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By E. Haccuria

Due to the growing share of electrical cars, the use of lithium ion batteries significantly increased. Recycling of these batteries is essential to safely dispose hazardous materials as well as to recover valuables (Cu, Ni, Co, etc.). In order to optimise the recycling process, accurate information is required on the slag phase equilibria for battery smelting. The slag is a multi-component system containing alumina from the battery cases, silica and lime from fluxes, and manganese and lithium from the electrode materials. The present study accurately determined the phase equilibria in the "MnO"-Al2O3-SiO2 sub-ternary system at metal alloy saturation. Particular attention was given to the accuracy and reliability of the final results by highlighting the different reaction pathways, mass transfer- and reaction mechanisms taking place in the system during experiments. This attention enabled improved experimental methodology and measurements. Based on the present study, further measurements will be undertaken in the quaternary "MnO"-Al2O3-CaO-SiO2 system at low "MnO" concentrations, which closely represents the future recycling industrial slags.

Jan 1, 2014

By D. Stone

"The Beal Mountain mine near Anaconda, Montana, is an interesting case history of a mine that has continued operations despite several large pit wall failures over a period of approximately 10 years. The largest event involved some 3.9 million tons of rock that has moved steadily at a rate of about 0.05 ft/day since 1992. This paper documents the geotechnical instrumentation and monitoring programs implemented at Beal to provide the company with a comfortable baseline that supported ongoing operations. The success and failure of a variety of stabilization measures will be discussed, along with typical instrumentation responses for several significant events. The success of Beal Mountain serves as a good model for mines faced with serious geotechnical challenges."

Jan 1, 2002

By H. Sahay

"DISCLAIMER NOTICEThe views and conclusions contained in this presentation are those of the author and should not be interpreted as necessarily representing the official policies or recommendations of the Department of Mines and Energy, Government of Newfoundland and Labrador. ABSTRACTThis paper describes the main testing procedures that should be carried out to evaluate the capacity and capability of the service brakes on the large-size diesel-electric trucks now in common use at open-pit mining operations. Brake testing programs in general are described, and then consideration is given to different techniques of service brake testing, brake performance evaluation, braking energy and typical performance characteristics.IntroductionThe first diesel-electric truck with a capacity of about 85 tons was introduced in 1963 by Unit Rig and Equipment Company of Tulsa in cooperation with General Electric Company, who successfully developed and supplied the most important component- motorized wheels. Since then, its popularity and numbers have continued to grow, and now more than 3700 such units are in operation in mines around the world. In Canada alone, their number is more than 700, including about 150 in the Province of Newfoundland and Labrador. Similarly, their size has also grown from that initial capacity of 85 tons to 350 tons.Although the introduction of the electric trucks proved to be invariably economical, their indiscriminate use was not without severe consequences. Many runaways and even fatalities occurred when the dynamic retarder failed while the truck was travelling loaded down severe grades. The service brakes in such cases proved incapable of stopping runaway trucks, mainly due to their limited capability and the fading caused by excessive heat produced during the act of braking itself. This eventually led the various governments, the open-pit operators and the manufacturers to seriously question the braking capability as well as function of both service brakes and dynamic retarders. Extensive testing"

Jan 1, 1980

By Ross Nairn

"Tulawaka Mine is a joint venture between Pangea Minerals Limited 70% (a wholly-owned subsidiary of Barrick Gold Corporation) and Northern Mining 30%. Tulawaka Mine is located in North West of Tanzania in the Lake Victoria Goldfields. Tulawaka Mine is scheduled to process 360,000 tonnes per year at an average 11.5 g/t head grade over a 4 year mine life.The Tulawaka Process Plant uses single stage crushing, single stage SAG milling, gravity, intensive leach of gravity concentrates, CIL, Zadra desorption and cyanide detoxification to process the ore. Gold recovery exceeds (or is expected to) 95%.The Process Plant was designed and constructed under the supervision of Ausenco Engineering of Australia. Site works commenced in January 2004, Process Plant construction commenced in July 2004 and the first ore was presented to the Process Plant 26th February 2005. The first gold was poured on 15th March 2005. All construction contractors were demobilized from site 7th April 2005. Major achievements have been 3.4 million hrs LTI free during construction and operation to date, training of the Tanzanian workforce at Bulyanhulu Mine prior to commissioning, a succession plan to replace expatriate personnel with Tanzanian staff and significant over performance of the Process Plant against design.INTRODUCTIONTulawaka Mine is situated approximately 160 km south west of Mwanza in the western Lake Victoria Goldfields District of north western Tanzania. The property is held under a joint venture between Pangea Goldfields Inc (70%), a wholly owned subsidiary of Barrick Gold Corporation, and Explorations Minieres Du Nord Ltée (30%). A location map showing the position of the operation in Tanzania is shown in Figure 1.Mine site staff at Tulawaka comprises a combination of Tanzanian and Expatriate personnel. All Pangea personnel live on site and commute to their place of residence for their R&R on a 9 weeks on 4 weeks off roster."

