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By A. Romero

"Most underground mines require heating or cooling systems to meet health and safety standards in the working areas. In regions where daily or seasonal fluctuations of environmental conditions exist, these systems tend to be over-sized and operate at low average load factors, which increase both investment and operating costs. One possible approach to reduce the temperature variations of the fresh air supplied to the mine, and so reducing the investment in large heating or cooling systems, is to use rock as a thermal regenerator or thermal storage system. Packed rock beds act as thermal inertial units, damping the atmospheric temperature oscillations to provide a more constant fresh air temperature. However, although regenerators exist in some mines, and packed bed rocks have been studied in detail in solar thermal power applications, there is a lack of tools to design new and sufficiently large storage systems of this kind. This work fills the existing gap by presenting a methodology to design and evaluate the impact on the ventilation systems of these regenerators. The method is validated using existing rock bed systems found in the literature and illustrates possible designs for small and large size regenerators in both hot and cold climates.1 INTRODUCTION “ It is quite amazing how often answers to extremely complicated problems can be obtained in this way, by using greatly simplified models of the real situation for the purpose of calculation.” Whiller, A. in private communication to J. S. Stachulak (1989). Motivation Most underground mines located in regions with extreme climatic conditions require the thermal pre-treatment of the fresh ventilation airflows, whether for heating or cooling. Heating commonly is needed in sites with sub-zero temperatures and is usually effected by direct or indirect burning of natural gas or propane. Alternatively, cooling is required in areas with hot temperatures, or in mines with high underground thermal loads where the fresh air at ambient temperature cannot provide enough refrigeration. Thus, when the ambient air conditions cannot provide by itself enough thermal power to ensure safe underground working conditions, a heating or cooling plant is required, which involves a considerable equipment and operation cost."

Jan 1, 2019

By S. V. Ulianov, A. Y. Senchenko, E. M. Kurchevskaya

"A process sample of old copper-containing tailings (-2.0+0 mm) from an iron-ore plant was tested in TOMS institute in order to develop a process for copper concentrate production. The study of tailings material composition indicated that -0.2+0 mm fraction was efficient and contained 97.26% of copper. Copper is predominantly concentrated in chalcopyrite which is present both in free form and in intergrowth with ore and rock-forming minerals. Study of chalcopyrite using an optical microscope and Qemscan SEM determined that the most grains surface was covered with oxide films. Additionally, a lot of grains had micropores on their surfaces that indicated a partial mineral leaching. The oxide films did not make it possible to float copper minerals from head material. It was recommended to apply an attrition operation to restore chalcopyrite floatability. In order to increase selectivity a regrinding stage for a primary copper concentrate was introduced to obtain P80 26 um. Taking these recommendations into account a flotation flowsheet was developed to process the old Cu-containing tailings. The flowsheet allowed us to obtain a concentrate with 23.5% Cu content at 71.5% recovery.INTRODUCTIONStudy of material composition is an important prerequisite for development of a minerals processing technology. Mineralogical study contributes to search for an efficient technology to recover valuables and for prospective methods of minerals processing.The use of such up-to-date analysis methods and equipment as Qemscan and Mineral-C7 allows one to determine minerals size, liberation degree, and mineral associations. Mineralogical data can reveal reasons for hard processing of minerals prior to development of the process.The advanced process mineralogical methods used to study Cu-bearing pond tailings indicated several factors that influenced development of the process. The first one is a barren size fraction which is to be removed from the process in order to considerably reduce the costs connected with grinding. The second is the presence of films on chalcopyrite surface; in this context, a preliminary attrition of the material was recommended. The third factor is finely disseminated Cu minerals in sulphide and in the rock, which predetermined selection of particle size for flotation. All these data led to significant decrease in time required to develop the process technology. The developed process flowsheet based on the obtained mineralogical data is present in Figure 1."

