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"The Antamina orebody is a geologically complex, polymetallic skarn located in the Peruvian Andes. A range of copper and copperzinc ores are campaigned separately through the concentrator producing copper and zinc concentrates as well as lead and molybdenum by-products.Ore type classification at Antamina is based on chemical assays; geological or lithological variations are not considered. Consequently, each ore type demonstrates a wide range of concentrator performance.Antamina is currently implementing a system to track ore movements from the shovel face to the concentrator. Relatively sophisticated models of run-of-mine and coarse ore stockpiles ahead of the mill will be used to account for the variable delay between mining and milling.One of the numerous benefits of ore tracking is the ability to directly associate the measured concentrator performance with a physical location in the orebody. Using existing mine planning software, three-dimensional models can be created of such performance measures as metal recovery, concentrate quality, operating cost, and reagent consumption.Such a database of information allows the true value of different ore sources to be determined. The value of an ore source is conventionally measured as metal content (individual or combined ‘equivalent’) with a cut-off grade applied to indicate some base level operating cost. However, the true value of an ore source should include dynamic factors like haulage cost, reagent use, smelter penalties, and even prevailing metal prices. If an ore source is considered from all perspectives rather than simply its potential to deliver value, the concept of ore becomes more flexible. Any ore source that reveals itself to be unprofitable can be eliminated from the mine plan or possibly blended with others to improve its value. The ability to predict mill performance, and thereby profitability, is a powerful tool.In an operation such as Antamina, where managing ore sources, campaigns, and the delivery of a variety of concentrates in a narrow timeline to customers is a daily challenge, prediction is the key to success and improved efficiency."

Jan 1, 2005

By John M. Jennings

The Hiawatha pitmouth is located at the head of the middle fork of Miller creek, at an altitude of approximately 8,100 feet. The coal is lowered to the tipple down a gravity tram two miles long, in 16 car trips which carry about 60 tons net. The average running time of the trip is about nine minutes. At the bottom of the tram there are two tracks for the loads and two for the empties. The trip, after being cut off the rope, is dropped by gravity onto a feeder, where the cars are uncoupled and fed one at a. time across the pitcar scales and into the dump. The dump is of the full revolution type, friction driven. On passing out of the dump the cars are returned by a kickback onto the empty car haul. They are caught at the bottom by a brake which consists of a pair of steel squeeze blocks operated by heavy springs. The empties are elevated to the double tracks, where the trips are made up. The feeder, scales, dump and empty car haul are all built on the solid ground of the hillside instead of being incorporated in one structure with the screening plant. The coal is dumped into a small hopper with a capacity of about 15 tons, whence it is fed by a reciprocating feeder onto a scraper conveyor which takes the coal over to the screening plant, a distance of 120 feet. This conveyor is 48 inches wide and travels at 120 feet per minute. The screen itself is interesting. It is of the Marcus Horizontal type.

Jan 1, 1925

By I. L. Nichols

The Bureau of Mines is investigating vanadium recovery from various domestic resources. A low-grade carnotite ore sample containing 0.34 pct V20S and 0.04 pct U30S was leached, and the results were compared with those of similar leach tests using a mill-grade sample containing 1.26 pct V205 and 0.19 pct U308. Experiments on the two ore samples included agitation leaching, pugging, salt-roast leaching, and autoclave leaching. None of the leaching methods investigated showed any significant advantage over another except autoclaving, which reduced the leach time but at the expense of greater acid consumption. Typical of the experimental results attained are the data gathered from the agitation leaching tests. Agitation leaching of the mill-grade ore with 100 lb/ton H2S04 resulted in extractions of 94 pct U308 and 33 pct V205. To increase the vanadium extraction from 33 to 91 pct required concentrations of 300 lb/ton H2S04 and 10 lb/ton NaC103. Complementary agitation leach tests on the low-grade ore required 200 lb/ton H2S04 to extract 94 pct U308, and 300 lb/ton H2S04 and 10 lb/ton NaC103 to extract 90 pct V205.

