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By Alvin Kaufman

To aid industry and Government in planning water developments the Bureau of Mines in 1963 canvassed mineral producers to determine water use in 1962. Water use by the mineral industry comprises only 2 percent of the water withdrawn by industry as a whole. The major water-using mineral industries are natural gas processing, phosphate rock, sand and gravel, and iron ore. Analysis indicated the following conclusions: 1. Total water use and new water intake is dependent on the quantity of material processed and the process water requirements of the particular industry or commodity. 2. Recirculation is dependent on the processing requirements of a particular commodity, the manner in which the water is used, cooling and condensing requirements in certain large industries, the quality of new water intake, and the need to treat new and discharged water. 3. Consumed water is dependent on the quantity of water recirculated, the temperature and humidity in the area, and the proportion of water used for cooling and condensing. 4. A threefold increase in water use by the mineral industry by 1985 is forecast.

Jan 1, 1966

By R. Loots

High temperature creep resistant alloys are used in power plants as steam pipes and headers where they have to withstand super heated steam temperatures of up to 600ºC for the lifetime of the plant. Frequently certain portions of pipes may be subjected to excessive temperatures and consequently suffer from creep damage. Damaged portions then have to be replaced by welding new steel to the service exposed steel. Such welds will have to be subjected to a post weld heat treatment (PWHT) to relieve residual stress. During such a PWHT it is then possible for the steel to crack in the heat affected zone (HAZ) by what is known as Reheat Cracking (RHC), also known as stress relief cracking. For the present purpose the RHC susceptibility of service exposed steel was compared with that of new steel by reproducing the temperature cycle to which the HAZ will be subjected during welding in a weld simulator. This was followed by PWHT where the influence of residual stress was studied by applying a constant stress to such a weld-simulated specimen in a creep testing machine. A sharp spiral was machined along the length of the cylindrical specimen to simulate the presence of a possible weld defect. In this way it is possible to subject all portions of the simulated HAZ to the same stress intensity. For the purpose of the investigation 1/2Cr-1/2Mo-1/4V steel known to be susceptible to RHC was chosen. During testing a constant tensile stress was maintained while the temperature was increased to follow the thermal cycle prescribed for PWHT. It was found that the new steel in all instances failed by RHC along the grain boundaries in the high temperature heat affected zone. In the case of the service exposed steel the creep strength of parent metal was lower than that of new steel as a result of exposure to a steam temperature of 540°C for about 15 years. Consequently, all of the failures during testing occurred in a ductile manner by necking outside the HAZ rather than by RHC. When service exposed steel is welded to new material, it is expected that the creep strain associated with stress relaxation will mostly be concentrated in the parent steel outside the HAZ. Accordingly, it should probably not be necessary to make any extra allowance for reheat cracking during repair of steam pipes manufactured from 1/2Cr-1/2Mo-1/4V steel other than that applicable for welding new steel.

Jan 1, 2004

By E. Roman, M. Brissette

"For its end of life, Kemess Mine processed its difficult-to-treat hypogene ore from the east pit, which is a lower grade copper and higher pyrite content gold ore. Metallurgical testwork showed that improvement was possible with finer grinding. In July 2010, small grinding media were introduced into the regrind mill resulting to a nearly 50% decrease of the power draw with a finer grind, an increase of 1.1% in the final copper grade, an increase of 1.2% in copper and 2.9% in gold recoveries at the cleaners. The overall recoveries of copper and gold increased by 0.6% and 1.3% respectively.INTRODUCTION Kemess Mine was located at 430 km North West of Prince George in North Central British Colombia. Its production started in May 1998 by Royal Oak Mines and was acquired by Northgate Exploration in 2000. Kemess reached its end of life production with its last of production on March 9th, 2011. The concentrator nominal capacity was 56,000 TPD consisting of a primary gyratory crusher feeding two identical conventional SAB circuits totalizing 35,840 kW available for grinding. Rougher flotation consists of four parallel rows of eight 130 m³ tank cells to produce a concentrate of 2% Cu and 6 gr/t Au which was pumped to a 5 m x 10.8 m regrind ball mill. The regrind installed power was 4,480 kW to generate a final grind of 80% passing 40 microns. This product fed the two stage cleaner flotation circuit, first stage being two parallel rows conventional 17 m³ Denver cells and second stage, two 3.3 m x 11.0 m flotation columns in series, designed and constructed by Kemess, to produce a 22% copper concentrate grade. The cleaner flotation middlings were returned back to the regrind mill (Froehling, Mohns, Roman & Grady, 2005; Froehling, Mohns, Rajwani & Fullam, 2007). The large copper-gold porphyry deposit was composed of 15% of oxidized supergene and leach cap ore and 85% of hypogene ore having three major alteration zones, mainly characterized by their sulphur/sulphide content. Chalcopyrite is the main copper mineral species with minor bornite. Gold was mainly present as free gold or electrum. The first alteration zone consisted of a phyllic alteration characterized by moderate and intense sericite alteration which had the highest mill throughput and twice the mine grade when in contact with the supergene horizon. The second zone consisted of a potassic alteration characterized by more competent groundmass and moderate to intense quartz stockwork which had a slightly lower mill throughput. Finally, a third alteration found beneath and on the east side of the deposit having almost no quartz stockwork and hardened groundmass, sulphide content is two to four times more. This zone had the lower mill throughput with significant drop in the mine average gold and copper grade. Through its mine life, the average life of mine copper head grade was 0.21% Cu and 0.64 gr/t Au. Head grades began to decrease from 2008 to be substantially lower in 2011 as shown in Table 1."

