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By F. Calizaya

The Western Mining Training and Translation Center is a consortium of four universities devoted to developing educational and training projects to reduce the incidence/accident rates in the mining industry. The center consists of Missouri University of Science and Technology, Montana Tech, the Colorado School of Mines and the University of Utah. It is fully funded by the National Institute for Occupational Safety and Health (NIOSH). The center manages 15 educational projects and one translation project. Hazard Identification and Risk Assessment (HIRA) for Small Mines is an educational project managed by the University of Utah. The objectives of this project are to identify critical mine hazards, assess risks; outline procedures to reduce accident/incident rates in small mines in the western United States and to train mine workers and super-visors on HIRA in each participating mine. To accomplish these objectives, the HIRA personnel visited 15 small mines in Utah, Colorado, Nevada and California. At these mines, most operators were employing independent contractors to perform several tasks ranging from drilling, blasting and mucking of rock at the work site to hauling ore to a processing plant. The operation of diesel-powered equipment was one of the major tasks performed by the contractors. This article presents three case studies in which contract workers performed a number of tasks with no reportable incidents but did so unsafely. It also highlights the training received by most contractors, the constraints imposed upon them and the safety hazards to which they are exposed most frequently.

Jan 1, 2008

By F. Russo-Bello

TauTona Mine (the renamed Western Deep Levels East) is using a longwall mining method with stability pillars to extract the Carbon Leader Reef at a depth ranging from 2097 metres to 3456 metres below surface. Seismic activity and the complexity and hazards associated with negotiating geological features at great depth pose the most severe hazard to deep level gold mining in South Africa. It impacts negatively on safety and the profitability of the mining. The critical success factors in longwalling at depth are: ?Sound mining strategies and face shape control ?Ensuring follow behind development infrastructure is kept up to date ?Support system design as a strategy to effectively limit the serious effects of damaging seismicity. This paper describes the following: ?Implementation of the required mining strategies ?The evolution of a new stope support system and problems associated with its development and implementation. How this support system was adapted for the successful negotiation of major geological features with known seismic potential and problems encountered ?Current initiatives to optimize follow behind development infrastructure ?Current initiatives in designing a new gully support system and problems experienced with its development and implementation. Finally, this paper will deal with the way forward, briefly covering areas like the above-mentioned gully support system, the proposed phased introduction of technology, and the use of alternative mining layouts.

Jan 1, 2000

By I. A. Mwamba

Plain and alloyed titanium aluminides of composition Ti-47.5 at.% Al were prepared by melting commercial-purity titanium and aluminium with additions of 1 at.% precious metal. The as-cast alloys were subjected to potentiodynamic scans in 5, 15, and 25 wt% HCl aqueous solutions at room temperature and compared for their abilities to spontaneously passivate. Addition of precious metals resulted in a general improvement of corrosion resistance by increasing the open circuit potential to more noble values. Addition of 1 at.% gold or silver to titanium aluminide did not significantly increase the corrosion potential (Ecorr) above that of the plain titanium aluminide alloy, indicating that gold and silver are not sufficient cathodic modifiers to improve the corrosion resistance of titanium aluminide in all the solutions tested. However, platinum, palladium, and iridium additions shifted the corrosion potentials to the position of the passive region of plain TiAl for all solution concentrations. This indicated that TiAl alloyed with these platinum group metals would passivate spontaneously by cathodic modification. TiAl alloyed with palladium performed the best in 5 wt% HCl solution with the most positive corrosion potential. In 15 wt% HCl solution, alloys with platinum exhibited the most positive corrosion potentials, while alloying with iridium or palladium revealed more negative corrosion potentials. Thus, palladium alloying led to the best cathodic modification of titanium aluminide in low-concentration HCl solution, whereas platinum and iridium alloying additions resulted in the best cathodic modification of titanium aluminide in high-concentration HCl solutions.