Jan 1, 2016

By Stephen P. Quin

One of the most prospective undeveloped greenstone belts in Canada, the Hope Bay greenstone belt extends over 80km in a north-south direction and is between 7 and 20km wide. The belt lies on the arctic coast and comprises mafic meta-volcanic and meta-sedimentary rocks that are bounded by Archean granite intrusives and gneisses. The greenstone package has been deformed during multiple events and is transected by major northsouth trending shear zones that appear to exert a significant control on the localization of mineralization. The Hope Bay belt contains a number of significant gold deposits hosted in a typical Archean greenstone belt and which occur in several geologic settings, including shear-hosted, lode-style and breccia-associated. Combined, the deposits host a measured and indicated resource of 1.6 million oz of gold averaging 15.7 g/t plus an inferred resource of 2.7 million oz of gold averaging 12.3 g/t. All deposits remain open to expansion. Exploration in the period 2000-2002 by Miramar includes more than 110,000m of drilling to more fully define the nature of the known deposits, while regional exploration has focused on identifying and evaluating new targets. In carrying out its programs, Miramar uses a number of techniques to improve its chances of exploration success. Activity in 2003 includes 43,000m of drilling focused on expanding resources in the Madrid area, testing the Boston deposit at depth and continuing the hunt for more discoveries. Meanwhile Miramar is advancing one of the deposits towards a production decision, based on a low capital, high return development scenario.

May 1, 2003

By R. White, C. Doyle, B. Martin, C. Díaz, J. Hrepic, G. Cuthbert

"A program initiated to improve the service life of the magnesite-chrome refractory in the top blown rotary converters at INCO’s Copper Cliff Nickel Refinery identified slag corrosion as a major wear mechanism. Target slag compositions were established to produce a non corrosive slag. A plant program to achieve this target slag by the addition of controlled amounts of do lime and MgO was instituted. Addition to rates of the slag modifiers was optimized based on the slag composition assays achieved and the requirement to address corrosion effects on the high alumina content ladle refractory. A60% gain in TBRC refractory performance was eventually achieved and has been maintained yielding a net annual savings of $700,000. A significant gain in available converter operating time has also been realized.IntroductionINCO’s Copper Cliff Nickel Refinery utilizes two top blown rotary converters (Fig. 1) to smelt a blend of semi-refined nickel concentrates, producing a granulated nickel matte suitable for the adjacent carbonyl based nickel refining operation. The converter, or TBRC, operation is a batch process composed of feed charging and melting, oxidation of the excess sulphur in the matte, and reduction of residual nickel oxide. The final matte, containing about 78% nickel, is tapped at 1600°C from the converter and is granulated in a sluice equipped with high pressure water jets."

Jan 1, 1997

By J. Allaire, P. N. Muteb

The Meadowbank mine is operated by Agnico-Eagle Mines Ltd. and is located in the Kivalliq District of Nunavut in northern Canada. The process plant operates presently at a throughput of 11,500 tpd with gold grade varying between 2.9 and 3.6 g/t. The ore supplied to the process plant come from two different open pits and it expected to come from a new open pit early 2014. The ore is mid-hard materials so the feed size ore have a bigger impact on the throughput. The average gold recovery is 94.0% with about 15-30% of gold recover from the gravity circuit. The Meadowbank mine commissioning started in February 2010 utilizing a flowsheet consisting of one stage of crushing followed by primary and secondary grinding. The commercial production began in March 2010. During the commissioning several difficulties were observed to reach the designed throughput of 8,500 tpd. It was evaluated to insert a second stage of crushing. The insertion of the secondary crushing have two objectives, first to reach easily the process plant design throughput as well as to evaluated the mill capacity increasing. The process plant team evaluated various expansion alternatives eventually deciding to install a secondary crusher, increased the SAG and Ball mills motors capacities and maintain the rest of the process unchanged. Early 2011, with the improvement in the gold markets, decision was made to install the secondary crusher between the primary crusher and the primary grinding. The process plant continued to operate on the initial design until early June 2011. This paper presents the SAG mill throughput increasing from the designed tonnage of 8,500 tpd to 2012 targeted tonnage of 11,500 tpd, secondary crusher commissioning and the difficulties to operate in Canada Artic (Nunavut).

