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By K. J. R. Ford

"Significant advances in mining and metallurgical operations have occurred over the last two decades, to define ore bodies via geological mapping, mineralogical analyses, geostatistical models and resource reserve modeling. Metallurgical process measurement, controls and modeling have also significantly progressed, with advanced instrumentation, control software, data capture systems and simulation tools. This progress together with development of the science of geometallurgy has allowed geologists, mineralogists, miners and metallurgists to start working together to define geometallurgical ore units and link these to mine planning and projected metallurgical performance. However, application of full geometallurgical programs into existing mine-processing operations or into new “greenfield” projects has been slow. This is mainly due to higher costs, limited dedicated technical resources and little or no corporate leadership. Additionally, life-of-mine cash flow models, the “dollars”, have been largely driven by ore tonnage rates, pay-metal grades and fixed pay-metal percentage recoveries, with few or no geometallurgical parameters included.Here defined, the “electrons” represent intrinsic ore characteristics embracing chemical composition, mineralogy, gangue components and their associated fundamental metallurgical processing phenomena. With its ever expanding treasure chest of measurement and modeling tools, geometallurgy paints a much clearer picture of the “electrons” for the geologist and metallurgist. But to fully understand and realize the real net value of ores over the life of a mine, a comprehensive link of “electrons to the dollars” is essential. Intrinsic characteristics of geometallurgical ore units are by definition linked to their metallurgical processing, but importantly need also to be related to operating costs, capital requirements and revenue.This paper discusses this important linkage under topics of ore geology and mineralogy, metallurgical processing, waste disposal, environmental issues, metal refining, project engineering and lifeof- mine models. Topics are illustrated by multi-metal references and by case examples from gold ore heap leaching and manganese ore smelting projects. 1"

Jan 1, 2013

By Marvin E. Lane

Viburnum, located in the foothills of the Ozarks in south-east Missouri, is the site of a modern lead mining and milling operation. Two mines are currently producing a total of 5,000 tons per day and a third mine is in the final development stage, with full production scheduled for late this year (1964). These mines will supply the tonnage required to keep the Viburnum mill operating at a full capacity of 7,000 tons per day for many years. The efficient underground mining operation is completely trackless. Drill jumbos are used to break the ore in a room and pillar method of mining. Transloaders, or St. Joe shovels, and trucks do the loading and hauling. The Viburnum mill has a world-wide reputation as a most efficient operation. This is due to the extensive use of automatic process controls, and labour saving materials handling facilities. Introduction THE St. Joseph Lead Co., has operated lead mines in southeast Missouri for 100 years. These operations have been confined largely to the area called the

Jan 1, 1964

By Gennaro Delisle

"Les Services TMG (Niobec) currently treat a niobium ore in the Chicoutimi area of Quebec. The mineral pyrochlore is present in a carbonatite deposit which assays 5% to 7% apatite.The ore is ground to minus 65 mesh and deslimed at 10 micrometres. Part of the carbonates are floated with a fatty acid emulsion and rejected to the tailings as a first step to ensure the proper conditions for the flotation of the pyrochlore concentrate. The carbonate concentrate is 65% minus 44 micrometres and is composed of dolomite, calcite, 22% apatite and a minor amount of silicates.A laboratory investigation was undertaken to determine the feasibility of extracting apatite from the carbonate concentrate. This led to the development of an economic flowsheet to produce high-grade apatite concentrate. The apatite is depressed with phosphoric acid, and the pH is adjusted at 4.8-5.0 with sulphuric acid. An additional amount of fatty acid is added to permit efficient flotation of the carbonates. 1n a second stage, the silicates left behind with the apatite concentrate can be removed by depressing the apatite with AIC3 while floating the silicates with a secondary amine.The results of the laboratory tests indicate the possibility of producing an apatite concentrate assaying 38.3% P2O5 (83.7% BPL) with a recovery of 66.2%. Pilot-plant tests undertaken at the concentrator yielded a concentrate of 34% P205 and recovery of 76%."

