962 and is in the Lower Freeport seas located in Armstrong County, Pennsylvania, approximately 45 miles northeast of Pittsburgh. The mine is a slope/shaft complex with the cover ranging from 100 feet at the slope to 450 feet near the back end of the property. The seam height ranges from 48 to 72 inches. A variety of roof conditions have been en¬countered, and an equal variety of roof control systems have been tried with varying degrees of success. Initially, places could be mined as deep as 100 feet without any permanent roof support. After several problems and the advent of the Health and Safety Act of 1969, roof control became more and more prevalent and predominant at the Jane Mine. It started with 4 foot conventional oolts. As the mine progressed in the northeasterly direction, roof conditions changed from the extremely good roof that was encountered initially, to bad roof; and other types of roof support were required. Sixty pound rails and 6 inch 25 pound to the foot steel I-beams were used for supplemental support in long life headings such as belt and track. Resin bolting was done in many areas for long life support due to the deterioration of the shale roof over time. Several areas had to be abandoned because of uncontrollable top.

Jan 1, 1984

By G. Skybey

The Directional Drill Monitor (DDM) is an advanced instrument for real time in hole surveying. It was developed by Du Pont (Australia) to suit Australian gassy coal mining conditions. The instrument consists of three units: i. Down hole transmitter ii. Acoustic transducer iii. Data receiver processor unit (i.e. up hole unit). This paper describes the function of each unit in detail and highlights the importance of the DDM in long hole drilling efficiency. Surveying is carried out independent of hole depth. Each survey takes 2-3 minutes to complete, the frequency of surveying to control borehole trajectory is determined by the accuracy required for a particular hole. The DDM is currently being used successfully for long hole drilling in a number of collieries. A high degree of directional control has been achieved. Borehole trajectories were controlled to within +l¬2m of the planned path with intersurvey distances varying between 6 to 18 metres over distances up to 700 metres.

Jan 1, 1992