Search Documents

By H. L. Martin

A joint federal-provincial approach in gathering basic information from industry on mine production, concentrator operations and ore reserves, initiated during 1978, is a major step toward the goal of achieving better government-industry communication through a common data base. Information from questionnaires sent out by provincial governments is shared with the federal government; correct interpretation is ensured by mine visits conducted jointly by provincial and federal representatives. Initial focus is on nonferrous metal mines. The federal Department of Energy, Mines and Resources (EMR) is responsible for a national analysis that is aimed at the early detection of worrisome developments-and their reasons-in the flow from resources to reserves and from reserves to supply of these metals. The heart of the national analysis is the annual construction of a "monitoring curve", showing each metal's aggregated mine production as it would decrease over time if no new reserves and production capacity were to be added from the year of analysis. For a given metal, a comparison of monitoring curves from year to year can reveal early signs of supply stagnation. In anyone year, a monitoring curve in conjunction with foreseen needs for Canadian mine output will show the production gap that would have to be filled from reserves and production capacity yet to be developed. The production gaps for copper, nickel, zinc and lead, the four leading base metals mined in Canada, are shown for the 20-year period ahead, in the context of production achieved during the past 20 years. It is shown that reserve levels commensurate with likely future production levels would require that average annual reserve additions during 1978-98 exceed the average annual additions during the preceding 20 years. Reserve levels of all four metals have declined in recent years. A review of significant discoveries in the last five years suggests that this fall in reserves is not the result of a lack of exploration successes, but of stagnation in development for various economic reasons.

Jan 1, 1979

By A. Philipp, A. S. Middleton, G. A. Harron, A. B. Durham

"Basal till sampling techniques were used to search for gold mineralization in the Timmins, Ontario area. During 1978 and 1979, approximately 90 overburden sampling drill holes were completed in Murphy and Hoyle Townships to test specific geophysically defined target areas. A conventional tricone dual tube drill system, and a unique retractable tricone/rock coring system were used to various advantages.First pass till sampling at 800 to 1200 ft intervals perpendicular to ice transport direction yielded anomalous gold values in till and carbonatized quartz pyrite-rich bedrock chips, and lateritic soil. Subsequent induced polarization-resistivity surveys defined the areal extent of a pyritic carbonate alteration zone. A second pass till sampling/limited rock coring program evaluated the gold mineralization potential of the alteration zone. Diamond drilling has substantiated the till and bedrock gold anomalies. Further evaluation of this mineralization is currently in progress. IntroductionThis paper describes an integrated geochemical/geophysical gold mineralization discovery case history in Hoyle Township. This discovery is on a property jointly owned (I983) by Canamax Resources Inc. and Du Pont of Canada Exploration Limited. Exploration of the property began in 1978 by Rosario Resources Canada Ltd. and in 1979 a joint venture agreement was signed with Du Pont of Canada Exploration Limited. Rosario Resources Canada Ltd. was taken over by Amax Minerals Exploration in 1981, which became Canamax Resources Inc. in 1982.The work described in this paper deals with a gold mineralization discovery located three-quarters of a mile northwest of the gold deposit currently being readied for underground exploration and development. Further drilling is required to investigate the economic potential of the gold mineralization herein described."

