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By R. M. de Carvalho

Empirical and phenomenological models of SAG and AG mills have dominated the scene for the last few decades, being used nearly every time a mill is scale-up or an attempt is made to improve its performance. However, in spite of their success, these models fail to describe the details of both the grinding environment and the breakage mechanisms. The result is a generally good track record, with a few – but very important – instances when these models fall short in their predictions, which can be very costly. Further, nearly all of these models require detailed calibration to industrial and pilot plant data, being costly in their application, although still with limited predictive capabilities. Recognizing these challenges, a new model framework has been proposed which overcomes the limitations of currently used models by coupling the microscale population balance model, the discrete element model and functions that describe internal classification at the mill grate. This model framework has already been applied to describe breakage in ball mills, with showing very promising results. However, in the case of SAG and AG mills a much greater level of complexity is involved in the model, given the interaction between the mechanical environment that is formed totally or, in part, by the ore itself, and the breakage response in the mill. The paper describes the model framework, demonstrating how it can address appropriately effects such as mill speed, mill filling and size distribution, on the response from the mill. Some of the advantages of the model are related to its intrinsic multi-component and dynamic nature, which allows simulating appropriately mixtures of ores of different competences in a transient manner.

Jan 1, 2012

By D. Saramak, T. Gawenda, Z. Naziemiec, A. Saramak

"Crushing and grinding operations are burdensome to the environment due to generating large amounts of dust, noise and waste during course of comminution operations. An indirect negative contribution of industrial crushing processes is an increase of carbon dioxide emissions, which is a byproduct in energy production processes. On the other hand comminution processes in ore enrichment technology are essential due to liberation the useful minerals from the gangue, and producing the demanded particle size fractions of products. The article concerns investigation over negative influence of comminution devices on environment in terms of dust and noise emissions. Several various crushing devices were under investigations, both in laboratory and in industrial scale.The obtained results were analyzed in terms throughput and energy consumption achieved in various crushing devices. The test results allow to conclude that an intensity of dust generation for individual types of crushing devices is varied. An additional aim of this article is to link the aspect of mineral processing with its environmental impact and hence an influence of raw materials beneficiation industry on the quality of life of residents in immediate and close vicinity.INTRODUCTIONMineral raw materials beneficiation processes are an integral part of multi-stage processing of coal, iron ore and non-ferrous metals, production of aggregates, cement, or precious metals and stones. One of the common features of all beneficiation operations is their high energy consumption, which is mainly connected with various operations of feed particle size reduction. Comminution processes in ore enrichment technology are essential to expose and liberate the useful minerals from the gangue, while in aggregate production processes – to obtain the demanded particle size fractions of products. Crushing operations are characterized by a high degree of negative impact to the environment and humans, mainly due to:• High electricity consumption, which indirectly contributes to the increased emissions of carbon dioxide into the atmosphere, since CO2 is often one of the byproducts of energy production;• Relatively rapid wear of device working parts, what forces a production of spare parts and involvement of steel industry;• Intense dust pollution generated during crushing and grinding processes. This is one of the factors affecting the air quality and the consequent deterioration of the living conditions of inhabitants near mines and processing plants; and • Excessive noise emissions."

Jan 1, 2016

By Qing-hua Tian, Xue-yi Guo, Fei Li

A novel process for the preparation of a-Al2O3 from secondary aluminum dross was studied. Factors including alkali dross ratio, salt dross ratio, smelting temperature, smelting time, leaching temperature, leaching time, and the solid-liquid ratio, which possibly influence the transfer of Al into water were studied, and the feasibility of using seed precipitation as a treatment technology of the leachate was discussed. The optimum condition of the process was determined as: alkali dross ratio 1.3, salt dross ratio 0.7, smelting temperature 500°C, smelting time 60min, leaching temperature 60°C, leaching time 30min, solid-liquid ratio 1:4. The results show that using the seed precipitation method, the leaching of aluminum reaches 92.71%, demonstrating the technique as a feasible and effective treatment technology of the leachate.

