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By Marc C. Bétournay

Under the requirement to define the economic advantages of using hydrogen fuelcell vehicles, a broadscoped project was initiated to quantify the costs associated with operating fuelcell vehicles and diesel vehicles. While ventilation savings are clearly an advantage for the former, a complete and detailed analysis of this and other operational and capital cost elements was undertaken in order to provide the mining industry with factual information in considering the changeover from one technology to another. This article outlines the results of the investigations to date defining the costs/benefits associated with: capital purchases, operations (electrical/diesel, mine air heating fuel, ventilation, openings, surface and underground infrastructure, maintenance, labour, etc.), delivery of energy, regulatory obligations, and mine production. Considerations are given to the type and size of a mining operation, the depth, and location in North America. Significant ventilation power, equipment and infrastructure savings are possible with the application of fuelcells for shallower mine applications. At depth, where heat is an issue, there is a reduction in benefit. Specific qualifiers such as minimum air velocities, dust dilution, shaft and heating system velocity limits must be considered on an individual mine basis. While the capital cost for fuelcell mining equipment is currently higher than diesel versions, the high cost of fuelcells is anticipated to be reduced significantly in the near future with the intensification of manufacturing. In either case the operational savings would probably be sufficient to balance the supplementary capitalization costs. A comprehensive cost-benefit report will be issued in the near future.

May 1, 2003

By B. Stimpson, K. Kosar, K. Barron

"A number of high wall failures have occurred in some of the Alberta coal strip mines of a type called by the authors ""Multiple block plane shear slope failure "". They cause rapid deterioration of the dragline bench, reduced productivity and maintenance costs.The failures are invariably associated with the presence of almost horizontal (± 5 degrees) bentonitic clay bands in the overburden, often immediately above the coal seam, and regress behind the slope as a series of tension cracks which isolate detached blocks. As many as 17 sets of cracks have been observed, at lateral distances in excess of five times the slope height from the toe.The theory of this mode of failure is presented and a limiting equilibrium computer program which will produce deterministic or probabilistic (Monte Carlo) solutions is briefly described. A case history of one such failure is documented. It is shown that the extent of the regressive failure is exceedingly sensitive to the water table elevation, leading to the conclusion that drainage, if practical, represents the best control strategy. IntroductionA number of high wall slope failures are occurring in some of the coal strip mines in Alberta of a type called by the authors ""Multiple block plane shear slope failure""(I). These failures cause a rapid deterioration of the bench from which the dragline operates, reduced productivity, and necessitate conregresses into the slope as a series of tension cracks and detached blocks. As many as 17 prominent cracks or crack zones have been observed, the last being in excess of five times the slope height distant from the toe.The block between each pair of cracks must be at limiting equilibrium, i.e. the factor of safety equals one. This condition can be used to determine the locations of the cracks which develop in sequence behind the face."

Jan 1, 1986

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 V I. McCallum

In 1954, Falconbridge Nickel Mines Limited began the program of deepening the Falconbridge mine workings for 2,000 feet below the bottom, or 4025, level in order to provide twelve new levels for production. Involved in this program were a production-sized internal shaft, .crushing, pumping and hoisting installations, the integration of the new shaft operations with those of existing shafts, and considerations of ventilation requirements and ground support at depth. This paper describes the major items of equipment, and the depth development of No. 9 shaft as a whole, and also emphasize the planning phases of the program. The factors involved and reasons for the equipment and layouts chosen are given. Introduction and History THIS paper describes the planning and development of that portion of the Falconbridge mine between the 4025 and 6050 mining levels. This mine, the oldest of several being operated by Falcon-bridge Nickel Mines Limited in the Sudbury district, has been a continuous producer for over thirty years.

Jan 1, 1964

By P. Condori

Palabora Mining Copper has a unique ore body composition which unlike other copper ore bodies has three main components in significant proportions; these are copper bearing ore, magnetite and dolerite. The high content of dolerite, a hard barren material, creates a high proportion of pebbles which normally are recycled back to the mill. This action increases the mill bearing pressure and power progressively until corrective actions are taken. A logical action to solve the pebbles generation would be to incorporate a pebble crusher. This in fact was done however two fundamental issues were encountered; the high content of magnetite and existence of a cyclone classification section. Due to the high content of magnetite, metal detectors have not been able to work efficiently to separate tramp metal which causes damage to the pebble crushers. The existence of cyclones in the circuit had a detrimental effect on the final product when the pebbles crusher was online producing coarse sandy material. Due to the difficulties in treating pebbles via conventional procedures and the unique Palabora ore composition and characteristics, sorting technology was investigated at laboratory and pilot scale. It was found that a combination of optical and electromagnetic sorting could efficiently separate dolerite from the copper bearing and magnetite ores. This paper reports the findings of the laboratory and pilot scale work performed and discusses the additional benefits to the circuit.