Jan 1, 2013

By J D. St George

Flyrock is the unexpected and undesirable projection of rock fragments during detonation of a blast and managing this risk is becoming an increasing challenge as urban development encroach toward mineral extraction sites. The basic factors associated with flyrock are discussed in relation to geology, blast design and management. Published data of historical flyrock events are investigated and compared with local information on incidents. Good agreement was found between the data sets. A computer program of flyrock was developed using projectile motion based on launch velocity and angle, rock size, shape and density. This deterministic model was coupled with a probability model to simulate blast events. The impulse time ?t of the explosive pressure was taken as the stochastic variable to replace launch velocity. The variables were calibrated to replicate field data. A simulation was executed to demonstrate the capability of the approach to defining flyrock danger zones. For a simplistic situation it was identified that an appropriate distance to be cordoned off for the blast danger zone is a radius of 120 m in front of the blast and 65 m behind.

Jan 1, 2001

By S. J. Aresco

The Bureau of Mines has been active in promoting the purchase of coal for Government use under specifications that define the requirements in terms of heating value of the coal, expressed in British thermal units, and the composition as shown by proximate analyses. To these, when required, are added the ash-softening temperature, the free-swelling index, and the Hardgrove grindability index. Under most of these specification contracts the bidders guarantee the quality of the coal, and that guaranteed by the successful bidder becomes the standard of his contract. The deliveries are sampled in accordance with instructions 5/ issued by the Bureau of Mines; the samples are analyzed in the Bureau laboratory to determine whether the coal is of the quality guaranteed by the contractor and, if it is not, a price adjustment is made. Analyses of the delivered coal and tipple samples (samples collected at mine tipples as coal is being loaded into railroad cars or trucks) provide valuable data for use in evaluating future bids. In addition, the continuous sampling of coal as delivered is a check on the practical results obtained in burning the coal.

Jan 1, 1962

To optimise the haulage of the large tonnages produced from stopes in this high production gold mine, an ore pass and feeder system was introduced. This ore pass was designed and developed in 2006 and the plate feeder installed and commissioned soon after. Cavity Monitoring Surveys (CMS) have been undertaken of the ore pass since it was commissioned. The original thickness that was surveyed of the pillar between the ore pass and decline in 2006 was approximated to be 19 m. Throughout its use large blocks of rock were reporting at the plate feeder. This was the first indication that the ore pass was failing somewhere between the tipping level of the 9410 level and the plate feeder at the 9330. Even so CMS surveys were not taken systematically to firstly identify the area of failure or for monitoring. In August 2008 after implementing management systems, large blocks were still reporting to the plate feeder. The decision was made to empty the pass enough to carry out a CMS survey of the area. Subsequently it was discovered that the area of failure was between the decline and ore pass and that since the first survey the pillar thickness had been reduced to approximately 9.7 m. Further management controls were put in place to slow movement of the material and to limit the use of the 9410 tipping level, but this did not alleviate the rate of failure entirely. The results of an initial inelastic numerical model indicated that most of the pillar was damaged and that failure into the decline likely if no changes were made. The numerical modelling predicted that even if material was no longer passing through the ore pass interaction between the voids was unavoidable and would most likely result in instability. The ore pass and plate feeder are an integral part of the haulage system for the remaining life-of-mnie. Two solutions were sought to firstly reinforce and stabilise the pillar between the ore pass and decline and secondly to rehabilitate the ore pass or find an alternate design that could still utilise the plate feeder. This paper summarises the analysis and resulting rehabilitation that was used to accomplish this.

Nov 25, 2010

By Victor R. Spironello

The Bureau of Mines has studied slags associated with ferrous melting, to evaluate substitutes for the auxiliary flux mineral, fluorspar. A waste material recovered from alumina reduction cells, "used potlining," was tested in an 18-in-ID basic-lined cupola and in a 1/4-ton, pilot-scale, basic oxygen furnace (BOF). One potlining fraction, No. 1 cut, is a carbonaceous (graphitic) material and represents the bulk of the entire cell lining. The other fraction, No. 2 cut, is a refractory backup. These materials contain significant quantities of fluorine and other elements fn various compounds. The potlining fractions were blended in ratios believed representative of available materials. Both lump and pelletized forms provided extremely fluid slag. No adverse effects were found on furnace performance, chemistry or strength of the metals produced, or the environment. Fluorine recovery in the slag and fluorine losses to the scrubber water and solids were about the same as when fluorspar was used. There was apparently no lining loss from the cupola, but there was lining attack in the BOF. Since the fluidizing ability of potlining is due in part to sodium and other elements such as aluminum and silicon, less fluorine is needed in the slags when potlining is used.