Jan 1, 2011

By J. Cui, Z. Meng, W. Zhao

"The decomposition of rare earth concentrates containing bastnasite and monazite by alkali solutions has been investigated. A slurry of the rare earth concentrate and the alkali solutions was calcined at different temperatures. The impacts of calcination temperature on the decomposition of rare earth minerals, the oxidation of cerium, the deportment of fluorine and phosphorous, and the adaptability of the method to the different grades of rare earth concentrate were all studied. A possible technique for comprehensive recovery is proposed in which the chemical elements such as rare earth, fluorine, phosphorus, calcium, iron and thorium could be recycled or recovered. INTRODUCTIONThe liquid alkali decomposition method is an important technique for the decomposition of mixed rare earth concentrates. The technique is relatively clean compared with the sulfuric acid roasting method. However, its operation is intermittent rather than continuous. The two-step solid-liquid reaction and the subsequent water washing make it not very suitable for mass production and application. An important reason for the disadvantage is that an additional process of ""calcium removal by acid leaching"" is needed for the technique. The content of calcium in Baotou rare earth concentrates is usually 6-12%. The requirement of the current technique is that the content of calcium should be reduced to about 0.2% after the calcium removal process. Otherwise, the decomposition and leaching of rare earth will be affected. Therefore, calcium removal by acid leaching is a necessity before the alkali decomposition process is carried out. Research on different alkali decomposition techniques has been conducted with prior calcium removal by acid leaching (Xu, 1995; Yu, 1993; Han, Li, Chang, Liu & Xu, 1985). There have been some investigations of direct decomposition of the bastnasite (Feng, Pan & Li, 1998; Zhao, Zhang, Yang & Liu, 2002)and monazite (Sun, Li, Li, Su, Zhao & Liu, 1998; Zhou, 2005) by alkali-liquid, however, all studies were carried out in solid-liquid reaction system, and continuous calcination studies have not been reported.This paper investigates the process of direct decomposition of Baotou rare earth concentrates by continuous calcination with concentrated alkali. Using the rotary tubular electric furnace, the mixture of rare earth concentrate and concentrated alkali solution were decomposed continuously. Satisfactory decomposition results were achieved. A method for clean processing and comprehensive resource recovery from Baotou rare earth concentrates is proposed."

Jan 1, 2012

By Pierre Sauvé, Alex Makila

"The Camflo stock of porphyritic quartz syenite is an irregular pipe-like body 200 metres in diameter injected along the axial plane of an E-W trending F2 fold, near the contact between the. Kewagama and Malartic groups. The stock plunges 55° to the NE, almost parallel to the fold axis. The stock, exhibiting pervasive, weak cataclasis with numerous, locally auriferous quartz veinlets, is cut by minor faults striking WNW and dipping about 45°NE.Two types of gold ore are distinguished at Camflo: ""porphyry"" ore and ""fault zone"" ore. ""Porphyry"" ore, accounting for 90% of the gold produced, forms wide zones within the stock. The zones are irregular in shape but tend to be oriented E-W in plan view (on a horizontal plane) and are elongated paraliel to the plunge of the stock. These zones are more abundant in the northern part of the stock. Differences between ore and barren porphyry are not obvious. ""Fault zone"" ore occurs outside the stock along minor faults, particularly in differentiated sills offer rodiorite and felsic granophyre. Gold grade is roughly proportional to the pyrite content which in turns depends in part on the original magnetite content of the host rock. Although both types of ore show some correlation between Au and S, the Au/S ratio is considerably higher in porphyry ore.The bulk of the gold is in quartz syenite and in felsic granophyre, the most brittle lithologies at the mine. Presumably, the vertically continuous, pipe-like stock, made up of shattered rock, provided the main channelway for the ascending mineralizing solutions."

Jan 1, 1990

By S. Gheibi, H. S. B. Duzgun, N. Dadashzadeh

It is widely recognized that there are various uncertainties in the analysis of rock slope stability owing to inadequate information for site characterization and inherent variability and measurement errors in geological and corresponding parameters. Therefore, reliability-based approaches that allow the systematic and quantitative treatment of these uncertainties have become a topic of increasing interest in rock slope engineering. There has been extensive reliability analysis of rock slope stability in the literature. However, all of these analyses are based on the First-Order Reliability Method (FORM). The probability of failure of a rock slope in the First-Order Reliability Method is provided by a linear approximation of the failure function. The FORM is most commonly used because it is efficient; its accuracy, however, deteriorates when the nonlinearity of limit-state function increases. A better approximation of the failure function can establish a more accurate value of the probability of failure. The Second-Order Reliability Method (SORM) overcomes this drawback with a cost of lower efficiency in terms of the number of function calls and computation expenses. Despite considerable studies of the application of the SORM in other engineering fields, a few attempts have been made in the slope stability studies. In this study, a recently published Second-Order Reliability Method with First-Order Efficiency (SORM-FOE) is implemented on a selected rock slope in Amasya, Turkey. The SORM-FOE overcomes the low efficiency of the SORM by decreasing the number of function calls. The probability of failure of both FORM and SORM-FOE are obtained and the results are compared. It is shown that, in high value of failure probability, SORM’s do not significantly affect the analysis and design. However, in very low failure probability, the discrepancy between FORM and SORM-FOE becomes pronounced. It is also shown that, where the second order methods are required, the SORM-FOE is more efficient than the conventional curvature fitting methods in terms of computational costs.