Jan 1, 2006

By Ben Murphy, Tomi Maksimainen, Antti Rinne, Ishmael Muzinda, Toni Mattsson

"A 500m3 flotation cell was commissioned during 2014 at the Kevitsa Mine in northern Finland working as the first rougher in the copper flotation circuit. Treating around 850tph with an average grade of 0.3% Cu, the TankCell® e500 has run with an average copper recovery of 71% and an average enrichment ratio of 50 during early optimization test work. In addition the cell has proved to be very energy efficient, being operating with a specific energy of around 0.4kW/m3 under production conditions.A review of the installation and operation of the TankCell® e500 at Kevitsa will be presented. Metallurgical performance of the cell during the operating period will be discussed on the basis of on-line analysis and the metallurgical samples collected during early optimization campaigns.IntroductionThe past three decades have seen a rapid increase in the size of mechanical flotation cells. This has been largely driven by demand from the copper industry to process larger volumes of material due to combination of declining head grades and the desire to benefit from the economies of scale (Lynch, Harbort and Nelson 2010). During this period Outotec has been at the forefront of development of large flotation technology with an enviable scale-up track record; 100m3 (1995), 160m3 (1997), 200m3 (2002), 300m3 (2007), 500m3 (2012) to the largest commercially available cell at 630m3 in 2013. This scale-up methodology has been developed, tested and validated (Elgueta et al., 2009; Coleman & Dixon, 2010) by the large installation base of cells in the 300m3 size class operating (over 400). The learning from large cell operation are then reviewed and incorporated back into new designs and scale-up practices (Murphy, Miettinen, and Yanez, 2014)."

Jan 1, 2016

Overview n Why we audited the IRS n How we did it n What we discovered n What we learned n Where we are going next The unbroken chain of safety The Internal Responsibility System (IRS) The network of roles and relationships in a workplace, through which everyone works with mutual commitment to reduce or eliminate hazards.

