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By A. Takala, H. Cline, D. Girard

"This article describes the project development and successful implementation of a particle sorter at the Kensington mine. The Kensington underground gold mine and associated milling facilities are located within the Berners Bay Mining District on the east side of Lynn Canal about 45 miles north-northwest of Juneau, Alaska. The paper describes the project from concept, through the development phase, including test work confirming feasibility, to a discussion of operational data and includes shared experiences from operation and maintenance. The project has proven successful in recovering gold that was previously lost to the waste dump and has already resulted in a positive return on investment. The payback was achieved in less than one year.INTRODUCTION The first paragraphs of this paper describe Coeur Mining and the Kensington Mine, its history, geology, mining operation and mill. The second part of this paper gives a quick introduction to sensor-based sorting in general and X-ray transmission sorting (XRT) in particular. This article does not cover costs or economic impacts apart from the pay-back period for the investment at Kensington. Interested readers find more in information on this in the literature (Kleine & Wotruba, 2010; Robben, 2013). Since the publication about XRT sorting at TGME in Pilgrim’s Rest (Kleine, Riedel, Ketelhodt, & Murray, 2010), this is the first time that a comprehensive data set on gold ore sorting using XRT is published. Following chapters highlight the aim of the project for Coeur Kensington, project development through testing and operation before concluding. ABOUT COEUR MINING’S KENSINGTON OPERATION Coeur Mining is the largest U.S.-based silver producer and a significant gold producer with five precious metals mines in the Americas employing approximately 2,100 people. Coeur produces from its wholly owned operations: the Palmarejo silver-gold mine in Mexico, the San Bartolomé silver mine in Bolivia, the Rochester silver-gold mine in Nevada, the Kensington gold mine in Alaska, and the Wharf gold mine in South Dakota. The Company also has a non-operating interest in the Endeavor mine in Australia in addition to royalties on the Cerro Bayo mine in Chile, the El Gallo complex in Mexico, the Zaruma mine in Ecuador, and the Correnso mine in New Zealand. In addition, the Company has two silver-gold feasibility stage projects—the La Preciosa project in Mexico and the Joaquin project in Argentina. The Company also conducts ongoing exploration activities in Alaska, Argentina, Bolivia, Mexico, and Nevada. The Company owns strategic investment positions in several silver and gold development companies with projects in North and South America."

Jan 1, 2017

By D. W. Zandee

"IntroductionHaving mined an economic mineral bearing rock which is of insufficient grade to warrant direct smelting, upgrading or beneficiation is required to make it an economically viable ore. The first step in effecting the concentration process is comminution or sequential breaking of the rock to allow liberation and subsequent recovery of the desired mineral particles. Desirably the ore could be broken down to the point where the valuable mineral is completely liberated to maximize the chance of recovering as much of it and in as pure a form as possible. The number of energy intensive stages of breakage required to attain this goal depends on many physical characteristics of the ore such as hardness, friability, toughness, grain size, existing fracture planes and particle shapes.Commonly the distinct stages are primary, secondary and often tertiary crushing and finally grinding. To increase the efficiency of each stage of size reduction, pre-treatment sizing is common.Crusher feed is commonly scalped of large pieces or screened for reduction crushers and classified for secondary grindingcircuits.Although crushing and screening is employed for many different products, this chapter reviews current practice in larger tonnage metalliferous ore treatment plants.Current practice is to use large gyratory or jaw crushers for primary crushing and some sort of gyrating or cone crushers for additional stages of reduction. For autogenous or semiautogenous grinding circuits only the primary crushing stage is necessary, the balance of the work being performed in the grinding stage. For conventional grinding circuits using ball mills or combinations of rod and ball mills, secondary, tertiary and even quaternary crushing stages are normally required.Older type crushers such as stamp mills, double rolls, sledging mills and hammermills are no longer common but are used for small tonnage special processes.Important to efficient crushing are a variety of sizing devices such as grizzlies and vibrating screens. Products at different stages must be stored to give surge capacity and allow controlled feeding rates through mechanical and gravity feeders. These storage piles, bins and feeders as well as crushers and screens are commonly integrated into a total facility by means of conveyor systems. Most often these are endless rubber belts driven by simple or multiple drive mechanisms usually consisting of electric motors and speed reducers."