Jan 1, 2015

By Ki-Bok Min, Bona Park

"In spite of the applications of DEM models to emulate the elastic modulus, Poisson’s ratio and strengths of isotropic rock in laboratory scale, extensions to anisotropic rock and larger engineering scale still have many challenges to overcome such as robust verification and validation against observed field case. In this study, a discrete element modeling was performed in order to represent a transversely isotropic rock such as shale, and the upscaled model was applied to foundation and borehole problems using PFC2D. The applicability of the upscaled rock mass model was verified by an analytical solution. The DEM modeling of stress redistribution induced by an applied surface line load compares well with Boussinesq’s isotropic solution and anisotropic solution given by Bray. The borehole model with a circular opening under biaxial stress was compared with both isotropic and anisotropic solutions resulting in good agreement. The abilities of DEM model to investigate the micromechanisms accounting for complex macroscopic behaviors and to predict the macroscopic behavior are discussed.The results demonstrate that DEM modeling has the potential to be applied to anisotropic rock mechanics problems on larger engineering scales such as wellbore stability problems especially in shale formations. Furthermore, the effect of anisotropy modeled through this study may be able to explain failure behavior of anisotropic rock encountered frequently in real cases.INTRODUCTIONAnisotropy is prevalent in various types of rocks, especially in shale that contain a set of parallel bedding planes. These weak planes in various forms are sources of large deformation and low shear strength and contribute significantly to the anisotropic behavior of rocks (Amadei, 1996; Cho et al., 2012; Kim et al., 2012). As elastic and strength behavior of anisotropic rock is substantially different from that of isotropic rock, proper consideration of anisotropy is of great importance. There is extensive of experimental evidence showing the anisotropy of shale. Niandou et al. (Niandou et al., 1997) analyzed the plastic deformation and failure mechanisms of Tournemire shale were highly dependent on the orientation of bedding planes. Cho et al. (Cho et al., 2012) presented various deformations and strengths of Boryeong shale obtained from laboratory experiments. Fjaer and Nes (Fjaer and Nes, 2013) performed uniaxial and triaxial compressive strength tests using Mancos shale, and they observed that the shear sliding along the planes of weakness before the occurrence of a major failure can reduce the strength of anisotropic rock. The recent research has demonstrated that the anisotropy of rock is an important consideration in rock engineering."

Jan 1, 2015

By W. G. Spencer, E. Sulis, P. M. Tiley

"Dome Mine's No.8 shaft is located about 1 km northeast of Dome's No.3 shaft and mill complex in South Porcupine, Ontario.No.8 shaft was brought into production in 1984 to hoist ore and waste for the complete mine as well as to service the mine from the 1600 level and below. Access above 1600 level is provided by No. 3 shaft.The rectangular shaft has six compartments consisting of two skips, a cage, counterweight, manway, and a service compartment. The over-all size of the shaft is 14 ft 7 in. by 20 ft by 7 in. (Fig. 1). A 7th compartment for shaft deepening is incorporated in the bottom 500 ft of shaft. The shaft is 5470 ft deep. The loading pocket is located at 5300 ft and the dump 61 ft below the collar.Surface facilities consist of a concrete headframe and a structural steel hoistroom which houses the cage and production hoists . The service hoist is a 12-ft diameter double drum, with one 1750 HP DC motor, and the production hoist is a 15-ft diameter double drum with twin 1750 HP DC motors.Both hoists have a single clutch and a single reduction gear train built by Orenstein & Koppel, and equipped with Canadian General Electric drives. The cage hoist is equipped with a disc pinion brake.The skips were originally aluminum 14-ton capacity, Rolla-chute design with steel liners on three sides and 4 in. of rubber on the back, to absorb the loading impact. The original guide rollers consisted of fixed 10 in. pneumatic tires. The face-to-face units were eventually replaced with larger airplane-type tires mounted on a spring loaded pivot arm."

