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By Thimons E. D, Derick R. L

In early June, 1986, a fire of unknown origin erupted in one of several main entries of a coal mine located in southwestern Colorado. The fire was aggressively fought with high-expansion foam. As a result of several strategic decisions, the fire appeared to be held in-check. Unfortunately, sufficient quantities of foam were unavailable and as soon as the foam was depleted, control of the fire was lost. To prevent further damage and the possibility of an explosion, all entries to the mine were sealed.

Jan 1, 1988

By Robert J. Timko, Edward D. Thimons, Robert L. Derick

In early June, 1986, a fire of unknown origin erupted in one of several main entries of a coal mine located in southwestern Colorado. The fire was aggressively fought with high- expansion foam. As a result of several strategic decisions, the fire appeared to be held in check. Unfortunately, sufficient quantities of foam were unavailable and as soon as the foam was depleted, control of the fire was lost. To prevent further damage and the possibility of an explosion, all entries to the mine were sealed. This paper recounts activities from the discovery of the fire to the initial reentry of the sealed mine including location of the fire, the decisions made to best combat it, and the steps taken to rapidly and securely seal the mine when this became necessary. The mine operator requested the Bureau's assistance in obtaining and analyzing mine fire gases taken from sampling ports placed through the mine seals. This report describes how samples were taken and what gases were analyzed. In addition, this work describes the initial exploration of the sealed mine by personnel using oxygen breathing apparatus.

Jan 1, 1987

By J. Bowders

A case study describing the expansion of Marion County landfill is presented. The exapansion area is located on a deep mined area. Finite element analyses were used to predict the surface subsidence and associated tensile strains on the landfill liner of the expansion site. A synthetically reinforced liner system is recommended to minimize the impact of the subsurface movements on the integrity of the liner system.

Jan 1, 1986

By L. C. WYMAN

THIS paper discusses what mining reports should contain when presented to the California State Corporation Department, to accompany applications for the sale of stock to the general public, but the principles involved apply to mine financing everywhere. Such reports should 'be as clear and concise as possible, endeavoring to tell the truth in as few words as necessary. Should the engineer be aware that his report is to be used in connection with the sale of stock to the general public, he should consider the public as his client as well as the person who actually employs him. It is always required that the engineer be favorably known and of recognized standing. The following points should be given careful attention : 1. A legal description and location of the property. 2. Transportation conditions, and how they may be expected to affect the operating costs. 3. A brief history of the property, its production record as to tonnage, grade, and character of the ore should be given. Should the property have suspended operations for any reason, this should be given. Any details of former operating costs should be given, especially as to how any improvement could be made, due to better methods which might be employed.' 4. The main geological features and whether they are such as to lead the engineer to believe that the mine may possibly 'be exhausted or the ore of such grade as to be no longer profitable. If he considers them favorable, he should give his reasons for the belief that further development work may produce profitable results. A map showing these features should be submitted, and it is not necessary that such map be highly colored or contain any more than the essential matters. 5. Other mines in the district, or those with similar

Jan 1, 1929

By G. Iabichino, M. Cravero, F. Nacci, L. Rinaldi, L. Vigliero

ABSTRACT: The proposed paper is aimed at analysing the failure of a steep cliff slope that, in the past, was exploited for open pit mining operations. A volume of over 400,000m3 of gneissic aplitic rock was involved in the slope failure. An evaluation of possible failure mechanisms is given, based on a structural description of the rock mass, a mechanical characterisation and on computation methods. The block theory and limit equilibrium techniques for sliding and block toppling were applied and a comparison is given with the result obtained by using DEM models. RESUME: Le document proposé vise l’analyse des mécanismes de rupture d’une falaise qui a été intéressée dans le passé par une mine à ciel ouvert. Un volume supérieur à 400.000 m3 de aplite-gneiss a été concerné par la rupture de la falaise. Le document fournit une évaluation des possibles mécanismes de rupture, basée sur une description structurale de l’amas rocheu, une caractérisation mécanique et des méthodes de calcul numérique. L’analyse a été développée à l’aide de la théorie des blocs et de la technique de l’équilibre limite, soit pour le glissement que pour le basculement des blocs, et une comparaison avec le résultat obtenu par application des modèles DEM a été effectuée. ZUSAMMENFASSUNG: Die vorgeschlagene Arbeit zielt darauf ab, den Einsturz einer steilen Felswand zu analysieren, an der in der Vergangenheit eine Mine im Übertageabbau betrieben wurde. Der Einsturz betraf ein Volumen von über 400000 m3 an Gneis und Aplit. In der Arbeit ist eine Einschätzung des möglichen Mechanismus angegeben, der zum Bruch geführt hat, und sie basiert auf einer strukturellen Beschreibung der Felsmasse, einer mechanischen Charakterisierung sowie auf der Anwendung von Berechnungsmethoden. Angewandt wurden die Blocktheorie und die Methode der Gleichgewichtsgrenze für das Abrutschen und Kippen von Blöcken. Darüber hinaus wird ein Vergleich mit den Ergebnissen angestellt, die unter Verwendung von DEM-Modellen erzielt wurden.

