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By Clifford Britch

"Kaiser Resources Ltd. has completed feasibility studies resulting in approval of $277,500,000 for the construction of the Greenhills Surface Coal Mine, located 35 km north of Sparwood, B.C. The mine is designed to produce 1,800,000 tonnes per year of metallurgical coal over a 20-year period by conventional truck /shovel methods.The project is in the detailed design phase, with some preparation work being carried out on the site. The major construction activities are scheduled to commence in the spring of 1981. Mine development and pre-stripping will commence in July 1981, with first coal production scheduled for mid 1983.This paper will be a general description of the project and will include a discussion of the geology, reserves, mine planning, coal handling and processing, surface facilities and related topics.IntroductionProject DefinitionThe Greenhills Project is designed to produce, by conventional truck I shovel surface mining methods, 1,800,000 clean tonnes of metallurgical coal per year for twenty years . The production/ life constraints were chosen as being meaningful for planning and evaluation purposes after appraisal of the market potential and are not limited .by economic reserves. Substantial additional strippable reserves are available which would facilitate expansion of the proposed mine life and/ or capacity. In fact , all the facilities are designed so that they can be expanded in the future to accommodate a doubling of annual coal production, and certain portions of the coal processing facilities are being installed with an annual capacity of 3,600,000 tonnes. The mine will also produce approximately 2,400,000 clean tonnes of oxidized thermal coal, mostly in the first decade."

Jan 1, 1981

Large scale fluid extraction from the Latrobe Group in the Gippsland Basin is occurring as a result of offshore petroleum production and coal mine dewatering in the Latrobe Valley. Groundwater extraction in the Latrobe Valley exceeds recharge and is causing an expanding area of declining hydraulic head in the Latrobe Group, and regional land subsidence which is well documented. Recently the government water sector completed the installation of a network of Latrobe Group monitoring bores in the remainder of the onshore part of the Gippsland Basin. The bores situated along the coastal area show a steady decline in hydraulic head of about 1 m/year which can be attributed mainly to petroleum production.

Jan 1, 1992

By B M. Washington, R R. Lovel, G J. Sparrow, M J. Fisher-White

"Work in CSIRO with several Australian ores has shown that?by heating a high-phosphorus goethitic iron ore and then leaching with caustic or acid, the phosphorus levels in the ore can be lowered to a commercially acceptable value of around 0.08 weight per cent phosphorus (P). With increasing heating temperatures up to 1300¦C, the extent of phosphorus removal increased. This is considered to be associated with mineralogical changes in the iron oxide phases in the ore during the heating.Little phosphorus was removed from four high-phosphorus goethitic iron ores containing 0.12 - 0.16 weight per cent P in leaches with boiling water. However, leaches with sulfuric acid removed ten to 20 weight per cent of the phosphorus and leaching with sodium hydroxide removed even more phosphorus to produce leach products containing between 0.08 and 0.12 weight per cent P from three of the ores.After heating at 300¦C, leaches at a low pulp density of 0.5 weight per cent solids with 1 - 5 M NaOH and 0.1 - 1 M H2SO4 gave leach products with 0.06 - 0.09 weight per cent P. The lixiviant concentration required and extent of phosphorus removal depended on the ore. At a higher pulp density of 25 weight per cent solids, leach residues with between 0.07 - 0.14 weight per cent P were obtained.More phosphorus was removed from ores heated at 900¦C. Leaches at both the low and high pulp densities, with 1 - 5 M NaOH and 0.1 - 1 M H2SO4, gave leach products with 0.05 - 0.09 weight per cent P. However, the lowest phosphorus levels were obtained after heating the ores at 1200 or 1300¦C and leaching with 1 M NaOH or 0.1 M H2SO4. Leach residues with 0.02 - 0.06 weight per cent P were obtained, even from leaches at the higher pulp density of 25 weight per cent solids. Leaching tests on heated samples under heap leaching conditions at ambient temperatures have shown that phosphorus levels can be reduced to similar levels to that achieved in an agitation leach within ten days.CITATION:Fisher-White, M J, Lovel, R R, Sparrow, G J and Washington, B M, 2013. Effect of temperature on removal of phosphorus from Australian Goethitic iron ores, in Proceedings Iron Ore 2013, pp 373-380 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

By T. Mills, M. Evatz

Newmont’s Lone Tree Mine, located in northern Nevada, is an open pit gold operation supported by a standard truck and shovel mine production fleet (Caterpillar 785Bs and Hitachi EXs). Newmont continues to search for ways to decrease capital expenditure and lower operating cost, hence the implementation of the Ultra Light Body (ULW) concept. The fundamental idea associated with the ULWs is the increased payload potential. Economic justification comes from successful achievement of additional payload, within the truck’s rated GVW, without changing its loading and hauling characteristics. The increase in payload capacity allows equivalent tonnage movement using fewer trucks.