Jan 1, 1993

By Colin Harvey

The paper industry in the Pacific Northwest is a potentially large market for industrial clays. Current trends within the paper industry towards higher quality newsprint, filled and coated papers, offers an incentive for exploration in the region since the current supply of kaolin is mainly from Georgia in the southeastern U.S. Freight rates commonly exceed US$100 per tonne. The kaolin resources of the Pacific Northwest require expensive processing technologies to meet the industry specifications and it is doubtful that any large development of the local resources can be supported at the present time. Competition from high-quality regional resources of carbonates is another factor weighing against the development of local kaolin resources while there is a potential threat from shipments ofAsian and Brazilian kaolin in the early part of the 21st century.

Jan 1, 2001

By P. B. Demarsh

"Location: - Mine village of Perron, 11 miles east of Val d’Or in northwestern Quebec.Life; - Started production 1932, ceased 1951.Ore: - Flat lying quartz tourmaline veins within granodiorite. Gold values related to coarse segregations of pyrite. Specific gravity 2. 80.Capacity: - 30 tons 1932-35; 125 tons 1935-37; 350 tons 1937-51.Assay: - Heads 0.21 to 0.34 oz. per ton. Grade obtained by mining 600 tons daily, sorting belt reject 250 tons, balance to milling; tailings loss 0.003 to 0.005 oz. per ton; extraction 9 8. 5%.Grind: - 70% minus 200 mesh. Detention time 26 hours.Treatment: - Cyanidation process.Reagents: - Per ton, aero cyanide (NaCn) 0.36; unslaked lump lime 1.50; zinc 0. 03; lead acetate 0. 006; steel consumption 2. 50.Water: - Equal mixture of swamp creek water and underground seepage.Cost: - Per ton, sorting and crushing 0.21; grinding 0.31; cyanidation 0.11; filtration 0.05; precipitation 0.02; miscellaneous 0.22; total 0.92.Power: - Hydro-electric, 25 cycle, 12, 000 volts reduced to 550 volts.Labour: - 95% Maritime and European, balance Quebec and Ontario.Tailings: - Absorbed over large swampy area on barren mining claims.Personnel: - 1 7 men including supervision.Flow sheet: - Canadian Mining Manual 1947"

Jan 1, 1966

By E. W. Garner

THE greatest problem confronting business managers in modern times is not production, but distribution. It would be easy to expand productive output, but it is difficult to find markets for all that is produced. The fundamental source of economic difficulty, in the view of many observers, is a lack of proper balance between productive capacity and consumptive capacity. The studies conducted by the Brookings Institution on the general subject, "The Distribution of Income in Relation to Economic Progress", have led to conclusions of far-reaching significance, which may be summarized briefly as follows: (1) We have never lived beyond our economic means or capacity. (2) We have not been suffering from general overproduction. (3) We have not as yet reached a stage in our economic evolution at which it is possible to provide adequate standards of living for everybody. ( 4) It would require less than a 25 per cent increase in purchasing power among the masses to absorb our full productive capacity. (5) Production-curtailment programmes-whatever their merit in meeting temporary maladjustments-can only lead to national impoverishment. (6) Our inability to find markets adequate to absorb the full output of our productive establishments is explained by the unequal distribution of income and consequent restricted flow of purchasing power through consumptive channels. (7) The slow rate of growth of consumptive demand serves to retard the rate at which new capital is constructed, and hence the rate of economic progress. (8) The primary need is a rapidly expanding total income, and a broader diffusion of this income among the masses of the people. (9) A broader diffusion of income can best be achieved by reducing prices as rapidly as increasing efficiency will permit. That is to say, the benefits of technological progress, mass production, and low unit costs must be promptly passed on to consumers in the form of lower prices-giving more for each dollar.