Jan 1, 2012

By David Clarry

?The issue ?reduction requirement ?Planning abatement steps ?Energy Management & Energy Efficiency ?Energy Recovery / Utilization ?Alternative Energy Professional Services Overview ?Capital Projects and Construction ?Project Development and Planning ?Engineering ?Project management ?Construction management ?Consulting ?Process ?Business ?Environmental ?Technology ?Operational Services ?In-house engineering services for operations

May 1, 2009

By G. Couture, J. -F. Dubé, F. Letarte-Lavoie

"The Xstrata Nickel Raglan concentrator, located in Quebec’s Nunavik region, has been in operation since 1997. Originally designed to process 800,000 tonnes per year of high-grade nickel ore, the plant’s current capacity is 1.3 million tonnes per year due to various optimization projects. In 2011 and 2012, the Metallurgical group, with assistance from BBA, updated and optimized the variables affecting SAG throughput in order to optimize process stability while maintaining maximum plant throughput. This project consisted of reviewing the control strategy and completely updating the pre-existing Fuzzy Logic controller. This paper describes the successful update of the controller including a complete revision of control rules, a new process control strategy and the integration of new process variables, most notably mill rotational speed. It will also show how this new approach has provided a simple and robust way of delivering both an increase in throughput while reducing process variability.INTRODUCTIONThe Xstrata Nickel Raglan Mine concentrator is located in Northern Quebec along the 62th parallel. It has been in operation since 1997 and was originally designed to treat 800,000 tonnes of ore annually. The Raglan Agreement held with the Makivik Corporation and the Inuit communities limits the annual tonnage of ore that can be processed at 1.32 million tonnes.The ore is processed in a 24 foot variable speed, semi-autogenous grinding mill (SAG). The conversion from Autogenous Grinding (AG) to SAG was completed in 2006. Before the ore is processed by the SAG, its size is analyzed by a Wipfrag camera that provides a granulometry estimate. A cone crusher reduces the SAG’s discharge screen’s oversize material. A 14 foot diameter ball mill ensures final grinding before the flotation stage.Many control strategies were used at the SAG over the past few years. Fuzzy Logic was first used in 2007 in order to control tonnage feed to the SAG. This first version only measured the value of the variables without taking into account their rate of change. Revisions were made over the course of the last few years to improve controller performance. Other variables were added such as slurry density and crusher Closed Side Setting (CSS). The previous revision of the Fuzzy Logic was done in 2009 with a review of the rules only."

Jan 1, 2013

By M. F. Lopez, A. J. Mattus, S. B. Magid, D. B. Mobbs

"The use of Caro's acid as a replacement for hydrogen peroxide for the oxidation of vanadium prior to solvent extraction resulted in a 56% reduction in oxidant usage when compared on an equivalent hydrogen peroxide basis. Comparative oxidation tests, using two actual uranium mill raffinates, resulted in premature precipitation of an amorphous ferric vanadate when hydrogen peroxide was used to oxidize one of the raffinates containing a high iron to vanadium ratio. The use of Caro's acid eliminated the precipitation problem, thereby permitting the adjustment of the raffinate pH to a higher, more favourable level prior to solvent extraction.Potentiometric characterization of the raffinates revealed that vanadium is responsible for the inefficiency of hydrogen peroxide compared with Caro's acid, with the oxidation efficiency of iron being the same for both oxidants. A limiting potential between 720 and 740 mV was observed when using hydrogen peroxide, while, when using Care's acid, potentials approaching 1200 mV were achievable.A four-stage, countercurrent extraction performed on the oxidized raffinates did not reveal any operational problems. The extraction efficiency and the extraction coefficient were found to be higher when Caro's acid was used as the oxidant, as opposed to hydrogen peroxide.IntroductionMost vanadium is recovered as a by-product of some other process: leaching of Carnotite uranium ores, oil fly ash, phosphate rock or chromite processing. The major source of vanadium has been from uranium mills incorporating acid leaching of Carnotite ore s at elevated temperatures and pressures. The by-product recovery of vanadium from uranium ores of the Uravan Mineral Belt, located in the Colorado Plateau , is a common occurrence when vanadium recovery is economically viable."