May 1, 2003

By O. Rabeau, M. Legault, F. Fallara

"Abstract - The development of robust 3-D geological models involves the integration of large amounts of public geological data, as well as additional accessible proprietary lithological, structural, geochemical, geophysical, and diamond drill hole data. 3-D models and maps have been available, particularly in the petroleum industry, for more than 10 years. Here, we demonstrate how robust 3-D maps can be used as interactive tools for mineral deposits exploration. In particular, we show how the interrogation of 3-D data sets can constrain exploration targets at depth. The main advantages of this technique for the mining industry are the homogeneity of data treatment and the validation of geological interpretations, taking into account geophysical and geochemical data. Data integration and cross-correlation of geology and geophysics can be achieved in two dimensions in any good GIS package. However, the added strength of 3-D analysis is the integration of separate data sets in three dimensions to build more complete, more realistic models, and in delineating areas of high economic potential at depth. Furthermore, these models can be modified and improved at any time by adding new data from ongoing drilling and geoscientific surveys. This paper presents two examples of 3-D models used for mineral exploration: the Joutel VMS mining camp and the Duparquet gold camp, Quebec. In both examples, the creation of the model is discussed and queries specific to the relevant exploration model are introduced. Eight potential exploration targets are generated at Joutel and seven at Duparquet. Although the targets defined are dependent on the details of the chosen queries, it is apparent that this technique has the potential to generate promising exploration activity that can engender new targets. Sommaire - Le développement de modèles 3-D géo-intégrés fiables implique l’intégration de grandes quantités de données géologiques publiques ainsi que des données lithologiques, structurales, géochimiques, géophysiques et de forages en provenance du secteur privé. L’industrie pétrolière dispose de modèles et de cartes 3-D depuis plus de 10 ans. Ce manuscrit présente un exemple d’utilisation de modèles et de cartes tridimensionnels détaillés comme outils interactifs dans l’exploration de dépôts métallifères. L’emphase est mise sur le processus d’interrogation des données en 3-D pour la définition de cibles d’exploration en profondeur. Les avantages principaux de cette technique sont l’uniformité du traitement des données et de la validation des interprétations géologiques, tout en tenant compte des données géophysiques et géochimiques. L’intégration de données et la corrélation de la géologie et de la géophysique peuvent être accomplies en deux dimensions dans n’importe quel logiciel SIG. La plus grande force de l’analyse 3-D demeure cependant la possibilité de construire des modèles complets et réalistes permettant de cibler des secteurs à haut potentiel économique en profondeur. De plus, ces modèles peuvent être modifiés et améliorés de manière interactive en ajoutant de nouvelles données provenant de forages et de levés géoscientifiques en cours. Ce manuscrit présente deux exemples de modèle 3-D géo-intégrés développés pour l’exploration minérale : le camp de SMV de Joutel et le camp aurifère de Duparquet, Québec. Dans les deux cas, la construction du modèle est décrite et les requêtes spécifiques à chaque type de gisement sont présentées. L’exercice a permis de définir huit cibles d’exploration potentielles à Joutel et sept à Duparquet. Bien que les cibles identifiées dépendent des contraintes utilisées pour les requêtes choisies, cette technique permet de générer de nouvelles cibles qui pourraient relancer l’activité d’exploration dans les camps miniers " matures " ou dans des secteurs moins explorés."

Jan 1, 2006

By Paul R. Clark

Outline ? Coal As An Energy Resource ? Canadian Clean Power Coalition ? Phase I of the CCPC Project ? CO2 capture options ? Gasification ? CO2 use ? Results ? Phase II Plans ? Next Steps TransAlta - Profile ? 9,726 megawatts (MW) of capacity ? 46,877 gigawatt hours (GWh) produced at an average plant availability of 88.4% ? 47 facilities in four geographies: Canada; the U.S.; Mexico; and Australia ? Fuel mix of coal (62%), gas (26%), hydro and renewables (12%) ? 23.2 million tonnes of coal mined ? 47.0 million gigajoules (GJ) of gas consumed

May 1, 2004

By H. A. Khorami, T. Wong, B. Velapatino, E. Asselin, T. Donnelly, E. Bryce, R. Dixon

A copper–nickel concentration gradient alloy was designed and fabricated through a powder metallurgy (PM) process. Copper and nickel powders were mixed with different weight ratios from 0 to 100% copper in 10% increments. Each powder mixture was poured inside a cylindrical die and compacted between two punches using a hydraulic press machine at pressures up to 70 MPa, resulting in approximately 1 mm thick layers (discs) of pressed Cu–Ni powder. These 11 Cu–Ni compacted powder layers, i.e. 0–100% Cu with 10% increments, were then compressed together with pressures up to 420 MPa to form a green part. The green part was sintered in a vacuum furnace at 1000°C for 10 hours to form a solid mass (PM part) of Cu–Ni concentration gradient alloy. In the sintering process, Cu and Ni particles fused together and formed a single piece due to diffusion of the atoms across the boundaries of the particles. The diffusion of atoms also converted the stepwise concentration change into a uniform concentration gradient. INTRODUCTION Copper has numerous applications in different industrial sectors due to its superior physical, electrical, mechanical and chemical properties such as its excellent thermal and electrical conductivity, resistance to fatigue, high ductility, and resistance to corrosion. Moreover, copper and its alloys have natural antimicrobial properties. This has opened up new applications for copper–based self–disinfecting surfaces in healthcare in an effort to mitigate healthcare–associated infections (HAIs). HAIs affect millions of patients each year, leading to significant morbidity and mortality as well as financial losses to healthcare systems (World Health Organization, 2016). HAIs spread through three main routes: (a) iatrogenic in which HAIs are a direct consequence of medical encounters and/or treatment; (b) person to person where organisms are transferred between healthcare workers, patients, and visitors; and (c) environmental where transmission occurs via contaminated surfaces and equipment (e.g. doorknobs, railings, toilets/commodes, medical carts). The use of antimicrobial copper–based materials on high touch surfaces could diminish pathogen transmission via this latter rout. This has subsequently led to the development of a variety of copper–based materials for use in healthcare settings and promoted the investigation of the antimicrobial efficacy of copper–based materials.