Jan 1, 2001

By D Morrison, P Goodman

The cycle time between adjacent stopes is critical to achieving the required production rate from 3500 Orebody; one of the Enterprise Mine orebodies. Current Mt Isa cemented fills require a 90-day cure period which had to be cut to 30 days while maintaining the required strength criteria. The Enterprise Mine Feasibility Study recommended that this be achieved by Cemented Slurry Rock Fill (CSRF), but that on-going research into paste fill be continued. The continuation of the research, in conjunction with overseas visits has recommended that the 3500 Orebody be filled with paste fill, with the potential to fill other orebodies. The paste fill will comprise cyclone and thickener dewatered copper tailings to which Portland Cement (PC) is added before vertical borehole delivery and undergound pipeline reticulation to individual stopes. The benefit is not only re-use of copper tailings but also much reduced binder usage.

Jan 1, 1998

By L. A. Cramer

Platinum metal was first utilized by the Inca civilization in South America where it was found in alluvial deposits along the San Juan and Atrato rivers that lie between the Andean mountain range in Colombia and the Pacific Ocean. The nuggets found in these rivers average between 77% and 86%platinum, 7%?8% iron and small amounts of copper, gold and other Platinum Group Metals (PGMs). The Incas often alloyed platinum with copper to increase its workability and fabricated various ornaments from it.

Jan 1, 2001

By R. H. Estrada-Ruiz

This work analyses a new approach to estimates bubble size distribution of froth surfaces using artificial neural networks (ANN).Also, the robustness of ANN to interpret images with illumination perturbations, produced by light problems or dirt attached to the window of the video camera is evaluated. The experimental work was carried out in a laboratory flotation column, instrumented with an image acquisition system. The images were processed making use of a perceptron model with a hidden layer, sigmoidal transfer function and unitary bias, and the ANN trained with a back propagation algorithm. The results of validation show that ANN are reliable for learning and producing generalized predictions of the froth mean bubble diameter and bubble size distribution, when the model is trained using a database that contains information on the illumination intensity. Keywords: image analysis, neural networks, bubble diameter, bubble size distribution.

Jan 1, 2009

By M. Syamujulu

In 1994, Nkana smelter commissioned a multimillion US Dollar Chilean Teniente converter (CT) in its operations in order to improve its smelting technological base. The main objective was to reduce the overall smelting energy costs and increase the sulphur capture for sulphuric acid production to support the company?s leaching operations at Nchanga tailings leach plant. A decade of CT operations have since ensued following its installation and start-up. This paper provides a technical and operational audit of the performance of the CT in the first 10 years of its operations at Nkana smelter.

Jan 1, 2005

By John C. Machisak

INJURY experience and employment data at all metal mines in the United States and Alaska are presented in this Information Circular for the calendar year 1958. This publication covers the metal mining industries, and contains statistical data on injuries and employment pertinent to metal mines, ore-dressing plants, and primary nonferrous smelters, refineries, and reduction plants. These operations report injury and employment information annually to the Bureau of Mines on a voluntary basis. The Bureau is authorized to make requests for such data, and the information submitted contributes immeasurably to the promotion of safety in the metal mining industry. Every effort has been made to make the annual canvass as complete as possible. This publication contains, along with production and development workers, the office employees at the mine location. The metal mines identified in this publication are classified according to the mineral ores mined; that is, copper, iron, lead-zinc, gold-silver lode, gold placer, and miscellaneous metals.