Jan 1, 2013

By Simon D. Handelsman

?Artisanalmining ?Application of human energy (e.g. pick & shovel) directly to produce minerals ?majority no technical skills and lack adequate working tools ?Small-scale mining ?Application of modern technology for limited scale mineral production Significance of ArtisanalMining ?15 million people in artisanaland small scale gold mining ?55 developing countries ?Use of mercury damages health of 100 million people

May 1, 2006

By P. Kordi, L. Egorova, R. J. Bowell, D. Gurevich, V. Sedov, A. Mikhailov, D. A. Holwell

ABSTRACT Reconnaissance exploration in the Kuene region Namibia has included geological mapping, soil and stream sediment geochemistry, and ground geophysical methods. Copper mineralization occurs within a well-defined belt with north-northwest to south-southeast trend, parallel to the regional strike of the sedimentary rocks. The occurrences vary in size but are generally stratabound. The mineralization is centred on the transition from calcareous shale and interbedded limestone, through carbonate rocks, to quartzite. Geochemical anomalies have been identified correlating to induced polarization (IP) and magnetic anomalies. The anomalies identify one potential metallogenic zone more than 17 km in strike length.

Jan 1, 2012

By Ian D. Pirie

Presentation outline ?Project location ?Geology ?Resources and reserves ?Results of the FEED study ?Social and environmental considerations ?Project path forward

May 1, 2010

By Peter Radziszewski

"Typical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5% to 20% ). Cornminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and linerkharge wear (50 %). Developing the ability of predicting mill wear will allow a more accurate estimation of wear costs. However, total media wear in a given ball mill grinding process is a product of three recognized wear mechanisms those being abrasion, corrosion and impact. Further, the contribution to total media wear of each of these wear mechanisms has not been well established. In this paper, a overview of abrasive, corrosive and impact wear mechanisms is made from which an empirical wear prediction procedure is proposed. Further, future directions in wear prediction research is outlined. INTRODUCTIONTypical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5 % to 20 %). Comminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and lined-charge wear (50 %). With respect to wear, world wide steel consumption as grinding media is estimated at over 600 000 tomes per annum (Malghan, 1982). As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs.Charge media wear prediction is actually estimated using the Bond abrasion test. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2000

By R. F. Pilgrim

"This first chapter of the SPOC Manual 1s intended as a general introduction to all the other chapters in which the mineral process evaluation and optimization methodology is presented. This chapter also provides a brief description of available software for mainframe and m1crocomputers and provides information on its distribution by the Technology Information Division. CANMETRESUMECe premier chapitre du manuel SPOC est un guide général qui présente les autres chapitres ou l'on traite de la méthodologie d'évaluation et d'optimisation des procédés miméralurgiques. Le lecteur y trouvera aussi une description sommaire des programmes de calcul développés pour ordinateurs centraux et pour micro-ordinateurs ainsi que la façon de les obtenir de la Division de l'information technologique du CANMETComputer simulation of mineral processes has slowly emerged as an established branch of activity in A&D centres and in several teaching institutions after decades of slow progress. Substantial efforts to develop application software based on years of research on process modelling are being pursued in many countries ( 1 .2,3,4.5). The late BO's should see a flurry of success stones about significant leaps of productivity achieved in several plants through systematic process evaluation and computer simulation studies.In Canada. the SPOC project (Simulated Processrng of Ore and Coal) was initiated at CAN MET in 1980 with the obJective of developing and transferrrng to the mineral rndustry a computer-based methodology for process evaluation and optimization. The proiect deliverables include:- an engineering manual covering most aspects of off-line computer simulation:- a bank of FORTRAN software accessible through data communication lines for user familiarization and distributed at a nominal cost in source form by CANMETThe yearly progress reports of the project contain detailed information on the project elements since 1980 16). The context of SPOC can be summed up by quoting Gordon Agar""s 1975 statement during a workshop on research needs in the mineral industry: ""What is needed is a joint effort between the academic community and one or more enliglitened industrial concerns. The effort will have to be an iterative one with data collection followed by testing the model modification of the model (s), more testing prediction, modification. data collection and so on (7). A few parallel in1- t1atives had been reported in papers or research contract reports (8, 9) but there was an obvious need for CANMET to bridge the gap between the R&D and the production sectors by undertaking a project that would document past applications and induce a new generation of initiatives through the consol1dat1on and transfer of suitable manuals and software. The effort took shape as the CANMET SPOC pro1ect in which methodology transfer was to play a key role (10)."