Jan 1, 1983

The Department of Mineral Resources has a statutory responsibility for mine site environmental management and control in New South Wales. The Department is focusing on an environmental management systems (EMS) approach, as a management tool to achieve æbest practiceÆ outcomes both within our own organisation and on mine sites. One of the DepartmentÆs tools is environmental risk assessment and management which incorporates the Mining, Rehabilitation and Environmental Management Process (MREMP). The MREMP was introduced many years ago in order to integrate the activities of key stakeholders who may have a material interest in the mine. The MREMP provides an ideal mechanism by which ædue diligenceÆ requirements and environmental management programs can be integrated into mining operations and lead to a sustainable post-mining land use.

Jan 1, 2002

By Derek T. Novotny

Blasting cap precision is of major importance to advanced blast design. New technologies have introduced the use of microchips within detonator casings in the hope that the electronic control of the timing element will increase the accuracy of the cap and therefore improve the results of perimeter control blasting applications. The electronic detonators have proven to be accurate in timing tests done to date. They boast cap scatter results to microsecond accuracy, but the question is, does this degree of accuracy actually prove to be beneficial during practical use in the industry, or are conventional technologies applicable.

Jan 1, 1998

By R. Hansson

Zincite and spinel phases are present in the complex slag systems encountered in zinc/lead sintering and zinc smelting processes. These phases form extensive solid solutions and are stable over a wide range of compositions, temperatures and oxygen partial pressures. Accurate information on the stability of these phases is required in order to develop thermodynamic models of these slag systems. Phase equilibria in the Fe-Zn-O system have been experimentally studied for a range of conditions, between 900 and 1580°C and oxygen partial pressures (pO2) between air and metallic iron saturation, using equilibration and quenching techniques. The compositions of the phases were measured using Electron Probe X-ray Microanalysis (EPMA). The ferrous and ferric bulk iron concentrations were determined using a specially developed wet-chemical analysis procedure based on the use of ammonium metavanadate. XRD was used to confirm phase identification. A procedure was developed to overcome the problems associated with evaporation of zinc at lowpO2s and to ensure the achievement of equilibria. An isothermal section of the system FeO-Fe2O3-ZnO at high ZnO concentrations at 1200°C was constructed. The maximum solubilities of iron and zinc in zincite and spinel phases in equilibrium were determined at pO2=1*10-6atm at 1200and 1300°C. The morphology of the zincite crystals sharply changes in air between 1200 and 1300°C from rounded to plate-like. This is shown to be associated with significant increase in total iron concentration, the additional iron being principally in the form of ferric iron. Calculations performed by Fact Sage with a thermodynamically optimized database have been compared with the experimental results. Keywords: Phase Equilibria, Fe-Zn-O, Quenching, Wet-Chemistry, EPMA, XRD, ISP.

Jan 1, 2004

By William G. Pariseau

The research reported here concerns ground control in multi-seam mining and relates mainly to the analysis of stress about multi-seam room and pillar mines. The two variables of most concern are distance between seams and alignment of pillars between seams. The first year results of the two year investigation apply to the development phase of mining when the extraction is typically less than 50 percent. The second year effort is concentrated on retreat mining and barrier pillar safety when the extraction is typically greater than 50%. The procedures in both cases are readily applicable to more than two seams. Although two seam mines are the most frequent type of multi-seam operations, because of the potential for three seam operations in the Western United States, a number of three seam development analyses are also presented. The results enable one to make a rapid assessment of stability of rooms and pillars during production and barrier pillar size during extraction of adjacent panels. In the event of low safety factors, more site-specific analyses can be undertaken.

Jan 1, 1983

By Jan Krason

Introduction The South African Witwatersrand is the world's largest gold-bearing province and gold producing district. But the production capacity of the very big mines (Vaal Reef, Driefontein Cons., Western Holdings, Western Deep Levels, and others) is not as large as that of the Soviet Union's Muruntau gold mine. The Muruntau gold-bearing ore body was discovered in 1956-57, and has been in operation since that time. In the early 1970s, this mine produced 70 to 80 t/a (2.3 to 2.6 million oz per year) of gold. Within the last few years it has become the world's largest producing gold mine. Russian authors do not publish ore reserve or rate of production data. But it is very likely that the Muruntau mining complex, which involves several ore bodies, is currently producing about 143 t/a (4.6 million oz per year) of gold. The author has confirmed this by personal information from several knowledgeable Russian geologists. This amounts to one-half of