Jan 1, 2012

By C. Machio

This paper reports the results of a study to determine the effect of process control agent (PCA) on the characteristics of Ti-Mg powders during milling. It has been shown that a 2% increase in PCA content leads to up to a 40% increase in yield of the milled powder but reduces the kinetics of the mechanical alloying process. The introduction of 4% PCA decreases the mean powder particle size by up to 30%, whereas 6% PCA increases the mean particle size by up to 230%. The characteristics of the milled powders will be discussed in relation to the role of PCA in the milling process. Keywords Titanium?magnesium, mechanical alloying, PCA.

Jan 1, 2011

By William J. Campbell

The purpose of this investigation was to extend the analytical services which are required for Bureau of Mines programs in developing and evaluating high-purity materials. Determination of trace elements in minerals and ores as an aid in geochemical evaluation and prospecting was considered. Serious consideration should be given to the merits of using the avail- able fluorescent X-ray spectrographic instrumentation and techniques for determining trace elements. Limits of detectability for trace elements in metal, oxide, solution or mineral samples range from 0.1 to 100 p.p.m, depending on the element being determined, overall sample composition, and the complexity of the X-ray spectra; limits of detectability range from 0.01 to 1 microgram for elements that have been preconcentrated by a chemical or physical process, for example, by ion-exchange membrane.

Jan 1, 1962

By Seong-Taek Yun, Chil-Sup So, Seung-Jun Youm

The greisen-bordered quartz veins are spatially, temporally and genetically associated with a late Cretaceous granite intruding an early Palaeozoic, limestone-rich formation. Paragenetic data indicate three stages of hydrothermal mineralisation, the first typically in or within 500 m of the granite stock and showing vertical zoning between deeper Mo and higher W, the other two more abundant in the upper and outer parts of the mineralised area. Fluid inclusion and stable isotope data indicate that hydrothermal veins formed over a temperature range of 120-395 degrees C from low-salinity fluids. Early wolframite and molybdenite deposition in stage I occurred at temperatures of 300-395 degrees C from magma-derived, low-salinity, CO2-rich fluid, mainly as a result of fluid boiling and associated CO2 effervescence. Progressively larger volumes of local meteoric water entering the system led to the successive deposition of scheelite in stage I vugs, stage II base-metal sulphides and stage III fluorite-carbonates. Remobilisation of W and its later deposition of scheelite in the outer and shallower portions of the system occurred at 215-275 degrees C and stage II and III mineralisation from lower-temperature fluids that became progressively lighter in stable isotopes. It is concluded that the influx of meteoric water increased with distance from the cooling pluton

Jun 18, 1905

By Daudi R. Waryoba

Keywords: Wire Drawing, OFHC Copper, Recrystallization, Orientation imaging Microscopy (OIM), Texture inhomogeneity Textural inhomogeneities have been investigated in oxygen free high conducting (OFHC) copper wire drawn at room temperature to a true strain of 2.31, and isothermally annealed at 250°C and 750°C for annealing times ranging from 10 s to 1 hr. Local orientations were mapped by means of orientation imaging microscopy (OIM). The microtexture of the drawn wires showed a strong <111> + weak <100> duplex fiber texture at the inner core, whereas the mid and surface regions had a comparatively weak texture. Annealing at 250°C resulted into a recrystallization which originated from the mid section, proceed towards the surface regions and ends in the inner core. Recrystallization resulted into a strong<100>+weak<111> duplex fiber texture. A similar textural inhomogeneity was observed during short annealing at 750°C. However, prolonged annealing gave rise to abnormal grain-growth that proceed from the inner core to the outer surfaces with a dominant <111> fiber component at the inner region and mixed components of <111>, <100>, and <112> at the outer surfaces.