Jan 1, 2019

By M. Maccagni, J. Nielsen

"Engitec developed the EZINEX® Technology some years ago to recover zinc from oxidized secondary sources. This process works at neutral pH and can take into solution any kind of metal, but iron is rejected and remains in the leach residue. Engitec has also developed the FLUOBORATE System, which is able to recover lead and copper from relevant concentrates. Some iron is leached into solution, but it is controlled through a solution purification bleed treatment. In this paper, we present the two above mentioned technologies. They are very different with regards to the chemical environments and the methods used to control the build-up of iron in solution, but both systems are suitable to treat iron bearing feeds.INTRODUCTIONIn the design of hydrometallurgical processes, the most important issues are the water balance, the filtration and the control of impurities. This is particularly important when the process includes an electrowinning unit. Iron has been one of the most problematic elements to remove, especially in acidic environments.Engitec Technologies has developed several hydrometallurgical processes divided in two different electrolyte lines: the Fluoborate Technology for the recovery of lead and copper from primary and secondary sources and the Ammonium Chloride System based on a chloride electrolyte for the recovery of different non-ferrous metals, among which zinc and lead are the most important. In both cases, iron is present in the feed and needs to be eliminated from the solution. It is not a problem for the quality of the metal produced, but when present in solution, it consumes the basic electrolyte reducing its free concentration which is very important both for both the leaching and electrowinning unit operations. Both process systems are very peculiar, the fluoborate case being much more intriguing."

Jan 1, 2016

By Junior Santos

Kinross Paracatu started the operation of the new gold processing plant in the end of 2008, being composed by one SAG mill and two ball mills. During the first year of operation the plant had several problems related to stability and throughput capacity, mainly due to low grinding efficiency of the ball mills. In this period many actions were taken to maximize the throughput rate. Among the actions are reducing the size of balls in the mills, increasing the ball charge in the SAG mill, changing cyclone configuration and return of circulating load to SAG mill. With better stability of the circuit, the second stage of improvements has begun with the objective of increasing throughput rate and also reducing operation costs. Among the actions are the optimization of ball charge in the mills and a better operational control of cyclone parts. After two years of operation was obtained a significant increase in plant capacity, but still requires the addition of a third ball mill in order to maximize the use of SAG mill, which has spare capacity.

Jan 1, 2012

By Ltd. (CYMCO) Technology Center Yunnan Metallurgical Group Co., S. Jian, T. Song, J. D. Miller

"As sulfide zinc ore deposits become exhausted, nonsulfide zinc ore deposits become the other resource for zinc production. However, flotation for zinc recovery from nonsulfide deposits is much more difficult than the flotation of sphalerite from sulfide deposits and may become an option for zinc production only if a suitable processing technology is developed. In this regard, research efforts are being devoted to the flotation of nonsulfide zinc oxide minerals. Some common methods which have been used for the flotation of zinc oxide minerals include sulfidization-flotation, activation, and direct flotation with anionic and cationic collectors. The traditional flotation techniques have low efficiency and difficulty in separation and recovery of the zinc minerals from gangue minerals, for example, smithsonite (ZnCO3) from other carbonate gangue minerals and hemimorphite (Zn4Si2O7 (OH)2·H2O) from silicate gangue minerals. In this study potassium lauryl phosphate was evaluated as a collector for the flotation of nonsulfide zinc minerals. Results from contact angle and microflotation experiments at pH 6–9 show that hemimorphite (Zn4Si2O7 (OH)2·H2O) can be floated from quartz and that smithsonite (ZnCO3) can be floated from calcite. However, when dodecylamine, which is a common collector reported in literature for the flotation of nonsulfide zinc minerals, was used as collector, no significant difference between nonsulfide zinc minerals and gangue minerals was observed at any pH value. Initial evaluation suggests that potassium lauryl phosphate is effective at low concentration and may be a promising collector for zinc oxide mineral flotation.INTRODUCTIONBefore the twentieth century, zinc was mainly produced from nonsulfide ores (Large, 2001; Gilg, Boni, & Cook, 2008). With the development of differential flotation processes in the late 19th century major production switched from nonsulfide zinc sources to sulfide zinc ores (Hitzman, Reynolds, Sangster, Allen, & Carman, 2003; Gilg et al., 2008). However, as it is becoming more difficult to find new sulfide ores, recovery from nonsulfide deposits is being considered. Nonsulfide flotation for zinc recovery is much more difficult than the flotation recovery of sphalerite from sulfide deposits and may become an option for zinc production if a suitable technology is developed. Zinc and lead oxide ores are difficult to beneficiate ores due to their similar floatability with gangue, high solubility in water, and the influence of slimes (Fa, Miller, Tao, & Li, 2005). In order to improve the performance of oxidized minerals, lots of effort is being devoted to this field of study (Ejtemaei & Mahdi, 2014)."