Jan 1, 1983

By Stanley Lefond

The health of the industrial minerals industry appears to be improving year by year. In fact, preliminary information suggests that only three commodities (asbestos, bromine, and sulfur-other) failed to show a production gain in 1984 over 1983. In 1983, 10 commodities failed to post gains and six remained constant. The total value of industrial mineral production in 1984 was $16.7 billion, up $1.4 billion from 1983. Cement, clays, gypsum, and crushed stone production increased. Since these are used in housing and construction, this suggests these industries are moving out of their recent doldrums. Barite and clay production also improved in 1984. Local US barite production, however, still has a long way to go to attain the 1981 high of 2.6 Mt (2.9 million st). Barite imports remain high. But they are not as high as the 1982 importation of 2 Mt (2.3 million st). Bentonite production in 1984 increased 20% in volume and 25% in value over the previous year. This suggests improvement in the drilling industry and an increase in demand for iron ore pelletizing grades. All in all, the industrial mineral industry is alive and well. And it shows signs of great improvement in the next few years. Many thanks to those industrial mineral technical committee members for their assistance in assembling this industrial minerals review: M. F. Dibble, H. W. Allen, G. H. Gallup, S. G. Ampian, D. W. L. Spry, H. G. Fleshman, and L. J. Czel.

Jan 5, 1985

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Jan 1, 2008

By Peter Zhang

Pumpable cribs are being increasingly used in longwall tailgate entries and bleeders for roof support under abutment pressure. Their high peak capacity and confinement-controlled yield characteristics from reinforced-bags make them relatively safe and reliable standing supports for highly productive longwall panels. The pumpability of crib material over a long distance from the surface greatly reduces material handling and makes the underground installation more efficient and flexible. However, the performance of pumpable cribs is dependent upon crib properties, crib pattern, quality of installation, and geological and mining conditions. For proper crib selection and safe support design, it is critical to understand the basic technical requirements and performance of pumpable cribs in longwall tailgate entries and bleeders. Based on more than 10 years of experience using pumpable cribs for longwall standing support, the requirements and performance of pumpable cribs in longwall tailgate entries and bleeders under different mining conditions are presented. Experience shows that the load-displacement curve from full-scale lab testing is a good indicator of pumpable crib suitability for certain applications. Crib peak capacity, stiffness, and residual characteristics define its applicability to certain gate entries and bleeders under different cover depth. Crib pattern and spacing can be determined by abutment pressure level, roof, floor, and rib conditions, local roof stability, and ventilation requirements. To achieve proper in-situ performance, quality control of crib installation is also important.

Jan 1, 2012

By D. Christofferson

This final report documents the cooperative efforts of the Engineered Systems Division and the Mining Equipment Division of FMC Corporation during performance of Bureau of Mines Contract H0252091. Improved operator compartments incorporating overhead canopies, logical arrangements of controls, and design features offering greater operator comfort were developed and retrofitted to the FMC 320A roof bolter and 6L shuttle care The vehicles were evaluated in underground operations at separate mine sites. The successful evaluation has resulted in incorporation of the designs as standard features of the 320A roof bolter, and the 5L, 6L, and 10L shuttle cars.

Jan 1, 1978

By Grant Goddard

Cameco?s ?Double U? Strategy ? Double existing production ? To 40 million lbs U3O8 annually ? By 2018 ? Existing operations ? Saskatchewan ? United States ? Kazakhstan ? Projects under evaluation ? McArthur River expansion ? Kintyre (Australia) ? Inkai Block 3 (Kazakhstan) ? Development properties ? Cigar Lake project Measures of Success ? Safe, healthy and rewarding workplace ? Supportive communities ? Clean environment ? Outstanding financial performance