Jan 11, 1984

By H. Thomas, I. N. M. Nday

"*Paper written on project work carried out in partial fulfilment of BSc.Eng (Mining Engineering) degreeThe optimization of the cycle time to increase productivity at Ruashi Mining was an on-site project investigation conducted from December 2017 to January 2018. The main objective of the investigation was to determine the actual cycle time components. These comprise the queuing time, loading time, hauling time, and dumping time of ADTs (articulated dump trucks). Data was also collected on two hydraulic shovels to identify where these could be optimized. The actual cycle time obtained was compared to the theoretical cycle time to ascertain the constraints, which affect productivity. By applying Systems Thinking to the cycle time, man-made constraints of the actual cycle time were identified and recommendations to mitigate them suggested to optimize the cycle time. However, rainfall is a natural constraint that adversely affects the actual cycle time, and which cannot be mitigated.In addition, both the theoretical and actual cycle time were analysed, as well as the theoretical and actual productivity for the year 2017. Based on the findings, conclusions and recommendations were made to optimize the cycle time. IntroductionThe aim of the project was to increase Ruashi Mining’s productivity by 39%, from about 7.537 Mt of broken material mined to 12..367 Mt targeted in 2017.BackgroundRuashi Mining is a copper (main commodity), cobalt (by-product), and acid producer located in the Democratic Republic of Congo (DRC), 10 km from the city of Lubumbashi. Figure 1 shows the location of the Ruashi Mining open pit.Ruashi Mining started production in 1911 and in 2005 the mine was registered as a DRC company owned by Metorex Group Limited. In November 2013, the Chinese company Jinchua Group International Resources Co. Ltd. bought a majority stake in the business. Ruashi Mining currently produces 38 000 t of copper and 4400 t of cobalt per annum. The company has contracted NB Mining (a contract mining company) to transport its ore. To optimize the productivity of the mine, considering ore flow delivery at Ruashi Mining, the cycle time is the indicator used in this project. The cycle time components were recorded on site using a stopwatch."

Jul 1, 2019

By G. A. Kordosky

The concept of selectively extracting copper from a low-grade dump leach solution followed by stripping the copper into an acid solution from which electrowon copper cathodes could be produced occurred to the Minerals Group of General Mills in the early 1960s.This simple, elegant idea has resulted in a technology by which about 2.2 million tonnes of high quality copper cathode was produced in year 2000. The growth of this technology is traced overtime with a discussion of the key plants, the key people and the important advances in leaching, plant design, reagents and electrowinning that have contributed to the growth of this technology. Some thoughts on potential further advances in the technology are also given.

Jan 1, 2002

By Y. Wang, K. S. E. Forssberg, A. Johansson, X. Lu

This paper presents the desulphurisation of a fine graphite powder below 80 µm by high gradient magnetic separation. The removal of pyrite (FeS2) and pyrrhotite (Fe1-xS) from the fine graphite powder was carried out with a modified version of the Sala dry HGMS with vibration assistance. In addition, a comparative study with dry and wet HGMS systems has also been performed. In order to establish the performance of the desulphurisation, a test series of varying parameters such as magnetic induction, matrix type/configuration and vibration were performed. As an equipment/process for removing sulphurous impurities from the fine graphite powder, the dry HGMS is deemed useful and comparative with the wet HGMS.

Jan 1, 2002

By W. H. Langer

Quaternary alluvial deposits in the modern stream channel of the Agua Fria River west of Phoenix, Arizona, is mined and processed to provide aggregate to the greater Phoenix area. The alluvium occurs as three distinct, stratigraphic layers. Samples from each of the three layers were examined for gravel clast lithology, grain size distribution, and engineering tests, which revealed noteworthy differences in properties. The engineering properties of the gravels are dependent on a combination of factors including gravel clast lithology, the distance the gravel clasts have been transported from their source areas, the age of the deposits, and the degree of weathering. Knowledge of these characteristics may affect the methods that aggregate operators employ to extract and process gravel from the Agua Fria River.