Jan 1, 2016

By R. G. Zaripova, N. D. Ryan, H. J. McQueen

The austenitic and ferritic stainless steels do not undergo any phase transformation that induces very fine grains in the product through thermomechanical processing as practiced in most other steels. Considerable grain refinement can be achieved by intense multistage deformation descending through warm to cold working; notably multi-axial forging of bulk material can attain high strains with diminished risk of cracking. Hot working to improve ductility yields finer grain sizes as the T is reduced and strain rate E raised; increased strain has limited effect due to the steady state regime. Ferritic steels develop subgrains with size diminishing as E rises and T falls with the advantage of retained strain hardening; recrystallization delivers grains several times the cell size. In austenitic steels, dynamic recrystallization (DRX) provides finer grains at higher E and lower T with a lower limit due to intergranular cracking. The DRX grains contain a strengthening substructure that can easily be retained by quenching; if static recrystallization is allowed to occur the grain size rises markedly. Multistage rolling with diminishing T is able to provide grain refinement to the same degree as straining at the finishing T yet provides better ductility. grains on recrystallization. As a result of the strain induced transformation to martensite, small subzero deformations of austenitic or duplex steels produce very high dislocation densities that yield marked grain refinement when they revert upon reheating.

Jan 1, 2004

By A. Desbiens, A. Pomerleau

During the summer of 1994, a new PID-based control strategy was installed on line B of the grinding circuit at the Kidd Creek Division of Falconbridge Ltd. Part I of this paper described the methodology used to select the appropriate control strategy using a phenomenological dynamic simulator. Part II presents the implementation of the control strategy. Prior to the controller installation, it was necessary to solve various problems related to the process, such as a cyclic behaviour of the rod mill feed tonnage, a problem related to the mill water source and the calibration of the flowmeters. Once these problems were solved, the control strategy was accomplished in two steps. Step 1 corresponds to the installation of control loops to regulate the pump box level and the cyclone overflow density. Step 2 describes the implementation of a loop to control the cyclone feed density. According to Part I, a loop to control the circulating load is also part of the original design. However, reasons are given to explain the rejection of this loop. For both steps, technical procedures such as filtering, process identification and controller tuning are detailed. Finally, the operator’s interface is briefly described.

Jan 1, 1997

By R. Loj

Katowice Coal Holding (KHW) is one of the largest coal companies in Europe. The company owns five coal mines with a daily output of around 60 000 tons of coal. In the Katowice city area, very good quality steam coal is extracted. The depth of mining varies between 400 and 1100 m and some of the mines rank amongst those most afflicted by methane in the Upper Silesian Coal Basin, with the methane hazard having a significant impact on production and safety. ?Myslowice-Wesola? mine, being a part of the Katowice Coal Holding is characterized by high methane and rockburst hazards. The mine is equipped with a modern surface drainage pumping station. At the mentioned mine, longwall 514 was selected for analysis of its methane emission. The longwall was extracting the thick seam 501 (over 4 m), at a depth of over 900 m, which was subject to high stress and high gas concentration conditions. At a distance of about 50 m above, seam 501 is covered by thin seam 416. The strata between these two seams are sandstone and mudstone layers. In the floor of seam 501, with a thin shale interlayer, is located seam 510, which is the thickest (over 10m) and richest coal seam in the Upper Silesian Coal Basin. In longwall 514 and it?s vicinity, methane was the dominant risk. For this longwall three methods of classical gas drainage were carried out. At first, drainage was undertaken using long boreholes from a cross-cut at level 665 m, located far from the production panel. Later, it was performed using typical boreholes drilled from the ventilation gallery (tail gate road). Finally, it was undertaken using a degassing drift executed in the upper coal seam 416. All of these headings, ie. cross-cut level 665 m, the ventilation road and conveyor road in seam 501 and the degassing drift in seam 416, were equipped with media for drilling and proper drainage pipelines. Due to the high rockburst hazard and some problems with caving, several kinds of de-stressing and torpedo blasts were carried out in both coal seams and roof rocks. The subsequent analysis and experiments allowed the establishment of some general rules for carrying out degassing operations effectively. The effectiveness of a degassing drift was demonstrated. In some instances, the system allowed more than 70% of the total methane released to be captured and fed into pipelines. The aforementioned blasts had a marginal impact on methane emissions and the efficiency of degassing methods. Cyclic drilling and regular switching of boreholes in phase with the mining progress and working face advance were shown to be essential.