Jan 1, 1989

By Neil Campbell

"The gently-dipping Devonian dolomite beds south of Great Slave lake in northern Canada are nearly every-where concealed by a cover of glacial drift and muskeg. Nevertheless, in 1898 or earlier, four small outcropping areas of dolomite, richly mineralized with lead and zinc sulphides, were discovered. Exploration work performed here up to the end of 1930, included intensive trenching and drilling of the showings, together with mapping and wide-spaced drilling in the surrounding region. This work proved that the discoveries are isolated pockets of mineralization much too small to be commercially valuable by themselves. No evidence of orebodies elsewhere was reported.Meanwhile, mapping by the G.S.C. in the Shield far to the east disclosed major Precambrian faults, some of which can be projected southwesterly (under a cover of Paleozoic sediments) to the vicinity of Pine Point. During the period from 1940 to 1947, consideration was given, by The Consolidated Mining and Smelting Company Ltd., to evidence of a theoretical possibility that, although apparently dormant in Paleozoic times, these faults may have provided a structural control and a source of ore-bearing solutions in some post-Devonian period. From this, it was postulated that the breaks might have created, in the overlying dolomite, a linear array of orebodies concealed by overburden but lying in a belt of considerable length and in a position predictable within practical limits. A program of drilling designed to test and exploit this specific concept was conducted from 1948 to 1955.he drilling within the new, theoretically-defined search area resulted in the discovery of many lead-zinc ore-bodies, none of which was exposed at the surface. Some of these are large and (as indicated by 1965 production records) very rich. They are situated 1 to several miles distant from, and apparently not connected to, the earlier-known surface discoveries. The drilling also disclosed that the ore, which consists of disseminated sphalerite and galena replacements together with colloform sphalerite and coarse galena cavity fillings, occurs mainly at certain stratigraphic horizons within a large, coarsely-recrystallized dolomite barrier reef. Considerable fracturing and folding was found. It has been suggested that reef development was influenced by post-Precambrian tectonic movements associated with the old faults in the basement complex. The ore may have been deposited by solutions moving from distant sources to favourable structural and chemical environments in or related to the reef.In more recent exploration performed in the course of preparing for mining production, the rate of ore discovery within the same general geological context has been greatly accelerated by the use of new geophysical techniques and drilling equipment. Reserves of 21.5 million tons, averaging 4 per cent lead and 7.2 per cent zinc, were reported in 1965, and production from open pits at the rate of approximately 248,000 tons of concentrates per year is in progress"

Jan 1, 1966

By M. Barati, K. Ebrahimi-Nasrabadi, P. W. Scott

"The conversion of cryptocrystalline magnesite to caustic magnesia was studied experimentally at temperatures of 400–1,200°C and for calcination times of up to 8 h. A differential thermal analysis– thermal gravimetric analysis showed that decomposition begins at temperatures as low as 340°C and is completed at 600°C. Specific surface area, iodine adsorption number, crystallite size, and loss on ignition were measured to characterize the product quality. It was found that high-quality caustic calcined magnesia can be produced in a narrow temperature (600–650°C) and time range (1–4 h).RÉSUMÉ La conversion de magnésite cryptocristalline en magnésie caustique a été étudiée de manière expérimentale à des températures de 400 à 1200 °C et pour des temps de calcination atteignant 8 heures. Une analyse thermique différentielle – analyse thermique gravimétrique montre que la décomposition commence à des températures aussi basses que 340 °C et qu’elle est complète à 600 °C. La surface spécifique, l’indice d’adsorption de l’iode, la taille des cristaux et la perte au feu ont été mesurés afin de caractériser la qualité du produit. Il a été trouvé qu’une magnésie caustique calcinée de grande qualité peut être produite dans un petit écart de température (600 à 650 °C) et dans une période de temps variant entre 1 et 4 heures."