Jan 1, 1981

By Glenn D. Lyle

As part of an initiative to improve value in mining, CEMI is developing a handbook on best practice in mine design. Various formats were examined, and of these, the wiki was selected as being most suitable. Being web-based, its global reach and capability of being continuously updated allows the wiki to evolve as new knowledge is generated, and to capture the diversity of experience of practitioners in the mining industry. While wikis have many advantages in terms of organizing content and searching, their ability to be edited by users is often seen as a negative characteristic for quality control. To ensure reliability of information content, author attribution, specific authorship models, and a peer review process have been implemented, in addition to the standard version tracking capability of wikis. The first phase of the project concentrates on topics related to underground hard rock mine design. The scope will be expanded to cover other aspects of mining as the project matures. Strategic partnerships with various organizations, such as the CIM, and CANMET, are being formed to provide a wide range of seed material to stimulate growth of the site. www.minedesignwiki.org

May 1, 2011

By D Marshall

The sub-level blasthole benching method of mining has been proved to be economical at Willroy Mines. How-ever due to the nature and attitude of the wall rock, this method is susceptible to serious hanging-wall dilution of the ore. Attempts to solve the dilution problem by various means were not satisfactory until long Stressteel bars were employed as rock bolts. The adoption of this de-vice does not require any special skill, and has become routine in the under-ground operation. Since the inception of this method, the dilution of the ore has been greatly reduced and the ov1!r-all costs of controlling this dilution are much lower. INTRODUCTION T HIS is a report describing the successful use of long Stressteel bars as rock bolts in preventing l1anging-wall failure and consequent dilution in open stopes. Willroy Mines Limited is located (Figure 1) near the town of Manitouwadge, Ontario, approximately l1alf-way between Sault Ste. Marie and Port Arthur.

Jan 1, 1963

By J. T. Furey

"A pilot plant study was conducted to evaluate the performance of column flotation in the recovery of gold from a gold/tellurides bearing sulphide ore, and to compare the performance of column flotation to conventional ""mechanically agitated"" flotation.Tests were conducted in a 60 liter glass column with ground flotation feed split from a concurrently operating conventional pilot plant, thus permitting direct comparison between the two processes. During the test program, the effects of wash water rate, gas holdup, froth depth, residence time and flow bias were examined in detail.The ore had an average ""as tested"" feed grade of about 5 g/t gold and 1 % sulphur, with the gold occurring primarily as inclusions of montbragite (Au2Te3) in pyrite, with minor amounts of native gold. Conventional flotation using 60 minutes residence time across four roughing stages recovered about 97% of the gold into concentrates grading about 50 g/t Au. A single stage of column flotation roughing using 25 minutes residence time matched the conventional circuit recovery of 97%, but produced grades of+ 150 g/t Au. INTRODUCTIONColumn flotation is the most significant innovation that has occurred in the mineral processing industry in recent years and has become an important alternative to mechanical flotation. Column flotation has been most widely accepted and applied in cleaning service where it has frequently proven to be superior to mechanically agitated flotation. Roughing in columns has not been ignored, but early testing indicated that columns may not be capable of achieving adequate recovery and, to date, columns have not been seen as serious competitors for mechanical roughers."

Jan 1, 1990

By Yum Xiaoquan, Turgut Yalcin

"A laboratory study was carried out to determine the potential role and effect of ultrasonic irradiation in the flotation of a Sudbury copper-nickel sulphide ore. The use of ultrasound was tested separately at the pulp preconditioning, collector conditioning and flotation stages. While no significant benefits were evident in the latter two cases, ultrasonic preconditioning of pulp improved the grade-recovery performance and had the prospect of reducing reagent consumption rates. Tests were conducted on both the normal and oxidized ore samples, and several variables were investigated, including ultrasound frequency (25 kHz and 45 kHz), ultrasonic power (up to 1.6 W/cm2), duration of ultrasonic treatment (up to 60 minutes), and the intensity of pulp agitation employed during the application of ultrasound. Ultrasound frequency of 25 kHz and duration of 15 minutes yielded satisfactory results. Changes were found to occur in the temperature, pH and Eh of the pulp during ultrasonic processing. The favorable effects of ultrasound observed in this study were attributed primarily to its surface cleaning capability.INTRODUCTIONThe use of ultrasound for the surface cleaning of objects is well known. Ore flotation is based essentially on the surface properties of minerals, and hence stands to benefit from ultrasonic treatment which may help restore mineral surface characteristics that might have been lost or modified through oxidation, etc. Expected outcomes of this method of treatment include improved flotation kinetics, higher recovery, better selectivity, and savings in reagent consumption, as has been demonstrated by various researchers (Aldrich and Feng 1999, Celik 1989, Ozkan 2002, Letmathe et al. 2002, Nicol et al. 1986)."