Aug 1, 2013

By J. H. Parliament

"The 15,000-tpd Similkameen open-pit copper mine, 10 miles south of Princeton, B.C., started production during 1972. The two main ore zones are separated by the deep canyon of the Similkameen River, with the Ingerbelle or western deposit being mined first. The location, design and construction of the plant were complicated by the steep terrain and the situation of the orebodies. Mining, utilizing 10-cu.yd electric shovels and 100-ton electric-wheel trucks, is conventional. Closely integrated with the preproduction mining was the construction of a 4-mile section of main highway to bypass the Ingerbelle pit. This entailed a very large fill, placed with mine equipment. Ore is crushed to minus 9 inches, stockpiled and reclaimed for direct delivery to three 32-ft-diameter by 14• ft single-stage autogenous grinding mills, each driven by two 4000-hp motors. Because of the friable ore in the upper part of the mine, 4-inch balls are added, with higher than design throughput attained. Classification is by 91>-inch spiral classifiers and cyclones. The flotation section comprises three parallel lines of ten 400-cu.ft rougher- scavenger cells and a single line of 100-cu.ft cleaner and recleaner cells. Concentrate is transported in end dump, closed trucks to Vancouver for dockside storage prior to shipment to the smelter. Tailings are piped across the canyon on a suspension bridge and through a tunnel to an empoundment area. Water is reclaimed and returned to the concentrator, forming a closed system. Dams are raised with cycloned ands, with the downstream section compacted."

Jan 1, 1973

By B. F. Giyani, D. Rahal

"Stirred media mills have proved to be more energy-efficient than the conventional ball milling for fine grinding applications. The mill performance is governed by a number of operating variables including mill stirrer speed, feed particle size, solids content and slurry rheology. A study was conducted to assess effects of mill speed, media size and density, and slurry density on the Stress Intensity (SI) in a VXPmill and thereby the relationship to mill performance. For the mill speeds tested, it was found that the optimum tip speed is 7 m/s. If the speed is too high (12 m/s) energy losses result due to mechanical friction losses and heat generation. If the speed is too low (3 m/s), the kinetic energy is not sufficient to promote breakage. The size reduction ratio increases with the increase in mill speed and the decrease in solids content especially in the coarse fraction particle sizes. The effect of slurry rheology on mill performance was shown by assessing the effects of increasing solids concentration and decreasing particle size. Rheological effects seemed to interact with mill speed such that the mill was less energy-efficient at higher solids content and at higher mill speeds particularly for grinding to fine particle sizes. There is an opti"

Jan 1, 2017

By H. B. Merlin

"The past demands for energy and the contributions by the major sources are shown for the world and regions, and for Canada. Three forecasts of future demand are presented. By the year 2000, the world will be consuming energy at a rate of 670-700 x 1015 Btu/year. Canada's rate is estimated at 20 x 1015 Btu/year. Electricity will increase its portion of energy from 7.5 per cent to 15 per cent, and more than 40 per cent of the primary energy w1ll. be required to generate 1t. World, regional and Canad1an forecasts are g1ven for total electricity generated and for electricity generated in nuclear stations. The integral world requirement for energy by the year 2000 will be over 12 x 10 18 Btu. Current recoverable fossil and hydro reserves are about 18 x 1018 Btu. The total reserves are estimated to be seven times greater.World production of electricity in nuclear stations by the year 2000 will be almost three times the electricity produced in the world today. The amount of fuel consumed in generating this electricity will depend upon the types and mix of reactors installed. The heavy water reactors, alone or in conjunction with fast breeders and liquid fuel reactors, are the most efficient users of uranium and thorium resources. The heavy water reactors are also the most efficient producers of plutonium which will be required as fuel for fast breeders."