Jan 1, 2015

By W. Mueller, M. Rocheleau, E. Dimroth

The paleogeographic evolution of the Abitibi Belt is subdivided into three phases, namely: 1) a phase of the growth of a submarine volcanic chain, 2) a phase of the emergence and growth of volcanic islands and 3) a phase of uplif t and erosion of volcanic islands. Isostatic calculations suggest that the volcanic sequence in the south of the Abitibi Belt rests on oceanic basement or on the mantle; the volcanic sequence in the north of the belt rests on about 13-14 km of tonalitic basement. 10-15 km of plutonic material accreted to the volcanic islands during paleogeographic phases 1 and 2, and caused the uplif t during paleogeographic phase 3.

Jan 1, 1984

By Han Xing, Cui Jianzhong, Qin Ke, Shao Bo, Zhang Haitao

In this paper, a new approach that uses hot-top DC casting process to producing multi-alloy ingots is present. The different type and size of composite and coated ingots (4045/3004/4045 three-layer composite ingots with section 630mm by 1500mm and 4045/3003 coasted ingot with F140mm/F110mm in diameter) are cast by this approach. The macro-morphologies and microstructures of the composite ingots, the temperature distribution and the element distribution in the interface zone were investigated, also and the interface bonding strength was measured. Results show that the solid supporting layer formed on the cooling plates plays a key role in the casting process of composite ingots. The solid supporting layer can prevent the blending of two melts by resisting the impact of alloy melt, which ensures the stable casting process and casting high quality composite ingots, because the contact height is increased using hot-top DC casting. In addition, fabrication and properties of 4045/3004/4045 three-layer composite ingots with section 630mm by 1500mm are reported in detail.

Jan 1, 2015

By H. P. Kang, Y. J. Feng

The hard roof in underground coal mines is generally considered as a main cause that induces disasters in working faces and entries, such as rock burst and windblast. Hydraulic fracturing technology is an effective approach to weaken hard roof to reduce ground pressure and control rock burst. The state and achievements associated with hydraulic fracturing technology applied in Chinese coal mines are summarized in this paper. Theories related to hydraulic fracturing are of the essence in fracture design to determine the fracture initiation and propagation pressure and predict the geometry and the characteristics of hydraulically induced fractures. Initially, theory of elasticity was employed to determine the breakdown pressure for a vertical borehole where the rock medium is assumed to be elastic. However, porosity and pore fluid have an influence on the borehole breakdown pressure, and then the poroelasticity theory was applied. Arbitrary deviated boreholes in practice are rather common and the maximum tensile stress criterion was employed to determine borehole breakdown pressure. The effect of natural crack in the rock medium on the initiation pressure has been taken into consideration through fracture mechanics. Meanwhile, considerable achievements have been made on hydraulic fracturing tools and equipment, and a set of mature hydraulic fracturing technologies for hard roof control are available. Some comprehensive investigations on the hydraulic fracturing for hard roof weakening are presented, and successful field applications in the Jincheng and Luan coal mining districts are introduced. Finally, the unsolved problems related to hydraulic fracturing are also put forward.