Jan 1, 1983

By S. Fuentes, J. Cáceres

"During the past fifteen years, mines utilizing the block caving method have improved performance and systematically reduced total mining costs.This paper describes improvements in mine design, mine planning, and mine management, including methods enabling the Chilean and international mining industry to ensure cost savings and productivity growth while using the block caving method. These improvements have encouraged the re-evaluation of underground operations at operating and closed open pits, and challenge surface mining applications as the only low-cost mining method. This paper describes a strategy to estimate the optimum open pit to underground transition time during the life of a mine. IntroductionDuring the last fifteen years, block caving mines have experienced huge changes in designs, planning and administration. These changes have permitted operators to drastically reduce costs, improve equipment productivity, and improve overall business performance. As a consequence of this situation, the block/panel caving mining system is challenging actual open pit operations to maximize the net present value (NPV)."

Jan 1, 2004

By P. A. Riveros, W. D. Gould, R. Molnar, D. W. Koren, P. Bédard

This paper reports the results of the field demonstration campaign of a new process for removing and destroying ammonia from mining/ metallurgical effluent solutions. The 50-day campaign was carried out at a uranium plant in Saskatchewan, Canada. The process was developed by adapting and integrating two well established technologies: zeolite adsorption/ elution to extract the ammonia from the effluent stream and to produce a concentrated ammonia eluate; and conversion of the ammonia in the eluate to innocuous nitrogen by applying bacterial nitrification-denitrification (NH3 = > NO3 - => N2). The data show that reasonable target criteria were met, under very challenging operating conditions. Problem areas and possible causes for poor performance are identified. Preliminary and generic cost estimates of the process are also presented.

Jan 1, 2003

By T. K. Dichmann, J. A. Finch

"Microflotation, laboratory batch, and in-plant minicell continuous tests were conducted to determine if copper ions not only activate sphalerite but also depress pyrite when the minerals are together. Microflotation tests showed that pyrite recovery increased with addition of copper ions in the pH range of 5-10 when alone, but decreased when sphalerite was also present. Batch flotation tests (Red Dog ore) showed that upon addition of copper, pyrite recovery decreased significantly. The continuous tests (Louvicourt Mine) confirmed this and indicated that with excess copper pyrite recovery can increase. The results suggest there is competition for copper and xanthate in favour of sphalerite. To take advantage of this competition it may be necessary to control the sequence and rate of copper, collector, and lime addition. INTRODUCTIONFlotation, in common with all physical separations, is an imperfect process that results in misplacement of minerals. One source of this imperfection is accidental activation by contaminating metal ions. An important practical example is the misplacement of pyrite to zinc concentrates. Flotation of sphalerite generally requires activation by copper ions, added as copper sulphate. Copper ions are also used to activate pyrite [l] hence there is the suspicion that accidental activation contributes to the recovery of pyrite (Py) with the sphalerite (Sp). However, recent literature has shown that copper ions may make Py less floatable if Sp is also present. Xu et al. [2] conducted continuous minicell tests in-plant at Kidd Creek Operations (division of Falconbridge Ltd.) and showed that while the addition of copper sulphate improved zinc/iron selectivity, it did so as much by reducing the recovery of pyrite as increasing zinc recovery (Figure 1)."

Jan 1, 2000

By Willemijn Husslage

Tata Steel Group At a glance ? Part of the Tata Group ?Tata Steel Group: ?Fortune 500 company ?Top 10 global steel producer ?Annual crude steel capacity of 28 million tonnes ?More than 80,000 employees ?Manufacturing operations in 26 countries across 5 continents ?Serving customers all across the globe ?Turnover 2009-10: $22.8 billion