Jan 1, 1969

By Arthur L. de Carle

"IntroductionThe Key Lake uranium-nickel deposits are located within a Mineral Lease centred at latitude 57 ° 12' north and longitude 105 °39' west along the southeastern rim of the Athabasca sedimentary basin in northern Saskatchewan, approximately 240 km north of La Ronge (Fig. 1). Access is provided by an allweather provincial highway from Pinchousc, situated about 200 km to the south, and a private airstrip maintained at the mincsitc.These deposits have been developed by Key Lake Mining Corporation on behalf of the Saskatchewan Uranium Joint Venture. Participants in the joint venture include Saskatchewan Mining Development Corporation (50%), Urancrz Exploration and Mining Limited (33 1/3 %) and Eldor Resources Limited (16 2/3 %).Considerable geological work has been conducted on the Key Lake orebodies and reported in summary papers published by Dahlkamp and Thn, (1977), Dahlkamp, (1978), Gatzweiler eta/., (1979) and Kirchner el a/. , (1980). However, intensive geological research is ongoing, particularly since many problems regarding ore genesis and controls remain unresolved . Exploration History and DevelopmentThe events (Gatzweiler eta/., 1980) which led to the discovery of the Key Lake ore bodies began in the summer of 1971 by following up on the reports of airborne radiometric anomalies and anomalous radon values in the Zimmer Lake water situated approximately 5 km to 6 km to the southwest. This work succeeded in locating a few high-grade uranium and nickel-bearing boulders associated with esker material. An integrated program, consisting of various geochemical and geophysical surveys along with diamond drilling, culminated in the discovery of the Gaertner orebody during July 1975. Subsequent fence drilling across the strike extensions of this ore structure located the Deilmann deposit, approximately I km to the northeast, in June 1976"

Jan 1, 1986

By T. Strong

"Selectively recovering pentlandite from high levels of pyrrhotite presents a technical and operational challenge. The liberation characteristics and also the chemical environment is critical to performing a successful separation. Reviewing the mineralogy to understand the characteristics of the ore and the plant flotation behavior is fundamental to designing a successful improvement program.At Savannah Nickel the challenges presented by a falling nickel price required a rethink of the flowsheet to improve the project economics through improving the concentrate grade and recoveries. Glencore Technology reviewed the plant performance and proposed and tested solutions to improve the grade performance of the concentrator with the support of the plant team. A pilot plant comprising a M20 IsamillTM followed by an L500/1 Jameson Cell was operated on various plant streams. In addition to the pilot plant, flotation surveys and a plant based laboratory program was used to supplement the plant understanding and within two weeks significant improvements were identified and demonstrated. The combination of the pilot operation and the plant based laboratory program was able to quickly demonstrate the improvements and test with equipment that is fully scaleable from pilot plant to full scale.IntroductionPanoramic Resources Limited is a Western Australian mining company formed in 2001 for the purpose of developing the Savannah Nickel Project in the East Kimberley region of Western Australia. Panoramic successfully commissioned the $65 million Savannah Project in late 2004. The process plant at Savannah comprises a single stage crusher, SAG mill, flotation, thickening and filtering stages to produce a bulk nickel, copper, cobalt concentrate. Metallurgical recoveries average 86-89% for nickel, 94-97% for copper and 89-92% for cobalt. The plant was originally designed for a throughput of 750,000 tonnes per annum, but has consistently outperformed the design specifications.Due to low nickel prices the Savannah Nickel mine and processing plant were placed into care and maintenance in May 2016. Prior to suspending production Glencore Technology was engaged to investigate options to improve the concentrate grade. With the discovery of Savannah North in 2014 and the Savannah North Scoping Study demonstrating LoM ~8 years and Mining Inventory of 6.07Mt @ 1.26% Ni, 0.64% Cu and 0.09% Co significant future potential exists. By completing pilot plant studies prior to shutting down, the plant engineering studies can be completed and the plant readied for restart by the time the nickel price improves."

Jan 1, 2017