Jan 1, 1987

By H. E. Urtubia, C. W. Bailey

"BHP-Billiton’s Minera Escondida Limitada is one of the world’s largest copper producers. It treats a combined 88 million tons of copper ore per year through its two concentrators: Los Colorados and Laguna Seca. Over the past three years, metallurgical sampling has been performed at the Laguna Seca concentrator to calibrate both the JKSimMet and JKSimFloat grinding and flotation simulators. In the past, the calibrated simulators have been used to predict the effect of changes in operating conditions in the grinding and flotation circuits separately. To assist in the development of future processing options, predictions were required of the effect of changing grinding circuit conditions on flotation performance. This was achieved by using an integrated grinding and flotation simulation. The grinding circuit performance was predicted under different operating conditions (such as tonnage rate and product grind size) using JKSimMet, and these results were then used in JKSimFloat to predict the flotation performance for the new grinding circuit product. This combination of the simulators allows both the effects of residence time and grind size on flotation performance to be investigated.This paper presents the first combined grinding and simulation modelling and simulation case study using JKSimMet and JKSimFloat. The modelling and simulation methodologies for the grinding and flotation circuits are discussed along with the circuit simulations that were performed. The simulator predictions are presented in conjunction with independent laboratory results of the new operating conditions. The ability to predict flotation circuit performance based on changes in grinding circuit performance is also discussed. This methodology provides improved predictions of flotation circuit performance when grinding circuit changes (including grinding capacity increases) are made.INTRODUCTIONOver the past decade, the grinding and flotation research groups at the Julius Krutschnitt Minerals Research Centre - JKMRC (Brisbane, Australia), University of Cape Town (Cape Town, South Africa) and McGill University (Montreal, Canada) have been developing a methodology for optimising grinding and flotation plants using actual plant data together with semi-empirical sub-process models. The project is being conducted as part of the Australian Minerals Industry Research Association (AMIRA) P9 Project, titled “The Optimisation of Mineral Processes by Modelling and Simulation”. The project first commenced in 1962 and is currently in its “Nth” phase."

Jan 1, 2007

By Andre Vien, Jean-François Dupont

"Over the past several years, Svedala has developed a continuous charge monitor (CCM) for tumbling mills. It consists of a strain gauge embedded into a rubber lifter and associated electronics. This paper describes the installation and use of such a device on a SAG mill at the Holt-McDermott Mine. Comparison to measured static volumetric load and bearing pressure and inclusion of the measurement into a control strategy and its potential for condition monitoring is discussed. INTRODUCTIONWhile the measurement of load is very important in all kinds of grinding applications, the biggest benefits are expected from AG and SAG milling. This is what motivated Holt-McDermott to take a closer look at this problem, with a view to optimizing SAG mill performance.In many cases, the SAG mill control problem can be stated as maximizing throughput while avoiding a volumetric overload, with the other constraints (e.g. power, grind size, upstream/downstream equipment constraints, etc.) playing a less important role. This is certainly true at Holt-McDermott where the SAG mill is the production bottleneck about 90% of the time, and anything that can be done to increase SAG mill throughput will have a significant impact on operating profit.Although the goal of maximizing throughput while avoiding volumetric overload is simple to state, it has proven to be much more difficult to achieve. The mineral processing industry has expended a great deal of time and effort to devise control strategies in the pursuit of this objective. While much progress has been made, there is still room for considerable improvement. Across the industry it is probably fair to say that overloads occur relatively infrequently, but the real losses accrue as a result of the tendency to run rather conservatively."

Jan 1, 2001

By K. G. Wheeland, J. E. Dutrizac, M. Wasserlauf, A. A. Condy

Numerous processes for the treatment of thiosalts in base metal milling effluents have been examined in the past decade. Flowsheets for thirteen processes are outlined, and cost estimates are presented for a base case of a 720 mi/h effluent flow containing 1400 mg/L thiosalts. The processes are ranked on the basis of Numerous processes for the treatment of thiosalts in base metal milling effluents have been examined in the past decade. Flowsheets for thirteen processes are outlined, and cost estimates are presented for a base case of a 720 mi/h effluent flow containing 1400 mg/L thiosalts. The processes are ranked on the basis of