May 1, 2011

By Rudy Tenbergen, Dick Farmery, Jim Naylor

"Paste backfill has replaced cemented rock fill as one of the components to improve ground · control and mine safety at the Macassa Operation of Kinross Gold Corporation in Kirkland Lake, Ontario. It is the first operation to utilize "" flash thickening ""of very fine gold tailings to a paste consistency in the ""Paste Production and Storage Mechanism"", a joint technical development of Inco and GL&V. The tailings paste is mixed with alluvial sand to produce an engineered quality paste backfill. The development leading up to the decision to proceed with this novel technology is discussed, as well as commissioning of the backfill operation.IntroductionThe occurrence of a major rockburst on November 23, 1993 prompted a review of mining and backfill procedures at the Macassa No 3 shaft . This shaft is some 7238 ft deep and lies in a mining belt that is prone to seismic events. The recommendations were to change the mining from underhand cut and fill to mechanized long hole stopping and to look at alternatives to the rockfill backfill system in place. The Macassa cemented rock backfill operation, while producing the desired fill stiffness had drawbacks as compared to paste fill. With paste fill, placement rates are faster, more openings can be filled, it is less expensive and less prone to the occurrences of compensable incidents.Macassa mining personnel visited the Sudbury operations of Inco in 1993, where development on paste backfill was in progress. As some of the R&D was taking place with funding from the Premier's Council Technology Fund of the Ministry of Trade and Technology of the Province of Ontario , Inco personnel at that time did some feasibility and conceptual consulting on the use of paste fill for the industry. Inco had developed the use of alluvial sand paste fill, an engineered blend of alluvial sand and tailings fines and was doing development work to eliminate the ""conventional"" way of producing paste from mill tailing by thickening and filtration. This latter development work was taking place with GL&V Ontario. That work led to the filing of a joint Inca/ GL&V patent application for the PPSM, the Paste Production and Storage Mechanism with all marketing rights being held by GL&V."

Jan 1, 1997

By A. C. Wilson

In 1991, local prospector C. ?Mickey? Clement recovered at least three alluvial diamonds from the Michipicoten River. Two of these stones were sent to the Department of Mineralogy, Royal Ontario Museum, where they were identified as industrial grade diamonds with weights of 1.05 and 1.13 carats. Both stones were described as frosted and graphite-inclusion riddled. In 1995, prospector Sandor Surmacz and geologist Marcelle Hauseux discovered diamonds in bedrock approximately 20 km north of the Michipicoten River. An 18.1-kg bulk sample of a xenolith-rich lamprophyre, taken from an outcrop at the side of the Trans Canada Highway north of Wawa, yielded 1 macro diamond and 5 micro diamonds. All but one was gem quality (Sage 2000). Since then, more than 36 occurrences of diamonds in bedrock have been reported in an approximately 30 km2 area centred 20 km north of the town of Wawa. The occurrences are hosted within a sequence of unusual, Archean-aged, heterolithic breccias previously considered to have no economic significance. Ten kimberlite occurrences also have been confirmed in the Michipicoten greenstone belt in the last 5 years. Two of these kimberlites are diamondiferous. To date, expenditures on property acquisition and diamond exploration in the Wawa are estimated at $120 million (CAN). This paper summarizes the geological observations made on these occurrences and discusses including the relationship between the occurrences and the regional geology.

May 1, 2003

By Tim Martin, Dave Tahija

"In April, 2014, Hecla's Greens Creek mill, near Juneau, Alaska, switched from sulphuric acid to the use of dissolved carbon dioxide for lead flotation circuit pH control. The Greens Creek mine uses paste backfill, in which mill tailings are mixed with cement and used to fill mined areas. Some backfill inevitably becomes part of mill feed during subsequent mining cycles, resulting in high feed pH levels and increased dissolved calcium in the lead flotation circuit. Sulphuric acid is traditionally used to lower pH in such cases but it produces undesirable side effects, including poorer lead-zinc separation and unstable froth. The amount of sulphuric acid used at Greens Creek was limited due to these effects, resulting in higher than target pHs when feed backfill content was high. Dissolved carbon dioxide can be added without these limitations so pH levels have been under much better control since the switch. Additionally, lead-zinc separation and froth characteristics have significantly improved.IntroductionThe Greens Creek Mine, owned by Hecla Greens Creek Mining is an underground mine located on Admiralty Island near Juneau, Alaska, producing silver, gold, lead and zinc in the form of lead and zinc flotation concentrates and a gravity gold concentrate (McDonald, Morgen & Hancock, 2012). The lead and zinc concentrates are produced in conventional flotation circuits preceded by a graphite removal flotation step.The underground mine utilizes cemented paste backfill, of which a certain amount is inevitably re-mined in later mining cycles and returned to the mill, mixed with fresh ore. The Portland cement used in the paste backfill is highly basic so that, from the point of view of the mill, the net effect is to introduce a good deal of lime to run-of-mine mill feed, producing a very high natural pH. This high pH needs to be lowered to obtain good metallurgical results."