Jan 1, 2015

By A. Idegami, M. Imamura, S. Heguri, Y. Ozaki

"Nickel powder can be produced by the reduction of a nickel ammine complex solution by hydrogen at high temperature and pressure. To minimize the deposition of fine nickel on the surface of the reactor and other equipment, seed crystals are added to the initial solution to promote nucleation of the nickel powder. Such seed crystals catalyze the precipitation of the nickel particles on the surface of the seed and enhance grain growth. In this study, nickel nuclei were precipitated by the addition of a polymer to the nickel ammine complex solution, resulting in the production of nickel powder. Thus, we herein propose a model that explains the effect of the polymer in the precipitation of nickel nuclei.INTRODUCTIONNickel powder is commonly produced via a hydrometallurgical process from the hydrogen reduction of a solution containing a nickel ammine complex at high temperatures and pressures, and this process was commercialized in the 1950s by Sherritt Gordon (Sherritt International, Canada, Benson & Colvin,1973; The Staff, Research and Development Division, Sherritt Gordon Mines, Ltd., 1958).In the above transformation, upon the injection of hydrogen gas into the solution, nickel is reduced as outlined in Equation(1), ultimately resulting in the production of nickel powder.Ni(NH3)nSO4+ H2?Ni+ (NH4)2SO4+ (n-2)NH3(n:2–6)(1)In this process, seed crystals are added to the reactor to suppress the precipitation of fine nickel (scaling) and to enhance nickel reduction (Benson & Colvin, 1973; Mackiw, Lin & Kunda, 1957). More specifically, iron sulfate and a small amount of dispersant polymer are added, which leads to the formation of nuclei. Upon subsequent hydrogen reduction, the reduced nickel can then be deposited on the nuclei, allowing nickel particle growth to take place. However, this results in iron contamination in the final nickel product, and so the use of pure nickel as a seed crystal is desirable to produce high quality nickel powders."

Jan 1, 2017

By W. R. Cowan

The Caledon Outwash is a large deposit of glaciofluvial gravel and sand located 55 km north west of down town Toronto. For the most part it was laid down in a meltwater channel fronting the Paris Moraine about 14,000 years ago. Applying the braided bar model of Eynon and Walker (1974) the principal facies present are imbricated coarse gravels (bar corefacies); crudely stratified, imbricated bar top sandy gravels; and trough crossbedded side channel sands and gravelly sands. Of lesser importance are crossbedded bar front gravels, shallow braided stream facies gravels, and stoss side facies sands and pebbly sands (rare). Lithologically these deposits are characterized by FIGURE I. Geological relationships of Caledon Outwash. Industrial Minerals in Canada up to 95% Ordovician limestone and siltstone in the gravel fraction. A siltstone content of up to 34% requires that beneficiation be carried out for the production of better quality aggregates. Resources remaining are estimated to exceed 500 million tonnes in 3000 ha of the most suitable part of the deposit. Present production is about 4.5 million tonnes per year, shared by several producers.

Jan 1, 1984

By E. LaRocque

Denison Mines Limited has been processing uranium ore in the Elliot Lake area for the past 20 years. As the original concept of uranium processing has been modified somewhat over this period, the present paper will describe the current recovery practices at the Denison mill from crushing through concentrate drying and packing . Special emphasis will be placed on the leaching, ion exchange and precipitation sections of the process, with consideration being given to the entire circuit. Descriptions will be given of equipment now being used and control strategies will be discussed.

Jan 1, 1979

By H. Golzarpoor

Life Cycle Analysis (LCA) is a methodology for evaluating the environmental impacts associated with a product during its life cycle. LCA is identified as the most reliable method for verifying environmental impacts; however, current LCA-based approaches have certain limitations for environmental analysis of construction products. Integration of the LCA methodology with Discrete Event Simulation (DES) provides a sound framework for modeling and analysing the environmental impacts of construction products. LCA and DES is one possible combination for analysing the cause and effect of various scenarios where time, resources, and randomness of input variables affect the outcome and, therefore, has the potential to address the shortcomings of LCA in construction. Recent studies in disciplines other than construction such as manufacturing systems have revealed positive effects on evaluation of environmental metrics while integrating LCA with DES; however, this integration has not yet been applied for environmental analysis of construction products. By implementing LCA data in a DES model, this research proposes an environmental model of earthmoving operations in a case study. Environmental variables are simultaneously assessed with production variables in the same simulation model and the integration of DES and LCA is discussed.