Jan 1, 1988

By James A. Brown, Marius Botha, Christopher A. Fleming

"The technical and economic feasibility or re-processing hundreds of millions of tons of tailings in a number of historic gold tailings dams in the Witwatersrand area of South Africa is being investigated. The objective is to recover uranium and residual gold from the tailings, as well as to oxidize the sulphides to eliminate long term environmental liabilities associated with acid mine drainage. The tailings from the re-processing plant will be re-deposited in a new area far from urbanization, allowing for the rehabilitation of the real estate currently occupied by these historic tailings.The first step in the process involves a mild-regrind and mild acid leach of the tailings (pH 2.5 to 3.0) to dissolve up to 40% of the uranium. This uranium is recovered directly from the pulp by resin in pulp (RIP) with a strong base resin. The RIP tails are then treated by flotation to recover a rougher pyrite concentrate, and flotation tailings that are discarded. The rougher concentrate is cleaned to produce a final pyrite concentrate and cleaner scavenger tailings, and the pyrite concentrate is oxidized in an autoclave to leach uranium and generate acid. The autoclave discharge is combined with the cleaner scavenger tails, and the acid and ferric ions generated in the autoclave are used to leach more uranium and oxidize more sulphide in the cleaner tails. The product from the concentrate leach circuit is processed by resin in pulp for uranium recovery, followed by neutralization, cyanidation and carbon in pulp for gold recovery.The rougher float tails may be further treated to recover gold that was not recovered in flotation, but which had been liberated from the old dumps by natural oxidation of some of the pyrite over many decades. The float tails would be neutralized and treated by cyanidation and CIL to recover this liberated gold.All parts of the flowsheet have been extensively tested in the laboratory, and the critical elements of flotation, pressure oxidation and resin in pulp have been demonstrated in pilot plants. The results of the hydrometallurgical investigations are presented here."

Jan 1, 2010

By T. Hadley, E. Catovic

"Northern Minerals Limited’s (ASX: NTU) flagship project is the Browns Range Rare Earth Project (the Project) located in Western Australia, consisting of a number of deposits with high value, heavy rare earth elements, in particular yttrium, terbium, and dysprosium. The deposits are dominated by xenotime mineralisation, which occurs within hydrothermal silicified and hematitic breccias. The Company delivered an initial JORC compliant resource estimate for the Project at the end of 2012 of 10,500 t of contained total rare earth oxides (TREO), which was increased to 28,084t of contained TREO in October 2013 and further enlarged to 47,997 t of contained TREO in February 2014. Concurrent with the growth of the Project resource base, significant progress has been made in developing the metallurgical process for the beneficiation and extraction of rare earth elements (REE’s) from the xenotime mineralisation. Development of the beneficiation circuit evaluated several process options and selected whole-ore flotation and magnetic separation, followed by flotation (hybrid circuit) as the two preferred flowsheets to pilot, following encouraging results on the bench of 83% and 86% recovery to a 20% TREO mineral concentrate respectively. Several downstream hydrometallurgical extraction flowsheets were evaluated. However, the following two preferred process flowsheets were ultimately selected for further detailed investigation during the prefeasibility study: 1. Sulphation bake, water leach, purification, oxalate precipitation, and/or calcination to produce a mixed RE oxide; 2. Sulphation bake, water leach, phosphate precipitation, caustic conversion, HCl leach, and oxalate precipitation to produce a mixed RE oxalate. Excellent heavy rare earth oxide (HREO) extractions (~ 90%) from beneficiation concentrates were achieved on the bench for both circuits, however flowsheet 1 was ultimately selected as the preferred flowsheet to pilot due to lower capital and operating cost. A mixed rare earth oxalate (42% TREO) or oxide (92% TREO) can be produced with low levels of Th and U of 3 ppm and 6 ppm respectively."

Jan 1, 2014

By Y. Nakamura, T. H. Okabe, S. A. Sunkar, K. Nose

Ruthenium (Ru) is known as a platinum group element with a very small production volume. In fact, the worldwide annual production of primary Ru metal is only about 20-30 tons, and it depends on the amount of Pt produced, since Ru is a byproduct from Pt smelting. The primary uses of Ru are as a material for resistors, as a catalyst for electrode materials, and as a sputtering target for hard disk memory storage. In order to recover Ru from the spent sand-blasting abrasive media generated during the cleaning of the apparatus used for Ru sputtering, a hydrometallurgical recovery process using hypochlorite solution has been investigated. In this study, various types of recycling processes that are presently under development are also introduced, and their features and differences are discussed. The advantages and disadvantages of the new recycling processes are evaluated, and the possibility of establishing new recycling processes in the future is discussed from a multilateral perspective.