Jan 1, 1936

By K. Regl, J. Berneder

"AMAG investigated different aspects to improve the resistance against grain coarsening in aerospace alloys during recrystallization and was successful in finding several opportunities to minimize this effect. The studies show that one aspect to counteract grain coarsening is the chemical composition and subsequently all phases that are formed during homogenization. Starting with a fine grain (~ 50 µm) can lead to a final grain size of few mm. On the other hand with a coarse initial grain size (approximately 150 µm) and the right chemical composition the grain size can stay almost constant despite different deformation steps. Additional we investigated different sequences for the component production to eliminate the problem of grain coarsening as a coarse grain has a detrimental effect on product performance.INTRODUCTION In the aircraft industry, aluminium sheets are most commonly deformed in their soft state, followed by solution heat treatment and natural or artificial aging. If the preceding deformation is in a critical amount, during the recrystallization at solution heat treatment grain coarsening can occur. A subsequent deformation to this condition leads to an orange peel surface, which requires additional polishing. Next to this extra effort the coarse grain size produces a drop in the mechanical properties, often below specification limits. If there is no subsequent deformation, this effect is not visible and without destructive testing there is no possibility to detect the error. As mentioned before, this effect occurs during recrystallization, which is a process that is driven by the stored energy in the microstructure. The two most important aspects for an increase of the stored energy are the increase of the grain boundaries fraction (smaller grain size) and the increase of the dislocation density caused by deformation. The increase in dislocation density during the deformation arise from trapping of existing dislocations and from the generation of new dislocations. During recrystallization all grains are built completely new, so all dislocations are removed (Humphreys & Hatherly, 2004). There are three requirements for recrystallization:"

Jan 1, 2018

By Linlin Zhang

Rare earth elements (REEs) are a group of lanthanides elements which have been widely used in the electronic, petrochemistry, metallurgy, and defense industries. Their demand has increased sharply in the past decade with the boom in electronic device production. Some countries now consider them strategic metals for the future and it has been forecasted that REEs may become a dominant contributor to global economy. In the wake of REE exploration, mining and processing, concern has grown over occupational and community health issues and risks. A portion of available literature on this topic however is written in the Chinese language, and thus research information related to health and REE mining is not readily accessible to mineral development planners in countries such as Canada, where REE mining is now evolving. As REEs also represent an important commodity, primarily stemming from Chinese production, it is important to begin to understand the potential health costs related to this mining sector. There are two main objectives of this paper. The first is to present an overview of REE mining with a focus on geological characteristics, global production and uses with particular attention to the main producer: China. The second is to present the results of a literature review (using both Chinese and English sources) on REE mining and health and identifies key occupational and community health issues that require consideration. This paper concludes by identifying some critical issues this sector needs to address to ensure the sustainable development of this mining activity.

Aug 1, 2013

By F. P. Hassani, Jacque Ouellet, M. Hossein

A laboratory investigation was undertaken to evaluate the effectiveness of different binding agents in a paste backfilling operation suitable to be used in a sulphide-rich environment. Different combinations of Portland cement and fly ash type C as well as ferrous slag were used as the binding agents. The unconfined compressive strength of the samples was measured for extended curing periods. The results indicate that the addition of 3% slag in the binding agent mix could produce strengths up to 0.5 MPa in the samples under investigation. For higher binding contents, the addition of Portland cement will lead to higher availability of calcium hydroxide and calcium aluminate, which are susceptible to deterioration from the phenomenon known as sulphate attack. Using slag, on the other hand, will reduce the permeability, availability of calcium hydroxide, and dilute the calcium aluminate ions reducing the susceptibility of the samples to sulphate attack. However, gypsum will form which could have the same detrimental effect on the strength of the treated matrix. Therefore, special attention must be given to the content and properties of agents used to activate the pozzolanic properties of the slag.

Jan 1, 2001

By M-A. Thibault, C. Gauthierand, J-P. Gagné, J-Y. Carrier, P. Rochette, G. Dufour

"Anodes consumed in the electrolytic process for the production of primary aluminium are considered consumables. They need to be replaced at regular intervals of approximately one month. The so-called anode butts which are removed from the cells contain the metal supports (rods), cast iron thimbles, crust and residual carbon material. Cleaning is required to remove the waste material before the rods are fitted with new anodes. An inspection is required to ascertain if repairs to the rods are necessary. Presently, this inspection is carried out visually. ST AS, in 2004 in collaboration with Alcoa Deschambault, initiated a project to develop an automated inspection system. The objectives were to find a technical solution to perform these functions with a very simple system. Such a system was designed around ""artificial vision"". The inspection system developed uses only one fixed camera and does not require a separate step in the process for the anode rod assembly to perform the inspection. This paper presents the initial steps of the work performed to develop the equipment and software as well as a summary of the results obtained."