Jan 1, 2005

By Paul F. Grundy

The ever changing conditions in the coal-mining industry and the increasing competition of other fuels such as oil, natural gas, wood refuse, etc., have made it imperative that the coal operator be in a position to supply a constantly uniform and clean product. The old idea of putting on the market "anything so long as it's black" no longer holds good. Today's users of fuel have to consider the amount of heat per dollar they obtain from the fuel they buy. Burning of coal has become largely mechanical, at least so far as industrial coal is concerned, and industrial coal is the kind the bituminous operator is for the most part called upon to supply. When we get into things mechanical, the matter of precision has a mighty bearing on efficiency. Precision in coal means standardization of quality, or uniformity of preparation in size and also in ash and moisture content. The trend of coal mining is ever towards more and more mechanization. Mechanization means coal-cutting machinery, loading machinery, conveying machinery, etc. Although such methods cut the cost, the cleanliness of coal so mined and loaded is not controllable to the same extent as when each operation is by hand labour. The operator has therefore the problem of mining his coal as cheaply as possible, but at the same time preparing a product which will to the utmost satisfy his customer. The answer is a mechanical cleaning plant. Before installing a cleaning plant, however, the operator must take into consideration the cleansability of the coal from his particular mine. This is accomplished by taking representative samples of the seam or seams to be treated, and subjecting these to screen tests, followed by float and sink tests on each size product. From these tests, it can be readily found to what degree of cleanliness the several sizes can be prepared advantageously, and also what will be the percentage recovery of cleaned coal. The accompanying tables give the results of representative tests on Michel coals of minus-2-inch size. All coal over 2-inc.h was intended to be hand cleaned. The coal-cleaning plants in general use at the present time come under two broad classifications, namely, wet and dry.

Jan 1, 1934

By H. Deveci, E. Y. Yazici, A. D. Bas

The leach solutions of waste of printed circuit boards (WPCB) may contain a wide range of base and precious metals with different concentrations due to heterogeneity of these wastes in composition. Therefore, selection of a suitable metal recovery method is of practical importance for downstream processing of pregnant leach solutions of WPCB. Sulphuric acid leaching of WPCB in the presence of a suitable oxidant is capable of dissolving base metals as well as precious metals such as silver. In this study, cementation of silver from sulphuric acid leach solutions of WPCB using metallic copper, aluminum and zinc was studied. The WPCB sample was assayed to have a silver content of 694 g/t. A synthetic leach solution containing 20 mg/L Ag in sulphuric acid solution (0.51 M H2SO4) was prepared based on the silver concentration in pregnant leach solutions of oxidative sulphuric acid leaching tests. Total sulphate concentration ([SO4 2-]T) was kept constant using sodium sulphate at 1 M. Metallic powders of copper (-850 +600 µm), aluminum (-425 +45 µm) and zinc (-45 µm) were used in the cementation tests. The effects of metal type (Cu, Al or Zn), dosage of metal powders (above the stoichiometric ratio by up to 2000-fold) and temperature (20-80°C) on the rate and extent of silver extraction from the solution was investigated. The effect of presence of copper on the cementation performance was also tested. Under suitable conditions high silver recoveries (>90%) was achieved using metallic copper and zinc while aluminum yielded the lowest silver recovery (i.e. 40% Ag). The kinetics of cementation by zinc was comparatively higher than that of other metals. Increasing the temperature was observed to significantly enhance both the rate and extent of silver cementation. SEM-EDS studies were performed to characterise the metallic powders after cementation.

Jan 1, 2016

By Pierre St-Julien, Réjean Hebert, M. Racine, Roch Gaudreau, Pierre Lacoste, Michel Rocheleau