Jan 1, 1962

By M. Abdellatif

Keywords: Magnesium, pyrometallurgy, Mintek Extensive testwork on atmospheric thermal magnesium production has highlighted certain important fundamental and operational factors that have a major influence on magnesium extraction and its extraction rate, the condensation efficiency of magnesium, as well as on the quality of the condensed metal. The possible impact of these factors on the selection and preparation of raw materials, recipe choice, furnace electrical energy consumption, and on the refining requirements of the crude magnesium is discussed. Selected pilot plant operational and metallurgical data are also presented.

Jan 1, 2006

By Edwin W. Meyer

The augered cast-in-place piling technique was used to install deep foundations 20 feet into weathered sandstone. The project involved the upgrading of the Air Quality Control System (AQCS) for a 560 megawatt coal fired power plant located in Stillwater, Minnesota. The site subsurface consisted of 20 feet of overburden overlying weak sandstone. The overburden consisted of sands and gravels with cobbles and boulders, and variable thicknesses of peat. Ground-water at the site was approximately 8 feet deep. Two foot diameter piers were selected to facilitate installation through the cobbles and boulders in the over-burden and to penetrate the sandstone 22 feet. During installation of the first pier, the weak sandstone broke down to sand-sized material that could not be cleaned from the pier hole without the use of slurry to hold the material in suspension. However, the use of slurry would significantly impact both schedule and cost. Since the sandstone could be easily drilled with the pier equipment, it was believed that 2 foot diameter deep foundations could be installed as designed using augered cast-in-place piling techniques. A brief test drilling pro-gram was conducted to evaluate the viability of 2 foot diameter continuous flight augers penetrating 22 feet into the weak sandstone. The test drilling indicated that augered cast-in-place installation techniques should be successful. Following a short delay to mobilize the additional required equipment, construction commenced. Installation of the 2 foot diameter augered cast-in-place piling to a depth of 22 feet into the weak sandstone was successfully performed.

Jan 1, 2006

By Milton Camino Ramos

MARSA forma parte de los yacimientos vetiformes de la franja metalogenética aurífera del Batolito de Pataz, y políticamente se encuentra ubicado en la provincia de Pataz, La Libertad. En este sentido, el presente texto describe los resultados del programa de exploración 2010-2011, donde se dio el descubrimiento del primer cuerpo mineralizado del yacimiento, el cuerpo Esperanza, entre otras ocurrencias. Asimismo, se muestra que los cuerpos mineralizados de la franja metalogenética de Pataz, se formarian en estructuras de primer orden con evidencias de reactivacion, cuyo sistema estructural funciona como entrampamiento y en varias etapas de mineralizacion.

Sep 12, 2011

By Gabriel Esterhuizen

The room-and-pillar mining method is used extensively in underground limestone mines in the Eastern and Midwestern U.S. The rock mass is typically a near-horizontal, bedded deposit at relatively shallow depth. A survey of 34 mines was conducted in which data on roof spans, rock mass properties and support practices were collected as part of a National Institute for Occupational Safety and Health research project into ground control in limestone mines. The average room width is 13.5 m (44 ft), accommodating large mechanized equipment. The results show that the immediate roof beam can consist of many individual layers and can contain several near-vertical joints. In spite of the room widths, the roof is naturally stable in many mines. Observed roof instabilities can generally be related to excessive horizontal stresses or unfavorable geological structures that cause block fallout or beam failure of the bedded roof rocks. The survey showed that the rock mass quality of the limestone does not vary significantly throughout the Eastern and Midwestern U.S. Roof conditions and the need for support were found to be closely related to the thickness and competence of the first layer of limestone in the roof. The findings highlight the need to identify local geological structures and the roof beam characteristics so that support alternatives can be evaluated. Monitoring and observational technologies that are available for identifying potentially unstable roof are presented, which include: roof characterization using the Rock Fall Risk Index, automated logging of roofbolt hole drilling, microseismic monitoring and displacement monitoring.

Jan 1, 2007

By Mark S. Stagg, David E. Siskind

The mining industry occasionally blasts near pressurized transmission pipelines and has requested guidance of safe vibration levels and setback distances. the Bureau of Mines and the Indiana Department of Natural Resources cooperated with AMAX Coal Company on a study of coal mine overburden blasting. Five buried and pressurized 250-ft pipeline sections were specifically installed on the Minnehaha Mine highwall near Sullivan, IN for testing to failure. Four welded steel pipes ranging from 6- to 20- in diameter and one 8- PVC water supply pipe were monitored for vibration, strain, and internal pressure for a period of 6 months while production blasting advanced up to the pipeline field. In contrast to previous studies of small-scale blasting representing construction activities, these tests involved overburden blasts of up to 2100 lb per delay in 12-l/4-in diameter holes.