Jan 1, 2006

By J. P. Nowlan

"This paper is an attempt to outline the structural problems at the Cochenour Willans mine, in so far as their details have been tentatively worked out. It must be emphasized that there are considerable data that can be interpreted in more than one way, and that further work may modify or completely change some of the hypotheses presented herein. Certain features, however, are sufficiently constant to warrant reliance on them when searching for extensions of orebodies.As the individual ore shoots usually have relatively small dimensions, but are often very high grade, structural details are of more than usual significance in laying out a mining programme. An illustration of this is a small shoot recently located and mined as a result of studying a minor slip cutting off a part of high grade gold ore in No. 319 stope. The shoot, which is bounded on three sides by slips, and on the fourth by the original stoping, was roughly triangular in plan, 20 feet on the base, and 40 feet in length. The average width mined was 4.5 feet, but the gold obtained totalled over 1,800 oz.The property lies in Dome township, east of McKenzie island, and consists of an original group of thirteen claims acquired from the Cochenour Willans syndicate, plus one additional group of five claims obtained by purchase from the Kelson Red Lake Gold Mines, Limited, and one other claim that rounds out the southeast corner of the block.The claims of the original group were staked by W. M. Cochenour, Dan Willans, and Horace Young, during 1926 and 1927. The Cochenour Willans Syndicate was formed in 1928, and until incorporation of the present Company work was carried on at varying intervals by several different interests, including Ventures Limited, Hollinger Consolidated, who sank a shaft to open up the 150-foot level, and by the Syndicate itself. The first ore from the property was shipped to the Gold Eagle mill by the present Company in 1939, and amounted to 4,959 tons, from which 2,360 oz. gold was recovered. The first brick was poured in December, 1939, and production has been continuous since that date. To the end of June, 1946, gold having a value of $5,982,000 has been produced from 345,033 tons of ore treated."

Jan 1, 1954

By F. J. West

This paper deals with Jones & Laughlin Steel Corporation's new Adams mine, near Kirkland Lake, Ontario. After a brief discussion of ore structure, type, reserve tonnages, etc., the main body of the paper discusses the establishment of an open-pit limit based on economic considerations. The method used by J & L to arrive at an economically sound open-pit boundary its outlined fully. Physical limitations to open-pit boundaries are mentioned briefly. The paper concludes with a few remarks on the mining equipment in use and on its productivity

Jan 1, 1966

By P. A. Geusebroek, N. A. Duke

Abstract - The Lupin mine, located in the central Slave province just east of the western boundary of Nunavut Territory, is a world-class example of a Neoarchean-aged banded iron formation (BIF)-hosted lode-gold deposit. At the minesite the gold-mineralized Lupin BIF, separating stratigraphically underlying psammitic wacke and overlying argillaceous turbidite sequences, delineates the Lupin dome, a hammerhead-shaped F2/F3 interference fold structure occurring at the greenschist to amphibolite facies metamorphic transition within the thermal aureole of the Contwoyto batholith. Detailed paragenetic relationships indicate that peak thermal metamorphism coincided with the switch from regional D2 compression to rapid D3 unroofing of the Neoarchean orogenic infrastructure. Gold initially precipitated with pyrrhotite, replacing amphibolitic BIF at the apex of the Lupin deformation zone, separating the east and west lobes of the Contwoyto batholith. Over the course of associated prograde/retrograde metasomatic overprints, gold was further remobilized during garnet and loellingite/arsenopyrite growth in chlorite-altered selvages of late-forming ladder quartz veins. A metamorphic model of ore genesis, with gold being scavenged and transported by metamorphic fluid that was shed and structurally trapped at the amphibolite recrystallization front, is favored over the previously proposed syngenetic and exogenic models of gold concentration that have tended to polarize genetic interpretations to date.