Jan 1, 2014

Overview ?Oil Sands Background ?Water Use in Surface Mining ?Tailings Disposal ?Water Balance for Surface Mining ?Implications of Emerging Regulations ?Water Management Framework ?Tailings Performance Criteria (Directive 074)

May 1, 2009

By Dave Bulled

"This paper describes the design of an SABC grinding circuit using CEET technology applied to grindability data produced from testing 100 samples and distributed across the blocks of a mine plan using a geostatistical approach. The data distribution required the identification of the location of the sample points and a geological plan of the 130 million ton Ruby Creek orebody together with a mine plan of about 11000 blocks to be mined during the proposed 18 year life of the mine. The mill design was then made with consideration for the estimated hardness of each block in terms of both SAG and Ball milling with allowance for annual periods when the ore is harder to grind. Operating specifications were selected to achieve an optimum of capital and operating costs through use of the CEET program. The geostatistics assigned the statistical error that arises from having a limited amount of sample testing data to the hardness estimate of each block of ore in the mine plan. The final design could then include a quantification of risk and determination of safety factors based on the accuracy of block hardness estimates and CEET model fitting.INTRODUCTIONThe traditional approach to plant design involves the extensive testing of a single large composite sample or a small number of composite samples that are reputed to represent the ore body. It is accepted that the tests used can accurately measure in the laboratory the grindability of the sample by a process representing that to be used in the plant. The size of equipment required to achieve a specified throughput and product size is then calculated from one of a variety of models that have been developed by metallurgists over years of trial and error. Since these tests and models are tried and tested, they are accepted as reasonably precise; however during the operation of the resulting plant the design is sometimes found to be inadequate. It is then suspected that the flaw in the design process lies in the samples not being sufficiently representative of the ore body, since using only a single or small number of composite samples does not recognize the variability of the hardness of the ore, nor does it allow for the lack of precision in the hardness value used in the design.This paper outlines a design procedure suited to an orebody that is described by a geostatistical analysis of a reasonably large number of small samples of drill core. The analysis required the identification of the location of the sample points and a geological plan of the orebody, together with a mine plan of the blocks to be mined during the proposed life of the mine. The mill design was then made using the estimated hardness of each individual block with an allowance for annual fluctuations in the grindability of the ore. The geostatistical analysis assigned a statistical error to the hardness estimates of blocks and of each production year such that the final design included risk estimates and safety factors based on the possible errors that arise from having a limited amount of sample data (Kosick et al. 2002)."

Jan 1, 2007

By M. C. Makepeace

"It is the objective of designed experiments to obtain more information for less material and process cost than can be obtained by traditional experimentation. In this paper, the fundamental concepts on which the statistical design of experiments is based are briefly re-viewed according to the following outline: The Experiment The Design The Analysis Two simple 2"" factorial experiments are presented as examples."

Jan 1, 1967

By Trevor Yeomans, David Yaschyshyn

"Placer Dome (CLA) operates the 90-year-old Dome Mine, in South Porcupine, Ontario.The milling operation currently processes 5 million tons of ore per year, from both the underground and Open pit mining operations, producing 320 000 ounces. The milling process is conventional crushing and grinding, with gold recovery by gravity followed by leach/CIP.Cyanide tailings are impounded in a large (325 ha) tailings area where cyanide is allowed to degrade naturally. Effluent is discharged seasonally during the summer and early fall months to maintain site water balance. Pond water is reclaimed for sole use within the mill process.In 1995, an Effluent Treatment Plant (ETP) was constructed for the purpose of removing dissolved metals and suspended solids. In 1997, an INCO S02/Air cyanide destruction module was added.During this time period, the treated effluent often failed toxicity tests required by the Ontario MISA Regulation, even though effluent chemical parameters were below permit limits. Two studies identified ammonia as the toxicant to Rainbow Trout, and free copper as the toxicant to Daphnia magna.From a literature study, and contacts within the Ontario Mining Association, the use of chelation as a final end of pipe treatment step to render any remaining free copper and metal fractions biologically unavailable was considered. Lab and plant scale tests followed.During the 1999 ETP operating season, the addition of the chelant EDTA (ethylene-diamine-tetraacetic-acid) commenced. To date discharged effluent has passed all toxicity tests.The paper will review the effluent treatment history, current effluent treatment plant operation, and detail the use of EDTA for successful toxicity control.History of Tailings Deposition, and Effluent Treatment From 1910 to 1995Over it’s ninety years of operation, the Dome Mine milling circuit has changed from the original 1910 stamp-mill mercury plant, to today’s modem 14 000 ton per day automated gravity and Carbon in Pulp based plant.By the 19803, a modern mill was operating on mainly underground mined ore. The plant was operating at some 3 800 tons per day, using fairly conventional processing routes.The number 4 tailings dam was being used for tailing’s deposition.All effluent treatment was by natural degradation of cyanide. Discharge was from decanting type structures during the summer months only, maximizing the use of natural degradation no tailings effluent was recycled for Mill use."