Jan 1, 2010

By S. Djaroudib

Keywords: iron recovery, hematite, hydrolysis, polyvinylpyridine The precipitation of hematite (Fe2O3) from ferric chloride media at ambient pressure and at temperatures below 100°C was studied as part of a program to recover a marketable iron product form metallurgical processing streams and effluents. Hematite was formed in preference to ferric oxyhydroxides (e.g. ß-FeO?OH) by controlling the precipitation conditions, especially by adding Fe2O3 seeds. The extent of the hydrolysis reaction was found to depend primarily on temperature and on the free-acid concentration. As to the latter, the controlled addition of an adsorbent of free acid, in our case the anionic resin polyvinylpyridine (PVP), was shown to allow for the nearly complete removal of iron as readily filterable Fe2O3. The hematite product typically contains = 63% Fe and = 1.5% Cl, and its composition does not change appreciably upon repeated cycling using regenerated PVP.

Jan 1, 2005

By J. E. Urosek, K. L. Cashdollar, D. A. Beiter, G. J. Shemon

The National Institute for Occupational Safety and Health (NIOSH) and the Mine Safety and Health Administration (MSHA) conducted joint research to evaluate explosion blast effects on mine ventilation stoppings at the NIOSH Lake Lynn Experimental Mine (LLEM). After mine explosion accidents, MSHA conducts investigations to determine the cause(s) as a means to mitigate or eliminate future occurrences. As part of these post-explosion investigations, the condition of underground stoppings, including the debris from damaged stoppings, are documented as evidence of the strength and the direction of the explosion forces. The LLEM data showed that a total pressure of 52 kPa (7.6 psi) destroyed the solid concrete block stopping, ~36 kPa (~5.2 psi) destroyed the hollow-core concrete block stopping, and 9 kPa (1.3 psi) destroyed the steel panel stopping. These results will assist investigators in determining the explosion forces that destroy or damage stoppings during actual coal mine explosions.

Due to the reduction in frictional resistance, the characteristics of joint infill have a major role to play on rock mass instability. This paper elucidates the laboratory behaviour of graphite infilled joints based on a large-scale shear apparatus under constant normal stiffness (CNS). Discontinuities in a graphite mine have been simulated using high strength gypsum plaster, with the joints infilled with compressed graphite layers. The Fourier series was employed to characterize the joint surface, and the method of using the Fourier coefficients for the prediction of dilation and normal stress with displacement is explained. Die Charateristik von Verbindungsfüllungen hat einen wichtigen Einfluss auf Felsmassenstabilität aufgrund der Verringerung des Reibwiderstandes. Diese Veröffentlichung untersucht das Verhalten von Graphit gefüllten Verbindungen im Labormaßstab mit einer Methode die auf einem gross-umfang Apparateprinzip unter konstanter Normal-Steifheit beruht. Unregelmässigkeiten in einer Graphit Mine wurden mit Gypsum Gips von hoher Festigkeit simuliert, die Verbindungen wurden mit kompressierten Graphitschichten gefüllt. Fourier Reihen wurden angewandt um die Verbindungs Fläche zu Characterisieren. Die Methode die verwendet wurde um die Ausdehnung und Normalkraft mit der Verschiebung mit Hilfe von Fourierkoeffizienten zu bestimmen wird beschrieben. Due à diminution des résistances par frottements, les caractéristiques des joints de remplissage joue un rôle majeur sur les instabilités de la roche. Cet article étudie le comportement en laboratoire des joints de remplissage en graphite à partir d’un dispositif de friction à grande échelle sous rigidité normale constante. Les discontinuités dans une mine de graphite ont été simulées grâce à un plâtre à haute teneur en gypse, dont les joints ont été remplis de couches de graphite compressées. La caractérisation de la surface des joints a été réalisée avec les séries de Fourier. La méthode utilisant le coefficient de Fourrier pour la prédiction de la dilatation et de la compression en fonction du déplacement est également expliquée.

Jan 1, 2003

By Dunkin H. H

Flotation tests in a laboratory flotation cell at pH 8-8.5 and other sui}plementary tests were carried out on sphalerite-quartz and sphalerite-pyrite-'quartz mixtures. Depression of sphalerite resulted vvhen the addition of copper sulphate exceeded an optimum value. Evidence that this depression was due to collector deficiency because of reaction between copper sulphate and collector is presented. It was not found possible to devise a direct analytical method suitable for .plant control of reagent addition. A possible means of control using potassium cyanide and a higher pH is discussed.INTRODUCTIONAt the Rosebery mill of Electrolytic Zinc Co. of Australasia Ltd., recovery 'of sphalerite in the zinc concentrate occasionally drops unexpectedly. Experiment in the mill and mill laboratory has suggested that one of the factors causing this may be the addition of more than an optimum amount of copper sulphate. On such occasions, an increase in the collector addition improves zinc recovery in the zinc concentrate, but only at the expense of sphalerite-pyrite selectivity.To examine the phenomenon more thoroughly, and if possible to find some means of determining the optimum addition of cepper sulphate and controlling the actual addition at this optimum, one of the authors (W. J. T.) was employed by the Electrolytic Zinc Co. of Australasia' to carry out investigations in the Mining Department, University of Melbourne, under the joint direction of R. B...