Jan 1, 2003

By M Nombe

Based on the analysis of a case study at POWELTON No.2 MINE, it was determined that mine stability problems had occurred due to floor softening. A fireclay (underclay) layer in the immediate floor was softening after coal extraction and failing under pillars load. The thickness of soft floor was considered as the governing factor in the determination of the magnitude of floor displacement. The effect of floor thickness on floor displacement was evaluated by creating seven models that were input into the LAMODEL program. The properties of the soft floor were determined from a bearing capacity test performed in the Mains close to the area where mine stability problems had occurred. Moisture increase was considered in the reduction of floor bearing strength. Various strain-softening properties were allocated to the materials representing the soft floor. The LOW GAP POWELTON No.2 MINE was modeled and LAMODEL program was used to determine stress and convergence on both the coal seam and the fireclay seam. The results of critical stress and convergence obtained from modeling were compared with the actual floor failure area in the mine.

Jan 1, 2002

By H. A. Kishawy, S. Kannan, M. Balazinski

Metal Matrix Composites (MMCs) are new generation engineering materials that possess superior physical and mechanical properties compared to nonreinforced alloys. However, the presence of abrasive ceramic reinforcements in the ductile matrix causes severe tool wear and premature tool failure. Flank wear was found to be the dominant wear mode while the main wear mechanism was abrasion. Analysis of the cutting tools using scanning electron microscopy has shown that both two body and three body abrasion are operational during machining MMCs. A methodology for predicting the tool flank wear progression during bar turning ofMMCs was presented in an earlier paper [1]. In the proposed model, the wear volume due to two body and three body abrasion mechanisms was formulated. Then, the flank wear rate was quantified by considering the tool geometry in 3D turning. The main objective of this paper is to investigate the contribution of two body and three body abrasion towards the tool wear volume during cutting MMCs.

Jan 1, 2006

By W. M. Gray

The determination of the state of stress existing at a point in any solid body requires the determination of the six components of the stress tensor relative to a convenient set of axes. One general method of doing this in rock is to make strain or deformation measurements associated with local relief of stress in a number of boreholes penetrating a limited volume in the vicinity of the point at which the stress is to be determined. The stress is assumed to be constant throughout this volume. Leeman1 has described a number of the instruments that have been developed for making deformation or strain measurements in boreholes. Panek2 has described the application of statistical methods to the determination of the average components of ground stress from sets of measurements made in boreholes by means of deformation meters of the type developed by the U.S. Bureau of Mines. In this paper attention is given initially to the strain cell developed for use on the flattened ends of boreholes by the Council for Scientific and Industrial Research (CSIR) , South Africa. The equations required for evaluating the average components of ground stress from measurements made by means of the CSIR strain cell are derived. These equations are similar in form to those applying to measurements made by the borehole deformation meter developed by the U.S. Bureau of Mines. The results of an investigation of the properties of the equations are described. These results are important in the practical use of both the instruments mentioned above and any instrument based on similar principles.

Jan 1, 1968

Agitation leaching data has generally not been analysed to extract the maximum amount of information that is available. Analysis of heap leaching systems has been developing over the last five years to include the underlying rate determining process. This has been found to be diffusion control for all copper leaching; irrespective of whether they are oxide or sulfide minerals. Extension of the diffusion model has enabled better scale-up protocols to be developed. Acid consumption in copper heap leaching operations has been found to be a first-order chemical controlled system. The analysis of the data is relatively straightforward. However the interpretation and scale-up is not as direct as for agitation leaching. The differences in the residence time distributions (RTDs) are required to properly account for the scale-up between test columns and field. The leaching of oxide copper in agitated systems has also been found to be diffusion controlled. However, there are significant differences in the control mechanisms. Initially the leaching control is the supply of acid to the mineral surface. This is limited by the diffusion through the surface boundary layer and is thus subject to the level of agitation. Higher agitation provides thinner boundary layers and faster diffusion rates. Similarly, higher grade material creates a thicker concentration boundary layer and relatively slower leaching. Acid consumption in agitated leach systems is also a first-order system. However, the interpretation of the RTD is easier from batch testing to continuous multiple CSTR series plant operations. Altering the method of acid delivery can significantly reduce acid consumption rates. Predilution of the acid minimises the local acid concentration near the addition point and can reduce the consumption by up to 50 per cent. Modelling of the combined leach rate(s) and acid consumption rates in agitated systems allows comparison of various leaching strategies and selection of the most appropriate. Similarly, the model allows the identification of economic limits to leach recovery; where the cost of acid consumption is greater than the value of the copper recovered. From this analysis selection of residence times for agitation leaching can be determined. The modelling tools are useful in identifying optimum conditions for the leaching process.