"Les unités volcano-sédimentaires de la ceinture de l'Abitibi à l’est de Val-d' Or ont été regroupées en quatre domaines litho tectoniques. Du nord vers le sud, nous reconnaissons les domaines d'Assup, de Garden lsland de Val-d' Or, de Trivio et de Villebon. Deux de ces domaines (Garden Island et Trivio) correspondent à des couloirs de déformation, d’envergure régionale et d’orientation approximative E-W. Les trois autres domaines sont caractérises par des volcanites defond océanique de type MORB (tholéiites et komatiites) qui sont en position inferieure; ces volcanites sont surmontées par des sédiments marins de type turbidite (domaine de Ville bon) ou évoluent dans le temps, soit vers un volcanisme bimodal d'affinités tholéiitique et calco-alcaline ( domaine d'Assup), soit vers un volcanisme d'arc insulaire (domaine de Val-d'Or). Il n'existe aucune certitude de corrélation litho stratigraphique entre les différents domaines et l'on doit considérer la possibilité que ceux-ci représentent l'accolage tectonique de terranes de nature et d'origine différente.Les principales caractéristiques structurales de la région sont regroupées en quatre étapes de déformation dont trois sont archéennes et la quatrième attribuée à l'orogenèse grenvillienne. Au moins cinq facies métamorphiques ont été reconnus: I) un facies régional faible propre à la province du Supérieur, 2) un facies de cornéennes a amphibole en périphérie du pluton de Pershing, 3) un facies hydrothermal dans la zone tectonique de Cadillac, 4) un facies régional moyen dans la zone de bordure de la province de Grenville, et 5) un facies régional élevé a l'intérieur de la zone structurale du front de Grenville.Le front de Grenville est défini dans la région par l’apparition subite des ortho gneiss au faciès métamorphique élevé et par la disparition du grain tectonique E-W. Le contact brusque entre les deux provinces pourrait correspondre à une zone de chevauchement de la province de Grenville sur la province du Supérieur. La région étudiée est l’hôte d’un gite majeur de sulfures massifs et de plusieurs gites et indices aurifères; elle présente des cibles favorables pour l’exploration minière dans le contexte tectonométa-morphique de l'influence grenvillienne. L'intérêt de la Formation de Val-d' Or pour les métaux de base et le potentiel aurifère des couloirs de déformation d'orientation approximative E-W, sont bien connus. Les structures grenvilliennes NE-SW, qui se développent vers l'est, représentent aussi des cibles intéressantes, en particulier dans les secteurs ou elles intersectent des cisaillements E-W kénoriens."

Jan 1, 1990

By E. Uzun

"In this study, H2SO4 and HCl leaching behaviour of an ore sample containing 1.87% of Ni, obtained from the Karaçam (Eskisehir-Turkey) lateritic deposit that started operation recently was investigated. Acid concentration, temperature and reaction time were investigated as variables by comparative leaching tests and the optimum leaching conditions were selected on teh basis of Ni, Fe, Co and Mg behaviour. For H2SO4 leaching, optimum leach parameters were 95°C, a 12 normal H2SO4 concentration and 90 minutes of reaction time, and under these conditions Ni and Fe dissolutions were 97.6% and 95.4% respectively. On the other hand, 70°C, an 8 normal HCl concentration and a reaction time of 90 minutes were found to be the optimum parameters for HCl leaching, and Ni and Fe dissolutions found were 97.1% and 97.8%, respectively.INTRODUCTION Today, nickel and various nickel compounds are obtained from sulphide and lateritic nickel ores (Moskalyk & Alfantazi, 2002). Even though about 72% of the world's nickel reserves constitute lateritic ores, only about 42% of nickel is produced from these ores. The decrease in production from sulphide deposits, on the other hand, and increases in the use of nickel in the iron and steel, and battery production sectors, and especially the production of nickel cast irons in recent years, have accelerated the importance of making use of lateritic deposits (Luo et. al., 2010; Chang et. al., 2010). In the beneficiation of sulphide nickel ores, mostly pyrometallurgical methods are used. On the other hand, about 39% of lateritic ores are beneficiated by pyrometallurgical, and 61% are beneficiated by hydrometallurgical methods (USGS, 2009). In laterite beneficiation, the ore properties are of significance; that is to say, limonitic ores are eneficiated by high pressure acid leaching or via the Caron process, nontronitic ores are beneficiated by high pressure acid leaching or heat treatment (smelting), and ores containing serpentine by pyrometallurgical methods (Berezowsky, 2004; McDonald & Whittington, 2008a; Dreisinger et. al., 2015)."