Jan 1, 1954

By R. G. Benson, P. R. Wotruba, K. W. Schmidt

Introduction The Fortitude deposit is located in Nevada's Battle Mountain mining district in the central Great Basin region of the Basin and Range province of the western US. The deposit is at Copper Canyon in the southernmost portion of the district, about 19 km (12 miles) southwest of Battle Mountain (Fig. 1). Mining activity in the district began in 1866 (Roberts and Arnold, 1965). Base and precious metals were mined by underground methods intermittently until 1955. Placer gold was first discovered at Copper Canyon in 1912. Moderate scale dredging of alluvial gravel at the mouth of Copper Canyon was conducted by the Natomas Co. from 1944 to 1955 (Johnson, 1973). In 1967, Duval Corp. began large scale open-pit copper mining. (Sayers, Tippett, and Fields, 1968). Copper prices decreased and precious metals prices increased in the mid-1970s so Duval's exploration efforts at Copper Canyon shifted to precious metals. These exploration efforts resulted in the discovery of the Tomboy-Minnie deposits and shortly afterward, the Northeast Extension deposit at Copper Canyon (Blake, Wotruba, and Theodore, 1984; Fig. 2). The Fortitude deposit was discovered in late 1980, after Duval had proceeded with precious metals production at Copper Canyon. Converting the existing plant facilities at Copper Canyon to a precious metals recovery system was completed in 1978 (Jackson, 1982). Production from the Tomboy-Minnie deposits began in early 1979. Precious metals production moved to the Northeast Extension deposit in 1982. In late 1984, after three years of preproduction stripping and mining of the Upper Fortitude ore zone, production from the higher grade Lower Fortitude ore zone began. Currently, Battle Mountain Gold Co. operates the Fortitude deposit. Production from the Battle Mountain mining district before 1967 included 4.6 t (150,000 oz) of gold, 65.3 t (2.1 million oz) of silver, 13.6 kt (15,000 st) of copper, 4.5 kt (5,000 st) of lead, and 13.6 kt (15,000 st) of zinc (Roberts and Arnold, 1965). At the time of Duval's 1967 start-up, ore reserves at Copper Canyon included 12.6 Mt (14 million st) in the East ore body and another 3.6 Mt (4 million st) in the West ore body. Both had average grades of 0.79% copper, 16.1 g/t (0.47 oz per st) silver and 0.86 g/t (0.025 oz per st) gold (Blake et al., 1984). The Tomboy-Minnie deposits collectively contained 3.5 Mt (3.9 million st) of ore grading 3.09 g/t (0.09 oz per st) gold and 9.6 g/t (0.28 oz per st) silver. The Northeast Extension reserves were identified as 1.4 Mt (1.5 million st) of ore with an average grade of 2.9 g/t (0.085 oz per st) gold and 5.1 g/t (0.15 oz per st) silver. Initial total reserve estimates for the Fortitude deposit were 14.5 Mt (16 million st) grading 5.14 g/t (0.15 oz per st) gold and 29.8 g/t (0.87 oz per st) silver (Jackson, 1982). Therefore, before large scale mining, the Copper Canyon area contained a minimum of 102.6 t (3.3 million oz) of gold and 541 t (17.4 million oz) of silver as mineable reserves. The Fortitude deposit reserves as of Jan. 1, 1985 were 10.6 Mt (11 million st) grading at 5.28 g/t (0.154 oz per st) gold and 28.5 g/t (0.83 oz per st) silver. Currently, the company is producing 6.8 t (220,000 oz) of gold and significant amounts of