Jan 1, 2005

By H. Kamiya, H. Hasegawa, M. Nojiri, S. Matsui, Y. Fukuda, M. Tsukada

"The present paper focuses on the action mechanism of anionic polymer dispersants in dense Si3N4 and SiC suspension. An atomic force microscope (AFM) was used to determine the relationship between the macroscopic suspension viscosity and the microscopic adsorbed structure of a polymer dispersant at the solid / liquid interface. The surface interactions within the suspensions were analyzed in various dispersant pH and additive conditions. The addition of an anionic polymer dispersant decreased Si3N4 and SiC suspension viscosity and increased electro-steric repulsion force on the non-oxide surface in solution at pH> 6 which was the isoelectric point of the materials. It was assumed that the increase of repulsive force was promoted by a hydrophobic adsorption of polymer dispersant on the non-oxide particles.IntroductionThe surface of silicon nitride and silicon carbide powders in a water suspension system has quite a complex molecular structure. In the case of Si3N4, hydroxyl and ammonium groups form on the surface and the pH of the suspension is increased from neutral to about 10 due to surface oxidation. The particle surface is negatively charged in alkaline solutions (pH=10), because the isoelectric point of silicon nitride exists in the pH range from 5 to 6. However, in spite of charging the negative surface charge, an anionic polymeric dispersant is generally used for the reduction in slurry viscosity.In dense SiC suspensions, an anionic polymer dispersant is generally used in order to disperse aggregates and reduce the suspension viscosity, too. Since an anionic dispersant is used in alkaline solutions higher than the isoelectric point of silicon carbide, the dispersant adsorbed on the negative charged surfaces of SiC.Few researchers have discussed and estimated the adsorption mechanism and structure of anionic polymeric dispersants on silicon nitride and silicon carbide surfaces. This is because it is difficult to observe the structure and measure the interaction at the solid/liquid interfuce directly. However, new developments in surface-force measurement technology have enabled the determination of non-DLVO type interaction for analyzing aggregation behavior2. The use of the AFM for analyzing the effects of various surfactants and co-polymers on colloidal dispersions and solid surfaces has also begun to rapidly expand to characterize the interaction between solid surfaces."

Jan 1, 2000

By Frederick B. Kuhnow

Harmonic energetic delamination of T2 building is being considered for demolition at a formerly known atomic lab. The T2 building is within a containment structure that is under a negative pressure of -49 Pa (-0.2 inches of WC) so that no remaining radioactive material is leaked out. The T2 containment enclosure is a portal frame-type system that spans east/west over 35.05 m (114’-11 7/8”) and spaced 2.9 m (9’-6”) apart. The west side of the portal frames is supported by a 64 cm (2’-1”)thick concrete wall which is backfilled to the west and open to the T2 basement to the east. The east side of the portal frames is supported by a 61 cm (2’-0”) thick concrete wall that is backfilled to the east and open to the tank farm vaults to the west. A concrete slab-type footing supports both the east and west walls. Transverse walls provide lateral support for the walls as part of the original design. The west wall has an elevated slab that provides further stiffening of the concrete wall between the transverse walls.