Jan 1, 2001

By Mike Johnson, Niels Verbaan, David Dreisinger

The Search Minerals Foxtrot project in Labrador represents a significant Canadian resource of rare earth elements rich in the sought after heavy rare earth elements (19.8% HREE/TREE). A preliminary economic assessment of the project indicated that the project is technically feasible and economically attractive. A variety of metallurgical test programs have been undertaken which have led to significant streamlining and simplification of the process flowsheet. The improved flowsheet should allow Search to reach production earlier and will use conventional processing equipment. The results of these test programs are presented and discussed in this paper.

Jan 1, 2014

By Dave Bulled

"This paper describes the design of a molybdenite flotation circuit using FLEET technology applied to kinetics data produced from laboratory flotation testing and distributed across the blocks of a mine plan using a geostatistical approach. The geostatistics assigned the statistical error that arises from having a limited amount of sample testing data to the kinetics estimates of each block of ore in the mine plan. The final design then included a quantification of risk based on the accuracy of the block flotation kinetics and model fitting. The design made reference to the extensive technical reports describing earlier metallurgical test work including some pilot plant work over thirty years ago.The testing involved conducting MinnovEX Flotation Tests (MFT) to determine the rougher flotation kinetics at the specified reagent conditions and rougher-cleaner tests to determine the change in kinetics resulting from regrind. These flotation kinetics were used within FLEET simulations to create a full-scale flowsheet and plant design including recommended equipment sizes and the forecast of operating results. Measurement of detailed flotation kinetics for a wide range of samples allowed the effect of size of grind and efficiency of classification (i.e. width of size distribution) to be predicted for the whole ore body.INTRODUCTIONAdanac Moly Corp. proposed to develop a 20,000 ton per day mill to produce high quality molybdenite concentrate from its Ruby Creek resource in north-western British Columbia. It is a relatively clean low grade porphyry-monzonite occurrence containing minimal quantities of other sulphides such as chalcopyrite, pyrite, sphalerite and galena. Molybdenite occurs mainly as coarse flakes with lesser amounts of finer sized values. An optimum plant design was required to can achieve the predicted operating results, at a known level of risk, while treating the variability inherent in the ore body."

Jan 1, 2007

By John Russell

"The South American mining community leads the world in recognition and adoption of the latest grinding mill relining technologies. Clearly this region has recognised the benefits of improved relining performance.Six of the seven RUSSELL TWIN 8 Mill Relining Systems, ordered to date, are either operating or being manufactured for South American mine sites.This paradigm shift, towards very short relining times, aims at displacing mill relining activities from the critical path of concentrator maintenance shutdown schedules.The first uses of the RUSSELL TWIN 8 Mill Relining System have seen some of the highest SAG mill new liner placement rates ever observed by RME, as would be expected with two, independent Machines working simultaneously in the one mill.The liner-replacement-rate, averaged over the whole reline, is good (27% improvement over single Machine System in the same mill) but short of full potential. This modest result can be attributed to reline crews unfamiliarity with RUSSELL TWIN 8 System operation.This paper compares a sample of conventional (single machine) RME Mill Relining System performance data with one ‘double head’ mill liner handler by others. Early performance of Escondida’s new RUSSELL TWIN 8 System performance is then compared with recent single Machine data from Escondida, immediately prior to RUSSELL TWIN 8 commissioning. RME and Escondida look to constantly improving RUSSELL TWIN 8 System performance as full operational familiarity is achieved.Also, the preparation of this paper revealed a need to develop a global industry standard for mill relining data collection. A web based platform, to develop and then distribute standardised data collection templates, is discussed in Annexure 1."