Nov 1, 2011

By A M. Purba and A Sjadat, A D. Wilson

Ongoing mining of the East Ertsberg Skarn System orebody below the Gunung Bijih Timur and Intermediate Ore Zone caves has been relatively successful and a ramp up to 80 000 t/d was achieved in the Deep Ore Zone Mine. However, the distribution of production, the overall drawpoint availability as well as the rate of change of the cave back profile has not followed the anticipated path. Several key factors influencing cave growth and progress, grade variation, fragmentation, wet muck and surface subsidence have been recognised over the life-of-mine that now have to be taken into account for planning the later stages of mining.The paper will discuss the influence of geological controls of lithology and structure – how this as well as other contributing aspects have impacted ore flow, contributing to some of the widespread panel damage, and been a key factor in compacted draw columns and caused an accelerated appearance of water ingress. This in turn has rapidly changed the distribution of wet drawpoints requiring remote mucking and special controls that also affect the call from the remaining dry drawpoints. Comment is also made on the potential final cave profile if it continues along the current trend and some discussion is offered about possible mitigation measures if this were to come about.CITATION:Wilson, A D, Purba, A M and Sjadat, A, 2016. Progressing cave performance into the later stages of the Deep Ore Zone mine, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 285–292 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By J. L. Hoover

The goal of the drill bit consumption study was to develop an equation(s) based on various components that affect bit drill wear and use the equation(s) to estimate the consumption of bits during production drilling. Drill bit data from various geographic locations were compiled and used to develop linear regression equations using the Microsoft Excel Software. Multiple drill bit consumption equations were developed and statistical analysis was used to verify the accuracy of those equations. The equations will allow mining companies to more accurately forecast drill bit consumption when evaluating new mine properties. This capability becomes increasingly beneficial to forecasting as the cost of consumables increases.

Jan 1, 2012

From time to time lead, which is normally considered to be a soft ductile metal, fails by cracking. The cracks which form appear to be of two distinct types-(a) those which follow a zig-zag course around the boundaries of crystals and (b) those which follow a more direct path, frequently two systems of cracks appearing at right angles.The former appear definitely to be associated primarily with the properties of the crystal boundaries, whilst the latter appear rather to be determined by the system of stresses applied. In both cases the lead appears to be a brittle rather than a ductile metal, for the failure occurs without the large amount of deformation usually associated with the tensile breaking of lead. The latter is, however, a function of time and when lead is deformed by long applications of stress, the metal appears to- be less ductile than when the stress is applied suddenly. Thus, if a given test piece of commercial lead is broken by a stress of 1100 lb. per sq. in. in several hours, it may extend 40 to 50%before failing, whereas when broken by a stress of 500 lb. per sq. in. acting for 12 months or more it willshow an extension of only 15 to 25 %. With lower stresses still, failure would take much longer and undoubtedly lower extensions would be obtained.As the stress is lowered it is found that the marked local contraction characteristic of ductile metals becomes less noticeable, and failure, instead of being by extension of *Lecture delivered to the First 'Ordinary Meeting of the Australasian Institute of Mining and Metallurgy, 21st August, 1935.

Jan 1, 1935

By M. M. Fine

Chalcocite flotation concentrates (26-34 percent Cu and 6-7 percent S) were roasted in air to determine sulfur elimination as a function of temperature, airflow rate, and residence time. Fluosolids roasting was evaluated batchwise and continuously in a 4-inch fluid-bed reactor. Kiln roasting tests were carried out in a 4-inch ID by 6-inch-long rotary drum. Pelletizing the very fine concentrates prior to roasting was essential to minimize dust losses and accretions. Concentrate pellets, sized minus 10 plus 20 mesh and bonded with 2 percent bentonite, were excellent feed for fluosol¬ids roasting and limited dust losses to 10 percent or less. One-fourth-inch pellets were best suited to kiln roasting. Both techniques were capable of producing copper calcines analyzing well under 1 percent residual sulfur when operated at 900° C with average retention times of 3 to 4 hours. Calcined pellets were strong, porous, and readily reducible.