Jan 1, 1986

By A. Grimod

Rockfall drapery systems are commonly used as a simple, fast and cheap measure against the rock falling trajectories affecting very steep slopes. The system basically consists of a steel mesh, which is held on the top edge with a longitudinal cable fixed by means of a suitable number of anchors. The effect of this kind of intervention is to control the falls of the rocks, which are driven toward the bottom with slow velocity, or are stabilized on place. It can be used on any kind of slope to protect sensitive target in the mining industry, roads and railways, and inhabited areas. The falling blocks, typically smaller than 0.6 - 1.0 m in diameter, pile up into a trench (or into a ?pocket of mesh?) at the bottom. In comparison to other types of rockfall interventions, the simple drapery is cheaper, and its maintenance is easier, but on the other hand it cannot be considered the remedy for shallow instabilities, because it can only mitigate the effect of the falling trajectories. The design of simple draperies requires the analysis of several factors like the slope features (height, gradient, morphology), the geological and dynamic features (nature of the ground or rock, type of instability, erosion problem, blocks size), the environmental condition (presence of vegetation, aesthetical problem), the installation problem (accesses to the slope, safety for the workers, safety for the surrounding areas) and finally the performance required (temporary or permanent intervention, required maintenance, cost). Finally the most problematic design-step is the choice of the suitable mesh, the top longitudinal cable, and the top anchors type. Because of the highly variable nature of rockfall behaviour these structures cannot be standardized, but they have to be analyzed and designed for each case. Maccaferri has developed a new software application (MacRO 2) has a practical tool to define the mesh and the related supporting structure composed by up-slope cable and anchors. The software, based on the approach proposed by Muhunthan B. et al. (2005), allows designers to verify the top longitudinal cable, the anchors, and the mesh drapery, and to plan for maintenance procedures. Even if the method seems quite simple and rough, it is effective and lets the designer correctly foresee the materials and the geometrical features to be used on the interventions. This paper analyses the condition for the simple drapery installation, the main steps for the calculations, and it will present a case study in a U.S. mine. Nevertheless, even if the software allows a quick and simple calculation approach, onsite observations are always recommended in order to achieve a good design, with the ultimate goal of protecting property and human lives.

Aug 1, 2013

By B. M. Monaghan

"The developments of the 1950s and 1960s have seen Canada become an international source of iron ore, trading in all geographical areas and hemispheres. Canada's production will significantly increase in the next decade. However, although the outlook is bright, iron ore producers cannot be complacent. Instead, we must weigh the factors that have brought about the development of the Canadian iron ore industry and examine the probable changes for the next decade that will affect its growth.Looking ahead in the 1970s, competition will be severe. Pressure on prices will continue as new sources of supply bid for long-term contracts and as steel producers continue to develop mines around the world. Rising costs will make these pressures more severe. We must balance the growth of the world's steel makers with the emergence of new supplies of iron ore. The iron ore industry is now part of a new era in which demand and supply are intermeshed, and the developments around the world are acting on the Canadian iron ore industry's future."