Jan 1, 2016

By S. Castro, L. Gutiérrez, L. Uribe

"Use of lime in seawater as a pyrite depressant decreases Mo recovery in flotation of Cu- Mo sulfide ores. When technology that uses sodium metabisulfite (MBS) to depress pyrite is applied, the process is carried out at pH 6.5–7.5, the range where Mo recovery is higher than the usual pH 11.0–12.0 when lime is applied. Redox conditions and chemical reactions involved in depression mechanisms of pyrite by MBS, or sodium sulfite, in seawater are investigated and discussed based on the acidifying effect of MBS, oxidation of sulfite ions by dissolved oxygen, and the role of catalyst of some metal cations (e.g., copper ions) adsorbed on pyrite’s surface.RÉSUMÉ L’utilisation de chaux dans de l’eau de mer comme agent déprimant de la pyrite réduit la récupération de molybdène (Mo) dans la flottation de minerais sulfurés de Cu-Mo. Lorsque l’on applique la technologie utilisant du métabisulfite (MBS) de sodium comme agent déprimant de la pyrite, le procédé se fait à un pH de 6,5 à 7,5, gamme dans laquelle la récupération de Mo est supérieure au procédé habituel utilisant un pH de 11,0 à 12,0 lorsque l’on applique de la chaux. Les conditions de réduction-oxydation (redox) et les réactions chimiques impliquées dans les mécanismes de dépression de la pyrite par le MBS (ou sulfite de sodium) dans de l’eau de mer sont étudiées et débattues sur la base de l’effet acidifiant du MBS, de l’oxydation des ions sulfites par de l’oxygène dissous et du rôle de catalyseur de certains cations métalliques (par exemple les ions de cuivre) adsorbés à la surface de la pyrite. INTRODUCTIONThe copper industry is aware of the difficulties of floating Cu-Mo sulfide ores in seawater. Molybdenum recovery is partially depressed when lime is used to reject pyrite. In plant practice, sodium metabisulfite (MBS) has been demonstrated to be the most promising option for replacing lime in seawater."

Jan 1, 2017

By Peter Rubie, Toni Kojovic, Fengnian Shi

"Mining companies have recognized that characterization of ore breakage plays an essential role in the design and optimization of comminution circuits.Recently the Julius Kruttschnitt Mineral Research Centre (JKMRC) has developed a new breakage model incorporating the effect of particle size, and a new breakage characterization testing device, called the JK Rotary Breakage Tester (JKRBT), which allows rapid testing of particle breakage under high energy single impact and low energy repetitive impact conditions.Positive feedback received from mining company sponsors to the new technology has prompted the JKMRC to further validate the device and the model through a formalized validation program using machines designed and manufactured by Russell Mineral Equipment (RME). The first unit was installed and commissioned at the Anglo Platinum laboratory in Johannesburg in March 2007 and four more major mining companies will participate in the validation program in 2008, including Barrick, BHP Billiton, Rio Tinto and Teck Cominco. The results from validation tests at the JKMRC and Anglo Research have shown that the JKRBT generates statistically similar breakage parameters to the traditional Drop Weight Test, but with significant advantages over its predecessor. Application of the JKRBT in geometallurgical testing with the AMIRA P843 GeM project has confirmed the JKRBT device offers a rapid and consistent method for determining the hardness of drill core samples. The JKRBT has the potential to revolutionize ore testing, with applications in laboratory impact breakage characterization of rock samples, drill cores, coal and other materials. It also stands poised to significantly advance the process control and management of comminution circuits since the concept is readily adaptable to on-line ore hardness measurement.INTRODUCTIONMining companies have long recognized that the performance of a comminution machine such as a crusher or an autogenous / semi-autogenous grinding (AG/SAG) mill depends not only on the machine operating conditions, but also on the “hardness” of the feed material to be broken. This recognition has driven significant interest in laboratory particle breakage characterisation and testing. The laboratory breakage characterization tests can be roughly classified in two broad groups based on their primary breakage modes: tumbling and impact/compression. The famous Bond ball/rod mill tests (Bond, 1952 and 1961) represent the typical tumbling test and have become the industry standard for estimating the specific energy (kWh/t) of rod and ball mills. However, as the energy application in the single particle impact tests can be well defined and controlled, the latter have received significant attention in the last 30 years. It was estimated that mining companies worldwide spent over three million US dollars on single particle impact breakage characterization testing in 2006."