Jan 1, 1970

By C. H. Herty

"The work on the physical chemistry of steelmaking, as carried out at the Pittsburgh Experiment Station of the United States Bureau of Mines in cooperation with Carnegie Institute of Technology and, the Metallurgical Advisory Board, has covered a series of fundamental studies on the action of oxygen in liquid steel and the application of the findings to open-hearth operation. The staff has varied from 14 to 23 men per year, some of whom have been Bureau of Mines physical chemists and metallurgists, others research engineers of the Metallurgical Advisory Board, and the remainder graduate students employed as research fellows at the Carnegie Institute of Technology. The work on the laboratory studies, and particularly on slag systems, has been in charge of G. R. Fitterer, associate metallurgist, Bureau of Mines, while the plant work has been under the direction of C. F. Christopher, research engineer of the Metallurgical Advisory Board. In all, some 16 physical chemists and metallurgists and 24 research fellows have taken part in the experimental work. Advantage has been taken of the opportunity for cooperation with the University of Pittsburg to have certain phases of this program made the subject of graduate work, particularly the laboratory studies on the deoxidation of steel with manganese-silicon alloys.The general method of studying oxygen has been, first, to determine the solubility of the various oxides in liquid steel. To date the solubility of Fe0 in steel has been determined, over a wide temperature range.3 The straight-line relationship in the plot of solubility versus temperature indicated that under a slag of 100 per cent Fe0 the oxygen was present in the steel as dissolved FeO. A further conclusion drawn from this work was that the solubility of FeO in the metal was proportional at a given temperature to the mol fraction of Fe0 in the slag if the slag contained a large excess of basic constituents. This result was determined from a study of the solubility of Fe0 under slags containing lime and iron oxide in various proportions. These data have been used extensively in determining the saturation of a bath of open-hearth steel under any conditions of slag composition, and results to be given later showed that the solubility curves are consistent with the results obtained in practice."

Dec 1, 1930

High Pressure Grinding Rolls Scale-Up and Experiences

Sep 13, 2010

By R R. Large

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The formation of orogenic gold deposits is commonly a two-stage process. The first stage involves preconcentration of gold in a suitable source rock, such as carbonaceous black shales, felsic volcaniclastics or komatiitic volcanics. The second stage involves the release of gold (commonly with arsenic and/or tellurium) from the source rocks during metamorphism or magmatic intrusive events, accompanied by the replacement of source rock pyrite by pyrrhotite. Recent research suggests that synsedimentary gold trapped in the structure of diagenetic arsenian pyrite in reduced carbonaceous facies of continental margin sedimentary basins is an ideal source of gold for many orogenic and Carlin-style gold deposits. The conversion of diagenetic arsenian pyrite to pyrrhotite during basin closure and greenschist facies metamorphism is the key process releasing gold and arsenic to the metamorphic fluid, with metal deposition in structural trap sites to form the deposits.CITATION:Large, R R, 2015. Orogenic gold deposits – a two-stage process of gold enrichment, in Proceedings PACRIM 2015 Congress, pp 335–340 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By Dimitrios Litsas, Michael Kavvadas, Panagiotis Sitarenios

"The paper discusses the results of 3D finite element analyses to investigate the effect of face support pressure and soil permeability on face stability conditions in EPB tunnelling. Two main types of face behavior are identified: Undrained conditions ahead of the tunnel face, typically occurring in soils with permeability less than about k=5·10-7m/s, are shown to be favourable for face stability; analyses show that an applied face pressure comparable with the initial hydrostatic pore pressure is sufficient for adequate face stability, even for soils of relatively low strength. Drained or partially conditions, typically occurring in soils with permeability higher than about k=5·10-6m/s, show that face stability is strongly dependent on the horizontal effective stress of the muck in the excavation chamber; the presence of even a small effective stress in the muck is very beneficial for face stability conditions and can be a possible explanation of successful EPB tunnelling through relatively permeable soils. INTRODUCTION Face stability is probably the most important aspect in urban tunnelling. Usually, Earth Pressure Balance (EPB) machines are utilized, to ensure that a reasonable amount of face pressure is exerted on the advancing excavation face, reducing ground loss and preventing potential instability of the face core, thus controlling the amount of disturbance caused by tunnel excavation to surface structures. However, the ideal amount of face pressure depends on several factors (ground and groundwater conditions, tunnel size and depth, muck conditioning, etc.) which are difficult to include in analytical calculations. For many years, conventional empirical or analytical methods (e.g. Anagnostou & Kovari 1996) were the only available tools for quantifying face stability and selecting the required face pressure. In the last few years, however, numerical analyses have gained power over conventional methods, allowing more accurate designs and providing useful information about the anticipated ground deformations due to tunnelling. Numerical analyses have also been used in assessing safety against face instability using face extrusion (horizontal face displacement) calculations (e.g. Prountzopoulos 2012)."

Jan 1, 2016