Jan 1, 2017

By Graham A. Karklin

"The Echo Bay Mines silver property is located 875 air miles north of Edmonton, Alberta. The mine is on the east shore of Great Bear Lake at Port Radium which is twenty six miles south of the artic circle, in the North West Territories.HISTORYThe area was first surveyed by J. Mackintosh Bell of the Geological Survey of Canada in the early nineteen hundreds. In May 1930, Gilbert LaBine staked claims at Port Radium called ""Cobalt"" and ""Cobalt I"". What he had discovered was an extremely rich vein of pitch-blend. These claims were the birth of mining in this area. The first ore shipment of pitch-blend was hand-picked and flown out of Port Radium in 1931. LaBine's discovery of pitch-blend became of world-wide importance with the outbreak of World War II. In 1939 the splitting of the uranium atom brought this area into world focus. It is generally believed that it was ore from this area that was used in the production of the devastating bombs which were dropped on Japan.It was at Port Radium that the leaching process for uranium extraction was pioneered. The resultant methods have been modified and adopted for use around the world. The mine was originally owned by Eldorado Gold Mines Limited. Echo Bay Mines leased the property in 1964, subsequently purchased it, and have been producing silver and copper every since."

Jan 1, 1981

By C. A. Morais, L. V. Resende

"The technological development observed in recent years has stimulated the application of rare earth (RE) elements. Amongst their various applications it is well worth mentioning their use in electronics, superconductors, nuclear medicine, as well as in special inks - as those used in radar invisible airplanes. TV and computer monitor screens present as coating a powder containing some rare earth elements, mainly Europium and Yttrium. The recovery of the RE elements from electronic scrap is extremely important as the economic and environmental issues are concerned. This paper describes the study of the Eu and Y recovery from e-scraps. The first step of the process was the leaching of the powder obtained by manually scraping the inside surface of the screen. After the optimization of the leaching, the dissolution of Eu and Y was higher than 98 w/t% and a pregnant sulfuric liquor containing 11.9 g L-1 Y and 0.72 g L-1 Eu was obtained. The second step was the separation of the metals found in the liquor by the solvent extraction technique. The separation of Eu and Y was carried out in a counter current circuit compound of 3 extraction stages, 7 scrubbing stages and 15 stripping stages. DEHPA was used as the extractant, whereas a diluted sulfuric acid solution was applied as scrubbing agent and a hydrochloric acid solution at high concentration was used as the stripping agent.INTRODUCTIONThe rare earths (RE) are a moderately abundant group of 17 elements composed of lanthanides, scandium and yttrium. Amongst the RE, promethium (Pm) is extremely rare, however several of the RE elements are not very rare and occur dispersed in a variety of forms (Greenwood & Earnshaw, 1984; Tyler, 2004).The innumerable applications of the RE elements are based on their particular properties, mainly their spectroscopic and magnetic properties (Martins & Isolani, 2005). The technological development observed in recent years has stimulated the application of rare earth (RE) elements. They have been used in agriculture, electronics, superconductors, nuclear medicine, automobile industry, special inks – as those used in radar invisible airplanes, X-ray screens, high intensity mercury vapour light bulbs, neutron scintilators, charged-particle detectors and optical memory reading systems (Maestro & Huguenin, 1995; Nagaosa & Binghua, 1997; Sánchez et al., 1999)."

Jan 1, 2012

By P. Lausch, Z. X. He, J. F. Archibald

"Documented underground fill failures have resulted because of the lack of adequate knowledge concerning fill strength and other physical parameters, as well as to a lack of consistent quality of fill materials being used by operators. On the basis of literature review and correspondence with backfill operators at 18 mines throughout the province of Ontario, non-systematic quality control procedures have been shown to exist, both within different levels of single operations and between separate mining operations, which act to provide inferior engineered fill materials and erratic application response in situ.Assessment of existing procedures set by mine operators for determining backfill physical parameters, and recommendation of optimum fill material testing and manufacturing guidelines are presented in this paper.A series of standard fill characterization test procedures are recommended to provide substantial, though not comprehensive, fill behaviour review. The purpose of these recommendations is to propose industry standards for backfill quality review. Data acquired could be utilized for fill design functions and as a means of comparing fill quality parameters between sites in a regular and more systematic fashion than is presently done."