Jan 1, 2005

By P. Rosario

The application of high pressure grinding rolls (HPGR) has been growing in the mining industry for the last 10 to 15 years. The major benefits supporting this trend are better energy efficiency, improved grinding capacity, and higher metal recovery in downstream processes such as heap leaching and flotation. In general there is limited quantitative knowledge on the true benefits of HPGRs relative to SAG mills in comminution, and about which situations one deals with better than the other. This paper will present a structure for comparison of the energy requirements for HPGR versus SAG mill considering complete circuits for comminution of precious and base metals hard ores. The work presented is from the design of four complete circuits based on ore data from two sites.

Jan 1, 2010

By Gale D. C, Giles D. E, James Parker A

A zinc-lead-iron sulphide concentrate from McArthur River can be leached at atmospheric pressure in aqueous solutions with HCI, FeCl3, CuCI2, HCl/02 HC1/02/ CuCI2, Cl2, and CI2/CuCI2 to extract &gt;98 per cent zinc, &gt;90 per cent silver, and &lt;15 per cent iron in 1-7 h at 95&#176;. Galena is more reactive than sphalerite and is converted to lead chloride. Except for leaching with HCI, elemental sulphur is the the main by-product of the leach with little or no sulphate ion detected. The rates and products of these reactions are compared. It is recommended that aqueous slurries of the concentrate (50 per cent w/v) are leached at atmospheric pressure with Cl2 in the presence of copper(II) chloride, which accelerates the extraction of zinc. Hydrochloric acid is a by-product which is consumed bYJeaction with fresh concentrate in air at 60&#176;. Copper is subsequently stripped from the leach solution and recycled by reaction with fresh concentrate at 100&#176;. The separation, recovery, and electrolysis of zinc and lead chloride is outlined since this process would produce Cl2 gas which can be recycled and used directly in the leaching stage.

Jan 1, 1976

Six years ago it was my responsibility to give the presidential address, and now that privilege has come my way again. On that occasion we were gathered atWhyalla and I took the opportunity of putting before members some observations on the problems we then faced in Australia, and some gains we had been able to chalk up during those early postwar years. You may recall that we were then just beginning to emerge from the difficulties of postwar reconstructions - of fuel and power shortages, and a grossly inadequate manpower for the tasks facing us. The expansion of manufacturing industry in Australia greatly stimulated by wartime activities, and by the postwar developments of many institutions - ran ahead of the supporting services and of the supply of labour to do thejobs. For many industries this unfortunate state of affairs had meant operating at rates well below their installed capacities and therefore levels of output insufficient for Australia&apos;s rapidly growing needs. When I last spoke to you these problems were still with us, though their intensity was beginning to lessen as a result of vigorous efforts by industry and governments to reduce them. But I must confess that my thoughts on that occasion were full of those problems and of the effect they were having on the iron and steel industry of this country-of the coal shortage, the labour shortage, and the many other difficulties we were then faced with - and all at a time when the stepping up of our postwar expansion programme was of vital concern to all of us.What a different picture it is six years later! How greatly has the scene changed for the better. In 1951 Australia&apos;s postwar immigration programme was only in...

Jan 1, 1957

By M. du Plessis

This paper describes the evaluation of a limit equilibrium model to simulate the behaviour of crush pillars in platinum mines. An analytical model was derived to calculate the residual average pillar stress (APS) values of the crush pillars. The values predicted by this model were compared to the numerical values obtained from TEXAN simulations. In general, the limit equilibrium model appears to be very attractive for simulating pillar failure as the gradual crushing of the outside of the pillar and the transfer of stress to the intact core can be replicated. The value of the TEXAN crush pillar model was further demonstrated by simulating an idealized layout with crush pillars between two adjacent panels. The simulations illustrated that oversized pillars will not crush close to the face and this may lead to seismic failure in the back area. An important finding of the study is that closure measurements may prove to be a very valuable diagnostic measure in crush pillars layouts. Distinct differences in magnitudes of closure were simulated for a scenario where the pillars crush as expected, compared to the scenario where an oversized pillar is left. Experimental closure data collected in a crush pillar stope provided further evidence regarding the value of closure measurements in these layouts.