Jan 1, 2004

By S. McPherson, M. Chalkley, R. Ellenwood, W. Nicot

"The Moa JV was formed in December 1994 as a 50/50 joint venture between Sherritt Inc. and General Nickel Company S.A. of Cuba. The joint venture consists of the mine and Pedro Sotto Alba processing plant in Moa, Cuba and the Corefco refinery in Fort Saskatchewan, Alberta, Canada. Throughout its now nearly 23-year history, the Moa Joint Venture has faced a number of global, operational and process challenges. Global challenges have included high input costs such as for fuel oil and sulphur, as well as low nickel prices. Operational and process challenges have included increasing ore haulage distances, decreasing ore Ni + Co grades, and increasing ore impurity contents. For example, increases in impurities such as Mg and SiO2 in the ore result in increasing acid demand in leaching and decreasing solids content of preheated leach feed slurry, respectively. In spite of these challenges, the mixed sulphide production capacity from the Moa Joint Venture has increased from 12,500 to 37,000 t/y Ni + Co since 1994 by judiciously targeting expansion projects and by continuously performing process and operational improvements.INTRODUCTION The Pedro Sotto Alba high pressure acid leach (HPAL) plant at Moa Bay, Cuba has a long history (Maslenitzkiy, 1969; Chalkley, 1996; Carlson, 1960; Roy, 1961; Carlson 1961). The plant was constructed by Moa Bay Mining Company and started up in late 1959 after several years of process development by Freeport Nickel Company, including operation of a 10 t/d plant in Hoskins Mound, Texas and a 50 t/d plant in Port Nickel, Louisiana (Southwick, 2004), as well as development work by Chemical Construction Company (Chemico) to precipitate metal sulphides (Roy, 1955a; Roy, 1955b). The Moa plant had an original reserve of 50,000,000 t of ore, graded at 1.35 wt% Ni, 0.146 wt% Co, 3.70 wt% Si, 1.00 wt% Mg and 4.5 wt% Al. At original design rates of 2,000 kt/y of ore, the design capacity of the plant was around 25,000 t/y Ni + Co as a mixed sulphide product, which would be shipped to the Port Nickel Refinery in Louisiana for processing. In 1960, the plant changed ownership and was restarted in 1961 under Cuban management. Between 1961 and 1989, the plant was also supported by involvement from the former USSR, supplying spare parts and consumables. In 1994, Sherritt Inc. partnered with the General Nickel Co. S.A. of Cuba and formed the Moa Joint Venture. This joint venture includes the mine and Pedro Sotto Alba HPAL processing plant in Cuba and the Corefco refinery in Fort Saskatchewan, Alberta, Canada. Under this partnership, ore is processed in Cuba to produce a nominally 55 wt% Ni and 5 wt% Co mixed sulphide product that is shipped to Canada and processed by Corefco to produce class 1 nickel, high grade cobalt and byproduct ammonium sulphate fertilizer."

Jan 1, 2017

By K. E. Waters, C. Marion, T. Grammatikopoulos, A. Jordens

"The Nechalacho deposit, located in the Northwest Territories, Canada, is a heavy rare earth element (REE) deposit. The main REE-bearing minerals in the deposit are zircon, allanite, bastnäsite, synchysite, columbite (Fe) and fergusonite. Previous work investigated the use of gravity and magnetic separation in the processing of this ore and showed that coarse particles following grinding were enriched in zircon, suggesting the potential for selective comminution. The current work investigates the possibility to selectively comminute gangue minerals and concentrate REE-bearing zircon into coarser size fractions. The ore was ground wet in a laboratory rod mill for various lengths of time (0, 10, 20, 30, and 40 min). The resultant products were analyzed using ICP-MS, to determine the distribution of zirconium and REEs in the different size fractions produced after grinding, and QEMSCAN, to identify mineral phases, mineral liberation and the distribution of minerals. Optimum grind sizes for the upgrading of zircon due to selective comminution were then proposed, and dense medium separation was investigated as a means to further concentrate zircon.It is important to note that this process is not representative of the currently selected process design or recovery for the Nechalacho deposit. Any application of this process to this deposit would require optimization in order to ensure appropriate grade and recovery targets are met.INTRODUCTIONThe rare earth elements (REEs) are the fifteen elements in the lanthanide series of the Periodic Table, as well as yttrium (scandium is excluded in this work). Traditionally, REEs are categorized into two groups: the cerium sub-group of “light” rare earth elements (LREEs); and the yttrium sub-group of “heavy” rare earth elements (HREEs) (Golev et al., 2014). The LREEs comprise the first six lanthanides (lanthanum through samarium) and the HREEs comprise the remaining nine lanthanide elements (europium through lutetium) coupled with yttrium (Golev et al., 2014). HREEs are much less abundant in the earth’s crust and on average are more expensive than LREEs (Seredin, 2010). Of the fifteen REEs, the production of neodymium, europium, terbium, dysprosium and yttrium has been determined to be most critical due to their scarcity, high demand and criticality in many high-technology and clean energy applications, such as in illuminated screens for electronic devices and permanent magnets in wind turbines (Seredin, 2010; U.S. Department of Energy, 2011). These five elements are classified as critical rare earth elements (CREEs) (U.S. Department of Energy, 2011)."

Jan 1, 2016