Jan 1, 1971

By Y. P. Malhotra, JE. C. Carter

"Syncrude's overburden removal requirement ranges from 20 million to 30 million bank cubic metres per year to meet tarsand demands for the production of 129,000 barrels of synthetic crude oil per day. The economic tar-sand lies beneath a variable thickness of overburden materials, composed of silts, clays, some sands of lacustrine origin, wind-deposited sand dunes, and out-wash sands and gravels.For the removal of overburden, Syncrude commissioned its own mobile equipment fleet in January 1981. This fleet consists of two Demag 241 hydraulic shovels, ten Titan 33-15B trucks and support equipment. Prior to this, all overburden removal was being carried out by contractors.This paper presents an overview of overburden removal at Syncrude, with particular reference to the performance of its own fleet.IntroductionThe Syncrude Tar Sands Plant was officially opened on September 15, 1978. After an expenditure of approximately $2.26 billion, the second and the largest commercial tarsands plant was on stream toward a permit production rate of 129,400 barrels of synthetic crude oil per day. The Syncrude Plant, a joint venture of six oil companies, the federal government and the Alberta government, is a major step forward toward the goal of Canadian energy self-sufficiency.The Mildred Lake operation of Syncrude Canada Ltd. Is located 40 km north of Fort McMurray and 480 km northeast of Edmonton in the Province o f Alberta.Synthetic crude oil leaving the plant is a blend of two-third s gas oil and one-third naphtha. The two streams combine in a pipeline and flow to Edmonton. The product takes six days to make the journey, arriving at four locations in the city Imperial's Strathcona Refinery, Gulf's Edmonton Refinery, the Interprovincial pipeline terminal and the Trans-Mountain pipeline terminal. Each Syncrude owner receives a share of oil equal; to its percentage of ownership in the project. At the present time, the oil is being utilized by Canadian refineries from Vancouver to Montreal."

Jan 1, 1983

By J. D. Clucas, D. B. Dreisinger

"The Starved Acid Leach Technology (SALT) has been developed by Search Minerals to extract nickel and cobalt VALUES (from nickel laterite ores (saprolites) and Caron plant residues. The technology consists of sulfuric acid atmospheric leaching of the ores or residues followed by neutralization and recovery of nickel and cobalt. The acid addition is carefully controlled to maximize nickel and cobalt extraction while minimizing overall consumption of acid by various gangue reactions. The final recovery processes can be tailored for maximum economic benefit. The use of resin-in-pulp processing is advocated as the most cost effective method for initial recovery of nickel and cobalt from lower grade materials. INTRODUCTIONCommercial processes for laterite ore treatment vary according to laterite type and local conditions. The two general categories of laterite ore are saprolites and limonites.Saprolites are generally treated by pyrometallurgical processes for reduction of nickel to matte or ferro nickel alloys. The high cost of drying, chemical reduction and electric furnace smelting requires that the saprolite grade for economic processing is of the order of 1.8% nickel or higher. There are of course large amounts of saprolite that fall below this “cut-off” grade that are not currently economic.Limonites can be treated by the Caron Process which involves drying of the ore, a reduction roast of the whole ore to reduce nickel and cobalt (and some iron) to the metallic state followed by ammonium – ammonium carbonate solution to leach nickel and cobalt. Nickel and cobalt recovery are low (70-85% for Ni and 20-50% for Co). The Caron plant tailings therefore contain high levels of unrecovered nickel and cobalt. Relatively few plants have used the Caron Process and new installations are not anticipated due to the high energy cost of drying and reduction roasting. Nevertheless, plants in Brazil (Niquelandia), Australia (Yabulu) and Cuba continue to operate.The high pressure acid leach (HPAL) process was pioneered at Moa Bay in Cuba in the 1960’s in the treatment of limonites. In the 1990’s the HPAL process was touted to revolutionize the nickel industry with low capital and operating costs for new installations. Murrin Murrin, Cawse and Bulong in Australia were started and in each case the technical difficulties of operating at extreme temperature and pressure conditions and the challenges in “ramping up” a complex hydrometallurgical facility conspired to cause these plants to fail economically. At present, Cawse and Bulong are closed and Murrin Murrin continues to operate after undergoing a financial restructuring."