Jan 1, 2011

The Ross Goldfield is a unique gold resource. From its discovery in 1865 there was a growing understanding of the geology of the deposit as mining moved from the surface at Jones Flat in the south to underground extraction at Ross Flat further north. Multiple layers of gold-bearing gravels were penetrated in deep shafts and six rich gold-bearing bottoms were worked. The ultimate bottom on the Old Man Group gravel surface has never been reached either at Jones Flat or at Ross Flat. Mining at Ross Flat was ultimately halted by uncontrollable flooding, by escalating costs and by a shortage of labour during the First World War. Mining at Jones Flat recommenced in the late 1980s and attention has once again been drawn to the potential magnitude of the resource at Ross Flat. Research of historic literature shows that there is potential for a bulk gold grade of 400 mg/m3 in a volume of 82 million cubic metres of gravels which lie in a basin between Jones Flat and north Ross.

Jan 1, 1993

As requested by my fellow members I shall endeavour to read a paper on the Treatment of the Auriferous Sulphide Ores of Kalgoorlie from a metallurgical standpoint. Before going into the subject it will be necessary for me to make some slight explanations. As most of you are aware many of our London friends holding interests in some of these mines think it necessary that their doings should be shrouded in mystery. The result of this is that in some cases it is impossible to get any official figures from the mines for this paper. I must ask you not to lose sight of the fact that in treating the sulphide ores of the Kalgoorlie mines we had absolutely no data to start on, and that to-day. we are just beginning to get a fair grip of tbe subject; and though very high extractions were announced when the first plants were started on sulphide ore, it was well known that this was only to maintain the public confidence in the mines to allow time to make necessary adjustments; and, as showing the advance made in the treatment of the ore, I notice several of the mines have published the results of the final or re-treatment of the residues of their first sulphide treatment. I have also had the opportunity afforded me of thoroughly sampling the residues of several of the principal mines, and find the results as now published absolutely correct.

Jan 1, 1901

By Mario Paredes Malca

El presente trabajo describe el sistema experto por circuito de molienda desarrollado en la Concentradora Toquepala, donde se aplica la tecnología del control experto-basada en conjuntos difusos y modelos matemáticos. Las ventajas que presenta la aplicación de los sistemas expertos comprenden el aumento de la capacidad de tratamiento, la disminución del costo de operación y del porcentaje del producto que no cumple las especificaciones. Asímismo, el sistema experto se integra al sistema de control para asistir al operador en la detección de los casos, donde las operaciones se puedan descontrolar, aunque ésta no sea realmente su principal función. Si se le configura y sintoniza correctamente, el sistema experto detecta, diagnostica en forma continua los problemas de operación y aplica modelos matemáticos y estadísticos de optimización, interactúando con un sistema de control DCS o PLC.

Sep 12, 2005

By P. van der Heever

The genesis of Simrac?s technology transfer practices and future initiatives are discussed in this report. Simrac produces a relatively large volume of research-based health and safety outputs that must be communicated to its tripartite stakeholders and to other interested and affected parties. The wide spectrum and complexity of the transferables, together with the diversity of end-user needs and skills of the recipient population, necessitate customizing of technology transfer practices. In an attempt to optimize this process, a variety of practices are employed. Information is transferred by means of publication and dissemination of a variety of reports (paper, website and CD).Knowledge transfer is directed to health and safety practitioners by means of workshops, seminars, best practice booklets, handbooks and textbooks. Promotion of recently completed projects is directed at targeted recipients by means of project launches. Only isolated cases of piloting and implementation of technologies have been undertaken due to lack of resources and past organizational focus. It is anticipated that Simrac?s new organizational structure and management system, together with end-user involvement by industry champions, from project inception to implementation, will enhance the effectiveness of the technology transfer process.