Jan 7, 1986

By Sir Robert Hadfield

SINCE the writer has been intimately connected with the development of manganese steel for many years, some remarks upon the early work with regard to the rolling and forging of this material might he of interest, and particularly as to the question of rails of special quality and of superior wearing qualities. In a paper on Iron Alloys, with Special Deference to Manganese Steel, read before this Institute at the International Engineering Congress in Chicago in 1893,1 the writer mentioned that, about 20 years ago, his firm in Sheffield produced railroad material, such as tires and axles, of forged and rolled manganese steel, thus showing that this product could be manipulated and worked up into the various required forms. The tires illustrated in the paper referred to above formed a particularly difficult class of work to produce in a special steel, but no special obstacles were met with Although there was no difficulty at that time in making and producing manganese steel in various forms, forged, rolled or pressed; nevertheless, on account of its higher expense, sufficient encouragement

Jan 2, 1914

By S. K. Kawatra, H. Haselhuhn

"In many mineral processing applications it is necessary to know the isoelectric points of the various minerals in the ore to properly perform flotation and other mineral separations. However, current zeta potential analysis technology used to determine the isoelectric point of a particle suspension cannot deliver the isoelectric point of each individual mineral, just an average of the mineral system. Traditionally, individual isoelectric points have been gathered by testing pure, synthetic minerals. These values can vary widely depending on the mineral synthesis procedure and the water quality used during the isoelectric point analysis. A more robust approach for determining the actual isoelectric point of a particular mineral in a mixture of minerals is needed. This paper details a method for determining the isoelectric point of an iron oxide mineral in a siliceous iron ore. The method uses a laser scattering particle size analyzer and a pH electrode to determine the pH at which the liberated mineral particles begin to flocculate as pH is decreased. The pH at which the average particle size rises dramatically is the isoelectric point of the mineral with the highest isoelectric point in the ore. This measurement technique was used on natural hematite, goethite and siderite ores as well as a synthetic mixture of pure silica and pure hematite. The results for synthetic hematite mixtures were comparable to literature values.IntroductionRecently there has been an increased interest in the processing of low-grade iron ores (Bolen, 2014; Carlson and Kawatra, 2008, 2011, 2013; Halt et al., 2014; Halt and Kawatra, 2014; Haselhuhn 2012, 2013; Haselhuhn et al., 2012a, 2012b; Kawatra and Halt, 2011; Liu et al., 2014; Manouchehri, 2014; Sandvik and Larsen, 2014; Semberg et al., 2014). The zeta potential of iron ore minerals is a very important and often overlooked aspect of iron ore mineral processing. Zeta potential is the charge at the shear plane of a surface in an aqueous environment (Carlson and Kawatra, 2013). Many ores consist of a variety of minerals differing in zeta potential. This property of a particle surface in water is responsible for either keeping the particles dispersed from each other or flocculated to each other. This is important during many operations such as flotation, desliming and thickening. Zeta potential is primarily a function of the pH of the water and the mineralogy of the particles, though certain ions within the water can also have a significant impact (Haselhuhn, 2012). Consider the following situation: silica (SiO2) particles and hematite (Fe2O3) particles are suspended in an aqueous solution. These two types of minerals have drastically different zeta potentials throughout the pH spectrum (Esumi et al., 1988). In alkaline conditions, both types of particles exhibit negative zeta potentials and will tend to disperse from each other. In acidic conditions (between pH of 3 and 6), silica has a negative zeta potential while hematite has a positive zeta potential, so these surfaces will tend to be attractive to each other and the particles will flocculate and settle. This phenomenon is visually described in Fig. 1. The pH at which the two minerals begin to attract to each other, or destabilize, is near hematite’s isoelectric point. The isoelectric point is defined as the pH at which the surface has a net zero zeta potential (Carlson and Kawatra, 2013)."

Jan 1, 2015

By Andrew R. Marchbank

The application of biologically assisted methods for metal extraction in the mining industry has a long history, but development of modern techniques has been relatively slow. There is, however, a resurgence of interest in the industry with the recent development of methods to successfully treat base metals concentrates. Two major pilot plant programs for treatment of chalcopyrite concentrates have been initiated in the past year and results from one are beginning to be released. This paper provides an overview of methods that have been applied in the past; from the earliest deliberate applications of the technology, through the copper and uranium heap and dump leach applications, the more highly engineered agitated tank leach gold applications and the developments currently taking place in the industry.

May 1, 2002