Jan 1, 1996

By Poul Hansen

"In mid-I986, a world-class mining enterprise was created in British Columbia, bringing together the copper mining operations of Lornex Mining Corporation and Cominco into a new entity, Highland Valley Copper. This paper will focus on the background and rationalef or the formation of this partnership, the resulting operating changes and the productivity and cost benefits expected.IntroductionSince the deterioration of the world copper industry in 1981, the future of the Lomex and Cominco mines in the Highland Valley of British Columbia looked increasingly precarious. Despite the expansion of the Lomex mill complex in 1981 to one of the world's most cost-efficient copper and molybdenum concentrators, persistent low metal prices had made the remainder of the ore in the Lomex orebody marginally economic and had significantly foreshortened the expected life of the operation.Cominco had acquired the Bethlehem Copper assets in order to consolidate its ownership of the Valley orebody, and was feeding the Bethlehem concentrator with ore from the Valley orebody. In spite of the high grade and free-milling character of the Valley ore, profitability of the Cominco operation was severely restricted because of the limited capacity of the Bethlehem concentrator and its location high on the hillside above the orebody."

Jan 1, 1987

By J. A. Stewart

"A common problem in the mining industry involves the overloading of established conveyor systems. This overloading can have a number of causes, such as new production technology or a changing market condition which increases demand. An overloaded system usually spells disaster and leads to clean-up problems, equipment failures and lost production. The Cape Breton Development Corporation's Lingan Mine had a serious overloading situation on its underground coal clearance system, and equipment was being damaged and production was being lost. It was rectified, in part, by increasing the coal clearance capacity through the use of booster conveyors. This paper discusses the problems which led to this situation, the booster concept and the installation of the first unit at the Lingan Mine.IntroductionThe Lingan Mine is a submarine coal mining operation with an annual production of 1.5 million tonnes from three advancing longwalls. Coal is conveyed 3400m up a 13% slope to the surface. Designed and installed in the early 1970s, the system's 820-tonne/hour design initially proved adequate; however, by the late 1970sthe system was handling surge loads of 1640 tonnes per hour due to the introduction of new face equipment.The coal clearance system consisted of three 485-Kw, two 210-Kw and one 105-Kw conveyors. Each conveyor was designed for 820 tonnes per hour using 1050-mm, Type 10 belt with a 35-degree troughing profile.There were many effects of the overloading on the system:• Loading beyond the physical capacity of the belt resulted in spillage along the conveyor, which increased operational and maintenance costs.• Overloading of the drive components resulted in premature equipment failures.• Spillage at the drive pulley often resulted in damage to the belt and pulley lagging and, under extreme conditions, the conveyor would require partial unloading before it would operate. This was a particular problem under wet conditions because the 485-Kw conveyor drives were fitted with single-drive drums."

Jan 1, 1982

By Chen Xia, G. Deschenes, Wan-Tai Yen

"Ammonium thiosulphate is a potential replacement for the cyanide in gold leaching. However, the high reagent consumption hinders its practical applications. In order to reduce the thiosulphate consumption, chemical additives were used. The gold leaching tests were conducted on a sulphide copper-bearing gold ore with 0.3M thiosulphate, 0.03M copper sulphate and 3M ammonium hydroxide solution at 44% solids and pH 10.2. Sulphate, sulphite, amino acid and chelating agents were initially used as additives. It was found that chelating agents (EDTA and NTA) could reduce the thiosulphate consumption from 30 kg/t to 17.8 kg/t with some increase of gold extraction. The improvement from other additives was not significant. Other approaches were also attempted to reduce thiosulphate consumption with some success. The thiosulphate consumption could be reduced to about 17 kg/t by reducing copper sulphate concentration fiom 0.03M to 0.0015M; adding 0.04M of EDTA; replacing copper sulphate with 0.001M of nickel sulphate. The increase of pulp density fkom 44% to 60% solids could reduce the thiosulphate consumption to 17 kg/t as well. Limiting the dissolved oxygen content by sparging nitrogen reduced the thiosulphate consumption further to 10 kg/t. It could be concluded that the regulations of copper and oxygen supply in the lixiviant were the keys to reduce the reagent consumption.INTRODUCTIONCyanidation, as the predominant gold leaching method, has been used in mining industry for more than a century. The low reagent requirement and the high performance are the reasons for its great success. However, cyanidation cannot deal with ores containing large amount of copper, nickel, arsenic and zinc minerals. This type of ore may consume more reagents and result in a poor gold extraction. On the other hand, the carbonaceous materials may preg-rob the gold from pregnant solution and make cyanidation a problem leaching method. Further more, the high toxicity of cyanide represents another challenge to modern gold mines. In the beginning of the year 2000, an environmental disaster caused by an accidental spill in Yugoslavia was broadcasted world-widely (Gurevitch 2000). Gold mine operators are facing higher social pressures and increasing costs on environmental protections. The usage of cyanide in mining industry has been banned in some countries and locations such as Czech Republic, Japan, Turkey, California, Colorado, Wisconsin and Montana States. Thus, a non-cyanide lixiviant is required."