Jan 1, 2017

By Charles W. Axten

Since the mid-1970s, samples of airborne and waterborne fibrous particulates have been collected in the area of the Northshore Taconite Ore Processing Facility by the Minnesota Department of Health (MDH), the Minnesota Pollution Control Agency (PCA), and the University of Minnesota. Indirect sample preparation has consistently been used although other aspects of the sampling methods and sites have varied and analytical procedures were altered over time as more accurate and precise microscopy methods were developed (i.e., phase contrast optical microscopy, transmission electron microscopy, transmission electron microscopy with energy dispersive spectroscopy). In the mid-1970s, levels of airborne fibrous particulate in the Silver Bay area averaged from 0.00030 to 0.03 f/ml. This level was significantly greater than levels of similar particulates in the St. Paul, MN area, although two of the Silver Bay sampling sites, considered individually, did not indicate levels of fibrous particulate markedly different than that seen in St. Paul. More recent sampling data (i.e., 1990?2001) indicate mean concentration of airborne fibrous particulates (amphibole-like fibrous particulates) of 0.0020 f/ml with a range of values from 0.0001 to 0.0140 f/ml. Such levels are not significantly different from those seen in other non-urban environments in the US and Europe. Concentrations of fibrous particulates in water samples were higher in the mid-1970 when iron ore tailings were being deposited in Lake Superior, but since the tailings have been deposited on land waterborne levels of fibrous particulate in the Beaver River have remained relatively constant averaging in the range of 7.5 MFL. This level is only slightly in excess of the current EPA drinking water standard for fibrous particulates. Review and consideration of this data is important in determining the potential health risks associated with airborne and waterborne fibrous particulates in the areas of the Northshore Taconite Ore Processing Facility. © 2007 Elsevier Inc. All rights reserved. Keywords: Taconite; Asbestos; Asbestiform fibers; Fibrous particulates; Cleavage fragments; Phase contrast optical microscopy; Transmission electron microscopy

Jan 1, 2007

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By S. L. Bessinger

In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross-section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Headlosses in mine entries are calculated using Atkinson's Equation (1). [H = (English) H = (SI) (1)] where H = pressure ,loss, in. of H20 (Pa) K = friction factor, lbf min2/ft"4(Kg/m3) P = perimeter, ft. (m) L = airway length, ft (m) Q = airflow quantity, ft3/min (m3/sec) A = flow cross sectional area, ft2 (m2)

Jan 1, 1987

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 Robert L. Grayson, Syd S. Peng

"An airborne dust study was undertaken at a Southwestern Pennsylvania underground bituminous coal mine by the Pennsylvania Department of Mines and Mineral Industries in 1964 for the purpose of making recommendations and requirements regarding compliance with health and safety aspects of The Bituminous Coal Mining Laws of Pennsylvania, approved by the legislature on July 17, 1961. The study was designed to accurately represent the airborne dust conditions of the subject mine. Sampling by midget impinger of specific worker locations throughout the mine was accomplished. Sample analyses yielded dust concentrations to mppcf, particle size determinations in cumulative per cent less than specified cut-off diameters, and determinations of free silica percentages in total dust samples. The data from this study is further analyzed, and the results are presented.INTRODUCTIONThe Bituminous Coal Mining Laws of Pennsylvania, enacted in 1961, prompted a study of respirable coal mine dust by the Department of Mines and Minerals Industry in 1964. In recognition of their obligation for ensuring the health and safety of bituminous coal miners, the department set out to characterize the airborne dust of mines throughout the commonwealth. The purpose of the studies was to make recommendations and requirements for compliance with the law and for control of airborne coal mine dust in Pennsylvania mines.One such study was undertaken at an underground bituminous coal mine operating in the Pittsburgh Seam in Southwestern Pennsylvania. The study was desig11ed to sample specific worker locations throughout the mine site and characterize the respirable dust relative to each of them. A midget impinger was used to gather respirable range dust samples during an 8-hour shift. Continuous miner operator, roof bolter, stoper bolter, loading machine operator, shuttle car operator, gathering motorman, mainline motorman, and rotary dumper jobs were sampled underground. In addition, barge mover, coal sampler, fine coal operator, heavy media operator, and boiler room attendant occupations were sampled on the surface. The surface sampling was over a 71-hour shift because of UMWA contract provisions."