Jan 1, 2010

By E. Wraith, W. W. Culver, B. Wesstrom, P. J. Mackey

Electrodeposited copper of the 1860s was of unprecedented purity, although the profound significance of this achievement was not fully realised until the dawn of the electrical age a decade or so later. By the 1870s there were three somewhat large (4 to 4.5 tonnes Cu/day) electrolytic copper refineries in Germany based on J. B. Elkington’s process, first patented in the United Kingdom in 1865. More such refineries were planned throughout Europe. At the same time, copper output in the U.S.A. was starting to expand and with the spread of European electrolytic copper refining technology across the Atlantic, the door opened to major increases in North American output. The technology was first taken up in 1879 by James Douglas at the Chemical Copper Company in Phoenixville, Pennsylvania, where he was Superintendent. Douglas was assisted in this pioneering work by Edward Weston. The copper refinery at the Chemical Copper Company was small, not unlike the first plants in Europe. The present authors contend that the electrolytic copper refinery at the Chemical Copper should be regarded as the first copper refinery in the United States. The Balbach Copper Refinery in New Jersey, which commenced operations in 1883 with an initial production of about 2 tonnes Cu/day, has been erroneously considered as the “first commercial refinery”; it rightly belongs in second place. The early developments at these plants are discussed in this paper. INTRODUCTION In the mid-1800s, the main uses for copper were as sheathing on the wooden sailing ships of the day to protect from marine growth, boiler components in locomotives and steam engines, along with some construction uses. Copper ores were generally smelted, and further fire-refined for market. The recovery of gold and silver, if present in payable amounts, was a slow and relatively expensive step. This situation regarding copper would change by the 1870s with advances in the generation and use of electricity. The introduction of the dynamo led to major developments in electrorefining. First demonstrated in principle by Faraday in 1831 (Faraday, 1832), the dynamo was subsequently adapted and further developed by J.S. Woolrich for electro-metallurgical applications in 1842 (Woolrich, 1842; Napier, 1845). It provided a new source of power with the unprecedented potential to significantly change the scale of metal production by electrometallurgy. Interestingly enough, copper would be needed for the electrical age just beginning. By 1869, advances in dynamo design offered relatively reliable large machines, which enabled J. B. Elkington in the United Kingdom to use the dynamo in the radically new Pembrey copper electrorefinery — the first copper electrorefinery in the world (Wraith, 2013). Initial production at Pembrey was small — in the range of 1 to 2 tonnes/day — while gold and silver were recovered more readily by the treatment of the slimes than with methods used hitherto.

Jan 1, 2019

By G. Thom, A. Kay, J. B. Hyde

"This paper will initially outline the specific types of copper castings used in blast furnace iron production. Tuyeres, tuyere coolers and stack cooling plates will be discussed in some depth, with particular reference to their dimensional configuration and purpose.Following this, the failure of such cooling members will be discussed, with special focus on tuyeres. Tuyeres are the most common casting which often requires replacement. Reasons for failure and its effect on operations will also be included.The latter sections of the paper will deal with the strict specifications demanded by ironmakers regarding copper castings. The justification for such rigid demands will be explained through the material presented earlier in the paper.IntroductionThe modern 5000-tpd blast furnace, worth millions, is a large capital investment to be sitting idle. Consequently, ironmakers strive to minimize the loss in iron production and revenue from furnace downtime. Two major causes of downtime are furnace relines and shutdowns to replace damaged tuyeresor tuyere coolers. Copper ancillaries are important items in the production of iron. They are consumable items, therefore they contribute to the campaign life or to the downtime in blast furnace production.Several combinations of lining and cooling systems have been used with varying degrees of success. Blast furnace refractories have evolved from ceramic to carbonaceous linings. The Russians and Japanese have successfully introduced stave-cooled furnaces that may have several forms, such as evaporative cooling. Stelco has two furnaces which use the Russian-designed cast iron system. However, this paper will be confined to the use of copper cast cooling members in a conventional internally cooled furnace.Figure 1 is a schematic cross section of the E blast furnace. The three major sections of a blast furnace-the stack, bosh and hearth-are identified. Further, the locations of the various copper members are indicated. This furnace produces an average of 5300 tons of hot metal per day."