Jan 1, 2012

By R. W. Watson, H. D. Peer

"This paper outlines the history of steel linings in the Autogenous Mills, the factors influencing the decision to convert to rubber lining, and the operating and cost experience with the rubber liners to date. INTRODUCTIONSherman Mine is an open pit iron ore mining, concentrating and pelletizing operation located at Temagami, Ontario. The operation employs 392 hourly employees and 81 salaried employees and has an annual capacity of over one million L.T. of high quality iron ore pellets produced from low-grade magnetite ore.The Sherman Mine is a joint venture; 90% owned by Dominion Foundries and Steel Limited (Dofasco) of Hamilton, Ontario, and 10% owned by the Tetapaga Mining Company, a wholly-owned subsidiary of the Cleveland Cliffs Iron Company of Cleveland, Ohio. The operation and management of the mine is by Cliffs of Canada, Limited. Development of the property was started in 1965, with pellet plant production beginning on March 17, 1968. At the present time, the orebodies contain sufficient reserves to sustain operations at current production levels for 30 years. Mining was originally carried out in three orebodies, the North, South, and West pits. The North Pit has been exhausted and mining operations ceased in the spring of 1980. A fourth pit, the East Pit, will be developed during the early 1980's."

Jan 1, 1981

By Pablo González, Rodolfo Espinosa, Jose-Luis Noyola, Sergio Cossio

"Minera Penmont, at La Herradura, is a joint venture between Industrias Peñoles, S.A. de C.V., and Newmont Mining Corporation. The mine is located in the northern Mexican state of Sonora. La Herradura consists of an open pit mine with two-stage crushing and heap leaching facility, the gold is recovered from the pregnant solution using the Merrill-Crowe process; production began in July 1998. La Herradura produces about 130,000 gold ounces per year, and treatment capacity is 22,000 tonnes/day. Ore mineralization occurs in a shear zone with quartz and native gold. Gold distribution is quite heterogeneous due to the coarse visible gold, in some coarser particles ranging from 100 to 600 microns.Sampling and assaying protocols where established during exploration and prior to starting up operations, based on nomographs developed from early heterogeneity tests. But, as the coarse gold problem became noticeable, it was necessary to develop new sampling protocols to have an acceptable relative standard deviation to control the fundamental error within ± 15%.An experimental program was developed and conducted on a selected ore block made of 195 blastholes; regular ore grade control procedures were applied to both sets of data to evaluate the difference in results between the protocol established prior to starting up operations (“existing”) and the new protocol (“leaching”). In the “existing” protocol, subsampling is done through riffle splitters, and coning and quartering. Fire assaying is used to estimate the gold content, on a 50 g sample. In the “leaching” protocol a 12,000 g sample was bottle rolled cyanide leached, and the leach residue was assayed on a 50 g sample using fire assaying, and a weighted average total gold content was calculated. Currently, a new alternative method is being evaluated, where sub-sampling is done using rotary splitters, and fire assay on a 50 g sample; in this “rotary sampling” protocol and for comparison purposes, gold assays were also repeated using the “leaching” protocol.Test results were compared in three groups < 0.30 g/t (below cut-off), 0.30 to 0.8 g/t, and >0.8 g/t. Overall. Results show that the protocol established prior to starting up operations underestimates the gold assays. Detailed technical results, alternative options, and their implications are discussed in the paper."

Jan 1, 2003

By J. Andrew Cornier, Mike Cooper

"Upon successful completion of the AG/SAG Mill project (Cooper et al., 2001) the Brunswick Mine concentrator was configured with single line processing through the AG/SAG mill, followed by two parallel ball mill and flotation circuits. The Stage Grinding project was undertaken to utilize a pre-flotation stage on the AG/SAG mill primary cyclone overflow and simplify the flowsheet by consolidating the two ball mill and flotation circuits, resulting in single line processing fi-om AG/SAG mill to final concentrate. Concurrently, the reclaim water supply capacity was increased to enable flotation cleaning at lower percent solids.Benefits from the Stage Grinding project include revenue enhancement from increased zinc grade and recovery combined with additional lead recovery. Further benefits included a reduction in the number of process pumps, reduction in loss associated with spillage to sumps and process simplification through efficient use of equipment layout.The Stage Grinding project was implemented in six phases over a period of seven months in 2000. Each phase was commissioned during scheduled concentrator shopdays with minimal additional downtime required. One year after project approval the Stage Grinding project was completed at a total cost of $1 1.8 million.INTRODUCTIONThe Noranda Inc. Brunswick Mine, located near Bathurst, New Brunswick, is one of the world’s largest leadzinc mines. Milling of ore began in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore fkom the No. 6 open pit located several kilometers away. In 1980 the pit was exhausted and the concentrator rationalized to accept a design feed rate of 10,500 tomes per day from the No. 12 mine. In excess of 100 million tomes have been milled to date, representing over three-quarters of the known reserve. The concentrator employs differential froth flotation to produce lead, zinc, copper and leauzinc bulk concentrates. The ore is massive sulphide, with approximately 20% valuable minerals, 60% pyrite and 20% silicates."