Aug 1, 2013

By D. Warkentin, M. Tomlinson, A. Nacu

"This study focuses on comparing the “blinding” effectiveness of various organic additives as part of a general review of using organic additives to mitigate preg-robbing and improve gold recovery during cyanide leaching of carbonaceous gold ores. The laboratory results compared the impact of various organic additives on the prevention of gold adsorption using both activated carbon and ore samples containing carbonaceous matter, generally referred to as total carbonaceous matter (TCM). Although several organic reagents were found to have a significant blinding effect on activated carbon, these results did not translate into increased gold leaching efficiencies with actual ore samples.INTRODUCTION Preg-robbing refractory ores, which contain carbonaceous matter, have been problematic in cyanide plants for many years. In industry, preg-robbing by carbonaceous matter may be mitigated using various techniques including flotation, to remove the carbonaceous matter; addition of blinding agents, to passivate the carbon surface and prevent adsorption of gold cyanide complex; roasting, to partially or fully destroy the carbonaceous matter; chemical oxidation (Abotsi & Osseo-Asare, 1986); pressure oxidation; or microbial deactivation (Dunne, Staunton, & Afewu, 2013). Carbon-in-Leach and Resin-in-Leach are also used as process alternatives in which activated carbon or resins are used to compete with the carbonaceous matter for gold cyanide adsorption. For highly preg-robbing ores, thiosulphate leaching may be considered as an alternative to conventional cyanide chemistry (Baláz, et al., 2000). Natural carbonaceous matter is composed mainly of hydrocarbons, organic acids, and graphitic/amorphous elemental carbon. Natural elemental carbon that behaves as activated carbon is considered to be the main reason for preg-robbing, although the true nature of this phenomenon is not fully understood (Tan, Feng, Lukey, & van Deventer, 2005). In general, the hydrocarbon fraction is assumed to have negligible interaction with gold, while organic acids, such as humic acids, may form complexes with gold, inhibiting its extraction in cyanide solutions. Organic reagents have been used for blinding carbonaceous matter in gold ores prior to cyanidation. Abotsi & Osseo-Asare (1987) tested various organic compounds, with two possible effects in mind. First, the organic compounds were expected to adsorb onto the surface of the solids as anionic/cationic species; the resulting electrostatic interactions between the Au(CN)2- anion and the negatively/positively charged surface would inhibit/enhance the gold uptake. Second, the organic reagents were expected to coat the surface of the carbonaceous component in the ore, limiting the available adsorption sites and reducing gold losses. Optimal reagent dosage depends on ore mineralogy, maturity and surface area of the carbo"

Jan 1, 2019

By K. Yamaguchi

"This Precious metals content copper scrap is treated with a copper smelting process, and copper and precious metals are recovered. The precious metal lost in the slag is increasing with increasing of treated scrap amount. To determine the portion of the precious metals chemically dissolved and that associated with the mechanically trapped matte in the slag will be a key factor to improve the recovery of those metals. The distribution of precious metals between copper matte and FeOx-SiO2 slag phases is determined at 1300 oC and a fixed partial pressure of SO2 of 0.1 atm for the matte grades from 40 to 70 mass% of copper. The distribution ratios show a tendency to increase when the grade of matte is increased. In addition, the solubility of platinum and rhodium in the FeOx-SiO2-CaO slag under the air and the solubility of platinum in the FeOx-SiO2 slag equilibrated with a pure platinum or the Pt-Cu alloy were determined at 1300oC in the range of oxygen partial pressure from 10-9 to 10-7. The solubility of platinum and rhodium in the slag tends to increase with increasing content of copper in the slag and oxygen partial pressure. INTRODUCTIONThe stringency of the supply and demand of platinum group metal (PGM) intensifies by localization for resources and increase of the demand for PGM, furthermore, smelting and the recycling of PGM in the nonferrous smelting process take an important role. Scrap containing a precious metal for the purpose of collection of PGM is treated by copper smelting recently. With increase of the treatment amounts, loss of PGM in slags is the serious problem.In considering of PGM loss in slags, knowledge of the distribution behavior for PGM between slag and matte phases or slag and copper phases and the PGM solubility in slags is important. The distribution ratios of silver and gold between the FeOx-SiO2 slag and the Fe-Cu matte were reported by Roghani (Roghani et al., 2000), Shinha (Shinha et al., 1985) and Takeda (Takeda et al., 1993) However, there is no data for the PGM distribution ratio. On the other hand, the PGM solubility in slags was reported by Sano (Sano et al., 1997) However, there is very few report for the solubility of PGM in the nonferrous smelting slags."