Jan 1, 1988

By David K. Parrish, Duane A. Labreche

Analyses of jointed, cracked, and fractured rock masses are needed in a variety of applications, including estimates of reduced elastic modulus for damaged rock, estimates of permeability changes as a function of stress changes, and estimates of relaxed regions due to excavation. A numerical model of a jointed rock mass, the compliant joint model, was used to study the response of a jointed rock mass near a circular excavation. The joint constitutive model was linear elastic in shear and nonlinear elastic in the relationship between normal stress and joint closure. Initial stresses were geostatic and representative of stresses in a granitic jointed rock mass approximately 300 m deep. The joint properties were based on values that are representative of joints in granite. The results from numerical simulations were compared with analytical results for stress near a circular opening in an elastic medium having reduced Young's Moduli to simulate jointing effects on the rock mass. The differences in response between the two models resulted from the differences in the symmetry of elastic properties, the shear modulus of the continua, and the constitutive relations in the two models. For example, the non- linear anisotropic elastic continuum described by the compliant joint model resulted in vertical stress concentrations near the side of the circular opening that were 25 percent higher than the vertical stress concentrations in an isotropic linear elastic medium. The equivalent Young's Moduli in the vertical and horizontal directions of the compliant joint continuum deviated by as much as 50 percent from the values assigned to the modeled region prior to excavations. The perturbations of the elastic Young's Moduli extended to distances of two to three radii from the opening.

Jan 1, 1986

By D. Preece

The mining and oil industry employs techniques such as geologic mapping, core drilling and electromagnetic and potential field techniques to explore for new deposits. Since new discoveries of major near-surface ore bodies are becoming progressively rare and since traditional exploration methods are unable to resolve ore bodies deeper than about 500m, high-resolution seismic reflection techniques show great potential for exploration to much greater depths down to almost 3km.

Jan 1, 2011

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay

"The Analysis of Retreat Mining Pillar Stability (ARMPS) program was developed by the National Institute for Occupational Safety and Health (NIOSH) to help the U.S. coal mining industry to design safe retreat room and pillar panels. ARMPS calculates the magnitude of the in-situ and mining-induced loads by using geometrical computations and empirical rules. In particular, the program uses the “abutment angle” concept in calculating the magnitude of the abutment load on pillars adjacent to a gob.In this paper, stress measurements from U.S. and Australian mines with different overburden geologies with varying hard rock percentages were back analyzed. The results of the analyses indicated that for depths less than 650 ft, the ARMPS empirical derivation of a 21° abutment angle was supported by the case histories; however, at depths greater than 650 ft, the abutment angle was found to be significantly less than 21°. In this paper, a new equation employing the panel width to overburden depth ratio is constructed for the calculation of accurate abutment angles for deeper mining cases.The new abutment angle equation was tested using both ARMPS2010 and LaModel for the entire case history database of ARMPS2010. The new abutment angle equation to estimate the magnitude of the mining-induced loads used together with the LaModel program was found to give good classification accuracies compared to ARMPS2010 for deep cover cases.INTRODUCTIONIn the early 1990s, the Analysis of Longwall Pillar Stability (ALPS) was introduced by Mark (1990) as a chain pillar design software and was generally accepted and used by the U.S. coal mining industry. Following the success of ALPS, the National Institute for Occupational Safety and Health (NIOSH) developed the Analysis of Retreat Mining Pillar Stability (ARMPS) program for designing retreat mining pillars using a similar approach as ALPS (Mark and Chase, 1997). The Australian mining industry also recognized the success of ALPS, and Colwell, Frith, and Mark (1999) calibrated the program to Australian conditions. The ALPS and ARMPS programs draw their strengths from the large databases that are used to calibrate them (Mark, 2009). However, following the Crandall Canyon Mine collapse in 2007, NIOSH had to reconsider the pillar design criteria used in deep-cover retreat mining (Mark, 2010). The ARMPS overburden load prediction algorithm was improved to more accurately predict the loading of narrow panels with high overburden depths by implementing the pressure arch concept, and this new version is called ARMPS2010."

Jan 1, 2019

By John A. Rumball, Martin R. Houchin

Data from autogenous and semi-autogenous tests in a 1.67m diameter mill have been analysed in terms of a new model for rock abrasion. This model defines specific rate as the rate of production of fines (material less than 13mm) per unit mass of rock, where the term “rock” is applied to all ore pieces of ore coarser than the grate size. In most cases, an allowance for preliminary rounding and cleavage was made. A method has been developed to analyse steady-state data, using piecewise fitting to the cumulative size distributions. The model was compared to the well-established models, which are based on the population balance method. The specific rate increased for all rock sizes as the volume of 100mm balls in the mill was increased. Tests using a coarser feed and smaller balls provided further evidence that once rocks are rounded, abrasion is due to localised surface damage, caused by impacts, rather than rotational shear. An allowance for rapid rounding of jagged rocks, followed by less rapid abrasion of rounded pebbles was also tested, but the analysis was not conclusive. The average specific rate is a useful measure of the capacity of a primary mill. The effect of power intensity (kW/t of ore) was also investigated as a means of extrapolating the pilot-plant data to large production mills.

Sep 29, 2003