Jan 1, 1982

By R. W. Lawrence

"In the past year, BioteQ Environmental Technologies Inc. have initiated design and construction of new plants to add to an existing list of plants that utilize the BioSulphide® and ChemSulphide™ process technologies for metal recovery and water treatment. In these plants, low cost biogenic sulfide or, in some cases, chemical sulfide reagent is used to selectively remove metals from contaminated water. The process produces high grade, marketable, metal sulfide products even from low grade feed solutions. The revenues generated from the sale of products can allow for water treatment at a profit in some cases, offset treatment costs in others, and generally reduce sludge disposal and management costs. The effluents produced from the process can meet stringent environmental standards. Current plants in operation and under construction range in sulphide demand from 50 kg/day to 3.7 tonnes/day, with the feasibility of larger plants with sulfide demands ranging from 6 to 43 tonnes/day under review.Biogenic sulfide reagent can also be considered for other applications in hydrometallurgy and mineral processing. Uses under consideration by a number of mining companies include: (1) replacement of sodium hydrosulphide in the SART process for cyanide regeneration and copper recovery in the processing of gold-copper ores, and (2) replacement of sodium hydrosulphide for use in the separation of copper and molybdenum in flotation. Biogenic sulphide, produced on demand at site, offers significant cost savings and provides the additional benefit of improved safety due to the elimination of the transportation, handling and storage required for chemical sulphide reagents.This paper reviews the applications of the BioSulphide® and ChemSulphide™ technologies for water treatment and metal recovery, and introduces projects where biogenic sulfide can be applied in hydrometallurgy and mineral processing."

Jan 1, 2007

By K Belanger, J. Mclachlan, F. Stober, G. Norval, B. Snider, A. Munsch, T. Koehler, E. Piantino, J. Carlos Queiroz

The project to rebuild the smelting furnaces at the Anglo American - Barro Alto ferronickel smelter, located in Goias, Brasil, was initiated in September 2013. To optimize the capital cost, while maximizing furnace integrity and robustness, a design basis giving flexibility to ore composition variation was developed in order to ensure sustained nameplate production. The resulting capital efficient, fasttracked project execution was accomplished through the creation of an integrated project team of Anglo American and Hatch through FEL 3 and FEL 4 phases of the project. The integrated team addressed the project challenges through a cooperative and versatile approach, created by access to knowledge, experience and systems in both Anglo American and Hatch. The team could operate at an optimized speed of execution, as demonstrated in the FEL 3 and FEL 4 completion durations, as well as efficiently mitigating any risks of increasing cost or delaying schedule by utilizing lessons learned and team knowledge from Anglo American on the original green-field project and Platinum smelters and Hatch on similar reference ferronickel furnace rebuilds. Other critical drivers of project cost minimization and revenue loss/generation (considering production revenue of each furnace) are scheduled durations for construction, commissioning and ramp-up. The integrated team commenced construction, commission and ramp-up planning almost one year in advance of the shutdown and this paper will discuss the strategies developed to reduce risks and optimize the construction schedule to mitigate revenue loss and optimize ramp-up time to sustained operation at nameplate capacity.

Jan 1, 2015

By D. A. Trotter, G. J. Goddard

"This paper describes a method of deciding on various design parameters for a sublevel caving layout. Layout design curves are obtained from the output of a computer optimization program. The program uses mathematical relationships derived from the theory of gravity flow of material to optimize the large number of interdependent design variables which are involved.IntroductionIn the design of an efficient sublevel caving mining system, a large number of interrelated variables need to be considered. Although many papers discuss individual variables, very little exists in the literature to show how these variables interact to achieve optimum results. This can lead to trial and error in sublevel caving design, which can prove costly, particularly because sublevel caving is a mining method which requires integration with sublevels above and below. With no pillars being left, experimental stopes cannot be isolated for study. Unfortunately, the complexity of layout design and the costs associated with trial and error have caused many planning engineers to choose more tried and tested mining methods.Sublevel caving works best to extract low-value ore surrounded by incompetent rock. All phases of the mining method can be highly mechanized, allowing low operating costs and high production rates. Good safety records are achieved, as workers are always in 'good' ground and the method is regular, making it comparable to an industrial process."

Jan 1, 1981