Jan 1, 2002

By Manuel Gomez

Bituminous coals having maximum fluidity values in the 40- to 25,400-dial-divisions-per-minute (ddpm) range were carbonized in a continuous, entrained-bed reactor. Carbonization was conducted in an 8-inch diameter, vertical reactor at temperatures from 955° to 1,345° F and at air-to-coal ratios ranging from 4.63 to 16.12 scf/lb of moisture-and ash-free coal charged. Carbonization rates from about 118 to 360 lb/hr, as-carbonized basis, were examined. It was demonstrated that carbonization temperature and air-to-coal ratio must exceed minimum values for continuous operation. Within the operating conditions discussed, highly fluid bituminous coal was carbonized to produce tar and nonagglomerated char. The yield of tar ranged from 21.3 to 38.9 gallons per ton and the char yield ranged from 1,109 to 1,563 pounds per ton, all units expressed on a moisture-and ash-free basis. It was demonstrated that the maximum fluidity and the average size of the coal along with the air-to-coal ratio and the carbonization temperature act together to influence the carbonization product yields and certain properties of the products. Prediction equations were developed from the data for product yields and selected properties of the char, tar, and gas.

Jan 1, 1968

By Kenneth Gavin

The accurate prediction of the shaft capacity of drives piles in sand continues to present problems for foundation designers. Research in this field has been carried out for many years and has yielded a wide range of design approaches -some of which have been shown by recent experimental work to be based on questionable assumptions. This Paper reviews some of the current popular design methods for piles in sand in the context of recent information obtained from high quality field tests conducted by institutions such as Imperial College London. A database of driven piles in cohesionless soils has been compiled from well documented (and reliable) case histories that have been reported in the literature. The measured shaft capacities of the piles in this database are compared with predictions made by the various methods and an assessment is made of the predictive performance and reliability of each method

Jan 1, 1996

By M. Kumral

The efficiency of an ore-processing unit depends on the consistency of the characteristics of raw material entering the plant. When the mined ore is highly variable in quality, the only way to ensure consistency is to homogenize the ore prior to feeding to the processing plant. The homogenization can generally be achieved by the bed blending operation. Given that the stockpiling and reclamation processes are very expensive, it is necessary to design the process in such a way as to minimize variabilities of specified properties of raw material. In this paper, for alternative stacking types, optimal stockpile geometry is found in three stages: First, stockpile input is simulated by sequential Gaussian simulation, and then the variance reduction ratios (VRR) as a criterion of stockpile efficiency are calculated for various stockpile geometry scenarios by a stockpile simulator written in FORTRAN. Second, multiple regression analysis is performed to model the VRR by the use of stockpile length, the number of layers and stacker speed as the independent variables. Finally, the model is an optimization problem. Decision variables are the stockpile length, the number of layers, stacker speed and stacking type. The genetic algorithms (GA) are used to minimize the VRR. The approach was demonstrated on data from an iron orebody. The problem was to reduce fluctuations of iron, silica, alumina and lime contents in the stockpile output. The results showed that the approach could be used for the bed blending design efficiently. Keywords: bed blending design, multiple regression analysis, genetic algorithms, iron ore, content fluctuation

Jan 1, 2006