Jan 1, 2002

By T. Robyn

The Bolivar Cu-Zn skarn deposit in the Piedras Verdes mining district of Chihuahua is one of many base and precious metal deposits in the north-northwest trending Sierra Madre Belt extending across the states of Chihuahua, Durango and Sonora in northwestern Mexico. The information presented here is drawn from two papers on the Bolivar property1,4 plus new information. Historical mining and exploration for polymetallic Cu-Zn-Pb-Ag-Au deposits in the Sierra Madre Belt dates to Spanish Colonial days. Access to the Bolivar Property (Figure 1) is by paved road (approximately 305 km southwest from Chihuahua) and approximately 80 km by all-season gravel roads to the villages of Cieneguita (population ~1,000) and Piedras Verdes (population ~500) adjacent to the property. [ ] The Bolivar property comprises an old Cu-Zn producer (the Bolivar Mine), exploration concessions adjacent to the Bolivar Mine, and a new camp at the mine that supports mining, exploration and maintenance. From 1980 to 2000, former operators extracted approximately 300,000 tonnes of ore that was hand sorted to an average grade of 5-6% Cu and 25-30% Zn. From 2005-2010, Dia Bras produced from the mine just over 600,000 tonnes at an average grade of 1.7% Cu and 8.6% Zn. Dia Bras transported mineralized rock 275 km by truck and rail to its Malpaso mill, located about 30 km west of Cuauhtémoc in central Chihuahua, from 2005 until October 2011 when the new Piedras Verde mill was commissioned only 5 km from the mine. (Note: the Company cannot use the term ?ore? because it has never completed a feasibility study.)

Jan 1, 2012

Extraction of copper from a chalcopyrite concentrate by means of a cupric chloride leach is described. Initially the concentrate is mixed with sulphur in a 9: 1 ratio, and heated to 350-400&#176;C. This converts the chalcopyrite into more readily soluble sulphides, while the iron rejected from the chalcopyrite is transformed to pyrite, which is insoluble in the leach solution employed. A two-stage countercurrent leach using a leach solution of 1 M cupric chloride, 3.5 M sodium chloride, and 1 M hydrochloric acid extracts 98 per cent of the copper, if conducted at 105&#176;C. Iron extraction is relatively low, the mole ratio of iron to copper extracted being 0.17, compared to 1.0 for unreacted chalcopyrite. After dissolution, copper can be recovered by electrolysis, cupric chloride leach solution being regenerated at the anode.Electrolytic power requirement is low, being approximately 0.5 kWh per kg of copper. The sulphur and, most of the iron entering as feed are rejected from the process in the form of pyrite, which remains in the leach residue.

Jan 1, 1974

By G. Styles, K. Lye, R. Tenbergen, M. Mayhew

Requirements to produce a higher bulk concentrate grade have necessitated extensive use of on-stream analyser data and computer programming to obtain the targets set for Frood Stobie Mill as required by overall milling considerations.A brief history of milling process changes associated with the changes from a low grade (3% Ni) bulk concentrate in the sixties to a high (6 to 8.5%). Nil bulk concentrate currently will be presented.The heart of the presentation will be the computer control schemes for ""self optimisation"" of the distinct flotation sections in the plant within certain restraints to achieve targets set by the mill metallurgist.

Jan 1, 1985

By John C. Machisak

THIS PUBLICATION presents injury experience and employment data for the nonmetal mining industry in the United States for 1960. Data for the mines and mills are presented separately, as shown in the list of tables. The Bureau of Mines is authorized to send questionnaires to the nonmetal mining industry for their voluntary submission of injury and employment data. As the information submitted contributes largely to the promotion of safety in the industry, special effort has been made to make these questionnaires as complete as possible.

Jan 1, 1963

By T. T. Campbell

This investigation was undertaken by the Bureau of Mines to determine if high-purity ductile molybdenum of uniform quality can be prepared by bomb reduction techniques. Molybdenum trioxide (Mo03) and mixtures of molybdenum dioxide (Mo02) and trioxide were reduced by magnesium in closed steel bomb reactors. Very high pressures in excess of 5,000 p.s.i. developed during magnesium reductions of Mo03. The addition of calcium oxide to the bomb charges caused a significant reduction in pressure. Reaction efficiencies higher than 95 percent were obtained during magnesium reductions of either Mo03 or mixtures of Mo02 and Mo03, but high pressures were encountered when the charges contained from 30 to 60 percent Mo02. Reaction efficiencies were usually low (less than 80 percent) when the bomb charges contained more than 40 percent Mo02. The charges with the best composition considering operating pressure, yield, and purity contained from 10 to 30 percent Mo02. The addition of small amounts of carbon to the reaction charges was helpful because some deoxidation occurred during the reduction and also during subsequent arc-melting of the regulus metal. The purity of the bomb-reduced, arc-cast molybdenum metal ranged from 99.2 to 99.8 percent and in some instances the metal contained as little as 5 parts per million oxygen.

Jan 1, 1962