Jan 1, 2009

By B. G. Cousins, B. O’Rourke, C. Spence

"Some non-flotation chemistries are added to flotation circuits directly and indirectly to control issues such as scaling and dust. Since the dosage levels are generally not very large, a standard laboratory float test may not be able to determine the effect a specific chemical has on the productivity of the flotation circuit. Hudbay and Ashland Water Technologies developed a laboratory testing protocol that uses kinetic float studies to test the effect of normal to extreme doses of antiscalant and dust chemistries (5 ppm to 500 ppm). Laboratory results indicate that this protocol helps to predict the potential interference that a particular chemistry will have on the flotation circuit.INTRODUCTIONFrom a chemical perspective, flotation is a very complex system involving more than the reagents used for flotation. Along with collectors, frothers, modifiers, suppressants and activators, there can be antiscalants for water treatment, dust suppressants used in the mining process, residual flocculants from recycled thickener overflows, grinding aids, and many other types of chemicals that are added upstream of a flotation circuit. Although these other chemicals are usually used at very low doses, it is important to know if they affect the flotation process detrimentally in the event of a spill or a build up from recycled water.Laboratory flotation testing of these chemicals characterizes their effects, but testing at expected doses in the float can mask the effects of the chemicals in the plant. Also, simple tests that do not measure the kinetics of flotation do not reveal the effects the other chemicals have on the flotation circuit. Thus, a new flotation protocol was needed to expose the full effects of chemical additives and to assist in choosing the chemical with the lowest potential for a negative effect.FLOTATION KINETICSMeasuring recovery kinetics in a laboratory flotation cell is a well-established procedure: analyzing concentrates taken in a timed sequence to determine the recovery rate of individual cells in a flotation circuit. The recovery from the early flotation cells are a true measurement of how well a flotation chemical is performing. The principles of kinetics can be mathematically determined by the following functions (Houston, 2006):"

Jan 1, 2013

By Corby G. Anderson

"Laboratory testing, plant optimization, and the interpretation and application of metallurgical technologies can be tedious, time consuming and costly. This paper outlines the use of proven statistical design of experimentation software for rapid laboratory testing and optimization of operating plants with limited representative sample utilization. This results in less costly required testing, a more thorough understanding of results and the ability to simultaneously optimize several variables and outcomes at once. At the Center for Advanced Mineral and Metallurgical Processing (CAMP), this testing methodology has been used successfully in many metallurgical development and plant optimization projects thereby confirming its real world utility and application. Accordingly, several applied examples in flotation and hydrometallurgy will be outlined in this paper. This will include flotation testing of a copper gold ore, flotation testing of a copper, cobalt and gold ore, optimization of an industrial hydrometallurgical jarosite circuit and optimization of industrial nitrogen species catalyzed pressure leaching of a copper enargite concentrate.INTRODUCTIONWhile flotation technology is now 100 years old, the testing, interpretation and application of lab results can be tedious, costly and time consuming requiring extensive sample collection and preparation. In addition, recent applications like hydrometallurgical pressure leaching and associated unit operations, such as precipitation, still remain more of an art than a quantifiable science. This paper will elucidate the application of STAT-EASE Design-Expert software to minimize required laboratory testing while rapidly optimizing multiple results and outcomes in a statistically valid manner. The STAT-EASE program is based on proven design of experiment fundamentals. In this paper, the real world application of this methodology for lab testing, plant optimization and process development will be elucidated."

Jan 1, 2006

By J. S. Brock

Magnetic, electromagnetic and gravimetric surveys have played an important role in the mineral exploration leading to the discovery of three major massive sulphide deposits in the Anvil-Vangorda area of the Yukon Territory. During 1964 and 1965 a saturation prospecting program, relying heavily on the integrated use of airborne and ground geophysical survey techniques, led to the discovery of the Faro deposit, which was subsequently proven to be a 63-millionton lead-zinc orebody. Locating the Faro deposit, the largest found to date in this overburden-covered region, involved an exploration sequence requiring the selection of a favourable geologic environment, aeromagnetic and electromagnetic surveys, property acquisition of anomalous areas and detailed ground geophysical surveys, followed by rotary drill testing, detailed gravimetric surveys and diamond core drilling.

Jan 1, 1973