Jan 1, 1993

By J. M. Newell, G. R. Peatfield

"The Red-Chris porphyry Cu-Au deposit is located between Ealue and Kluea lakes, southeast of the village of Iskut on the Stewart Cassiar highway in northwestern British Columbia. It was explored by Texasgulf Inc., in a joint venture with Great Plains Development Company of Canada, Ltd. and Silver Standard Mines Ltd., during the period 1973 to 1980. Exploration resumed in 1994, following acquisition of the interests originally held by Texasgulf and Great Plains Development by American Bullion Minerals Ltd.Prospectors were first drawn to the area by prominent limonitic stain zones exposed in creek gullies in the western part of the property. These zones are developed on the alteration halo at the western periphery of the deposit. The potential ore zones do not crop out and were first exposed in a bulldozer trench excavated in 1971. Subsequent exploration was by percussion and diamond drilling, guided to some extent by induced polarization surveys. In 1976, estimated resources contained within two crudely designed open pits were 41 million tonnes with an average grade of 0.56% Cu and 0.34 g/t Au, with a cutoff grade of0.25% Cu. New resource estimates based on a major drilling program completed by American Bullion in 1994 were released as this paper goes to press. These estimates were calculated using various copper cutoff grades; at the 0.4% Cu cutoff, the ""in-situ geological reserve"" is given as 100 million tonnes grading 0.58% Cu and 0.46 g! t Au (American Bullion Minerals Ltd., 1995).The deposit consists mainly of pyrite, chalcopyrite and increasing bornite at depth, in zones of quartz-vein stockwork hosted by an elongate body of altered monzonite porphyry of latest Triassic or earliest Jurassic age. The stock intruded volcanic and sedimentary rocks of probable Stuhini Group affinity and is in faulted contact with sedimentary rocks of the Jurassic Bowser Lake Group. Gypsum stockworks are developed in the higher levels of the subvolcanic porphyry system. Late-stage quartz-carbonate veins carry small amounts of galena and sphalerite. Complex faulting within the deposit makes drill-hole correlations difficult. Alteration assemblages include an early potassic phase, almost completely destroyed by later phyllic alteration, and surrounded by a propylitic assemblage dominated by ankeritic carbonates in altered sedimentary rocks.The deposit has characteristics of bath the alkalic and calc-alkalic classes of porphyry copper deposits. The development of quartzvein stockwork suggests a calc-alkalic affinity, but the hostrocks contain little or no free quartz, molybdenum is present in trace amounts only and gold values are significant. Copper grades vary between wide extremes with unusually high grades in some sections of the deposit. The weight of evidence available in the 1970s led us to place the Red-Chris deposit firmly in the alkalic family. The quartz-sericite alteration assemblage was attributed to breakdown of potassium feldspar in a near-surface alteration environment in which seawater may have had a significant role. These interpretalions are retained here, but preliminary conclusions from the most recent work suggest that they may require some modification."

Jan 1, 1995

By S. Banisi, H. R. Iran-Manesh, M. R. Shayestefar, H. Shekarchyan

"Flotation circuit of the Sarcheshmeh copper mine processes 41,000 t/d of ore at 0.9% Cu. Chalcopyrite and chalcocite are the main copper bearing minerals which are concentrated in three stages of roughing, cleaning and recleaning. Due to major changes in the reagent addition regime, make-up balls combination, and throughput to the primary mills an increase of 2.6% (from 87.2% to 89.8%) in the sulfide copper recovery was obtained during a period of six years. In order to investigate a rather large variation of monthly recovery (i.e., from 89.0 to 91.2%) of the plant after implementation of the changes, a study was undertaken to trace mineralogically possible causes of such a behavior. During a five-month period, monthly composite samples of feed and tailings of the flotation circuit were taken. At the first stage, the polished sample preparation method and the mineral counting procedure were optimized through a detailed study. Liberation degrees and types of locking of valuable minerals in both streams were determined in three size intervals of +75, +38 and –38 microns. It was found that for both recovery levels (i.e., 89% and 91.2%), there was no significant difference in the liberation degrees of valuable minerals in the flotation feed; the average liberation degree was 57.1% with a standard deviation of 3.1%. On the other hand, in the case of high recovery level (i.e., 91.2%) there was a reduction of 15% in the amount of free valuable minerals in the tailings stream. The highest reduction was found for the -38 micron size range in particular for chalcocite. It was concluded that compared to the grinding circuit the change in the flotation circuit had the significant effect on the plant recovery increase.INTRODUCTIONThe Sarcheshmeh copper ore body contains 1 billion tonnes averaging 0.90 % copper and 0.03 % molybdenum. It is located southeast of Iran and currently processes 41,000 t/d of ore. The concentrator has been in operation since 1982. Last year, the concentrator produced 130,000 tonnes of copper and 4,000 tonnes of molybednite concentrate. An expansion to the plant which will increase the throughput by 50% is under construction and will be effective in 2003.After three stages of crushing, the ore with a F80 of 12.7 mm is fed to 8 parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of the first stage of flotation is reground and the tailings constitute the main component of the plant final tailings. A concentrate with an average grade of 30.5% Cu is obtained after cleaning and re-cleaning stages. At the present time, the average sulfide copper recovery of the plant is 89.8% which varies depending on the operating conditions and ore type."