Jan 1, 2002

Peñasquito is Mexico's largest gold producer, consisting of two open pits - Peñasco and Chile Colorado - containing gold, silver, lead and zinc. Over the next three years, mining activities in the pit are expected to be focused on lower grade ore in the upper parts of the Peñasco pit while stripping is emphasized to ensure an economically optimal pit shell design to maximize the net asset value of the operation. By 2019, Peñasquito’s gold production is expected to benefit from an improvement in mined grades as it recommences mining higher grades at the bottom of the Peñasco pit and significantly enhanced metallurgical recoveries with the planned completion of the Pyrite Leach Project ("PLP").

Jan 1, 2017

By M. Yamine, R. J. Ring, D. J. Waters

"An experimental program has been carried out by the Australian Atomic Energy Commission to compare the effectiveness of Caro&apos;s acid as an alternative oxidant to pyrolusite in the leaching of uranium ores from the Alligator Rivers region of northern Australia. Batch leaching tests on Ranger, Nabarlek, Koongarra and Jabiluka ores showed that uranium extraction was unaffected by choice of oxidant, but that Caro&apos;s acid reduced acid requirements by 15 - 25%. Lime requirements for neutralization of tailings liquors were reduced by 20 - 30%.The batch studies were extended to a four-stage continuous leaching circuit with a capacity of 400 gore h-I. Comparative leaching tests confirmed the reagent savings obtained in the batch tests. Following these encouraging laboratory results, lnterox Chemicals Pty Ltd carried out plant trials using Caro&apos;s acid. Reagent savings in the mill were similar to, or better than, those obtained in laboratory test work. The trials also established that Caro&apos;s acid has a number of operational advantages over pyrolusite.IntroductionUranium is most commonly extracted from its ores by leaching in sulphuric acid. Simple hexavalent uranium oxides and compounds are readily soluble but uranium in the tetravalent form is insoluble and must be oxidized for dissolution to occur. The most commonly used oxidants have been sodium chlorate and manganese dioxide. They act indirectly by maintaining iron as the ferric ion which oxidizes uranium by a surface electro- chemical reaction. An Fe3+ concentration of about 1 - 2 g L-I is usually considered adequate for the effective dissolution of uraninite(1)."

Jan 1, 1984

By André van Wageningen

Long term mine scheduling is one of the most difficult optimization problems. The problem is said to be NP-complete in that the best possible solution cannot be found any quicker than checking all possible permutations and combinations. In current mining operations, it is important to report changes in long term plans as soon and as accurately as possible. Where geological data can change rapidly, such as in narrow vein deposits, the need to quickly identify and optimize different long term alternatives becomes a difficult task and can lead to many permutations of the long term mine plan. The question arises, is the new plan optimal or not? Can the company objectives be met while adhering to operational constraints? Although a number of optimization techniques have been used in the past, many include significant simplifications or fail to produce acceptable results within the required timeframe. Genetic Algorithms (GAs) are among the most promising optimization techniques within operations research. They are based on natural selection and evolution of solutions to complex problems. They have been known to quickly converge within a few percent of the optimal solution by examining only a fraction of the possible permutations. The GA itself contains very little problem specific information and, as such, in this study, has been linked to a discrete-event simulation (DES). The DES is able to evaluate each sequence proposed by the GA and determine the effectiveness and value of each schedule. The DES feeds information regarding the feasibility and quality of the solutions back to the GA during the optimization process. In this paper, GAs and DES are described and the coupling of the two methods is outlined. The GADES module is then used to ?optimize? a narrow vein theoretical case study and the results discussed in a broader mining context.

Oct 1, 2004