Jan 1, 2003

By Euler De Souza

This paper presents a study conducted to assess sillmat behaviour during undercut mining and to provide a comprehensive understanding of the effects of arching and sillmat thickness on stability. The design study investigated the effects of such parameters as stope width and height, closure strains, orebody geometry and inclination, wall roughness, fill properties and cure time on the stability of cemented sillmat during undercut mining. Arching effects have been quantified based on centrifuge modeling studies for the different mining conditions. Observations have indicated that the degree of arching is a function of stope geometry, wall roughness, fill cohesion, and wall closure. Arching effects increased with decreasing stope width, increasing backfill cohesion and frictional resistance, increasing wall roughness and with wall closure.

May 1, 2002

By D. du Plessis, R. Stobbs

The Canadian Clean Power Coalition (CCPC) was created in 2001 to protect and enhance Canada’s vast coal and other carbon-based resource wealth, and to ensure that environmental public policy decisions recognize that these resources are a Canadian asset, not an environmental liability. To achieve its mandate, CCPC has undertaken to demonstrate technologies for use in new coal-fired power plants to control air emissions including carbon dioxide (CO2), nitrogen oxides (NOx), sulphur dioxides (SOx), mercury (Hg), and particulates (PM). The first phase of the demonstration project assessed three existing technologies for coal-fired power plants that reduce all emissions, including CO2, to levels equivalent to or below that of gas-fired power plants. The technologies considered in this phase included integrated gasification combined cycle (IGCC), ‘oxyfuel’ combustion, and amine scrubbing of pulverized coal. These technologies were evaluated with three coal types: Eastern bituminous, western sub-bituminous, and lignite coals. This paper presents the studies, conclusions concerning the capital and operating costs of these technologies, the cost of electricity produced (with CO2 removed), and the cost per tonne of CO2 avoided. It also discusses the potential for using captured CO2 for enhanced oil recovery, and possible technology enhancements for improving the competitiveness of gasification of low-rank western Canadian coals. The second ongoing phase of CCPC’s work includes the design and construction of a commercial-scale plant to demonstrate the feasibility of integrated gas combined cycle (IGCC) and/or polygeneration technologies and the potential for using the CO2 generated for enhanced recovery of conventional oil.

Jan 1, 2006

By Edith Lafontaine

Modern underground mines operate fleets of mobile production equipment including drill jumbos, long-hole drills, load-haul-dump machines, trucks and bolters. Their availability has a substantial impact onto production cost and thus onto mine profitability. Through appropriate collection of pertinent data and information, use of adequate performance measures and their analyses, mine operators can identify the most promising and cost effective opportunities to get more from their equipment and to reduce the overall cost. This paper presents results of the study aimed to review and analyse means and measures used to quantify equipment availability and its contributing factors (reliability, maintainability and maintenance support) in several Canadian underground mines. It overviews current reporting procedures, sources, quality and quantity of the data and information collected as well as examines performance measures and their definitions used by the surveyed mines. Problems concerning data acquisition as well as meaningfulness and usefulness of the above mentioned metrics are also discussed. The paper concludes with some suggestions and recommendations concerning several aspects of equipment availability assessment. It is believed that the latter may become a more useful tool for mine operators in their efforts to reduce production cost than it seems to be the case at present.

May 1, 2007

By J. W. Thomson, H. J. Lloyd, R. J. Taylor

"THE HOLLINGER MINE has been producing gold and by-product silver almost continuously for forty-eight years. The value of current production is $10,000,000 annually from one million tons of ore. Its management •has, particularly over the past decade, increasingly . concerned itself with the problem of profitable operation in a time of steadily rising costs and a fixed price of gold. At one period in the life of the mine, before World War II, production resulted in yearly profit margins that were stable and relatively thick. In winning gold from a series of inconsistent vertical ore lenses, operating management was much better protected than it is today against a general rise in cost or dip in grade of ore. The tissue-like thinness of present-day margins demand more careful handling. A sharp rise in costs, if protracted, could easily rip the tissue to shreds and close down the business. The counter-measure of raising grade and gutting the mine would bring on the end perhaps more rapidly, without having contributed much to the recovery of our national re-sources, nor to community welfare. The alternative is to keep costs under control."

Jan 1, 1960