Jan 1, 2012

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

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

Jan 1, 2012

By Wm. D. Jr. Sheldon

THIS paper presents practical data to aid in the selection of a mine ventilating fan, and outlines considerations to be made to ensure satisfactory operation and optimum efficiency of the fan. Much has been published on the general subject of mine ventilation, but the writer has found few specific data on factors and considerations which govern the selection of a fan-data which should be available to the operator if his fan is to give satisfactory service within the range of the high-efficiency portion of its performance curve and deliver the specified air capacity when operating against the anticipated range of resistance of the mine. The total quantity of fresh air furnished by the ventilation system must be adequate to provide, and to ensure at all times, safe and healthful working conditions in the mine. It is assumed that the mine operator is able to determine the air quantities required for the various parts of the mine.

Jan 1, 1941

By J. Ji

Oxygen requirement during acidic pressure oxidation of sulphide concentrates can he as high as 800 kg/(m3.h) in the first compartment of an autoclave when the total oxidation residence time is shortened to 30 minutes. To achieve such a high oxygen transfer rate while maintaining satisfactory oxygen utilization, a pilot plant campaign was launched at the Placer Dome Research Centre in 1998 to measure the oxygen mass transfer coefficient in a 5 m3 pressure vessel at a total pressure close to 500 psig with a 30 HP gassing agitator. Maximum oxygen flowrate to the pressure vessel was designed at 3 600 kg/h. In this paper, a great deal of information is provided on the gassing agitator, pressure vessel. measurement procedure and results. Without oxygen sparging, the oxygen mass transfer coefficient is 2 600 h-1 with a mixing power input of 5.0 kW/m3 in solution. With oxygen sparging, the oxygen mass transfer coefficient increases to 3 500 h-1 with a mixing power input of 4.0 kW/m3 in solution.

Jan 1, 2004

The primary steel industry of the Prairie Provinces consists essentially of three companies whose total annual melting capacity, as of January 1, 1963, was 333,000 net tons. These companies have steel rolling facilities. In addition, there are five other companies producing steel castings by melting scrap in electric furnaces which, with the plants having rolling facilities, gives a total annual furnace melting capacity of 380,000 tons. Since 1954, the annual production of steel ingots and castings has ranged from a low of 66,588 tons, in 1954, to a high of 211,138 tons in 1961. The present facilities on the Prairies for making steel ingots from scrap, and rolling them to produce light merchant mill products, appear to be sufficient for many years ahead. Steel scrap, as electric furnace feed to make steel ingots for the present line of products, will probably continue to be available at reasonable prices either on the Prairies or from the north-central United States. It is recognized, however, that residual elements in scrap may be-come of greater significance and necessitate more rigid control of scrap quality.

Jan 1, 1963

Flotation works very well for particles that are typically in the range 20 to 150 µmin diameter, for base metal ores. In this range, it is possible to get quite high recoveries in conventional flotation machines. Outside this range, the recoveries decline progressively, whether it be with the very fine or the coarse end of the size spectrum. In this paper, the reasons for the drop-off in recovery are briefly reviewed, together with strategies for future work. Recent progress in the author's laboratory will be described. These include the Concorde Cell, which takes advantage of the high dissipation rates in the shock wave created when a gas-liquid mixture reaches supersonic velocities. The environment thus created is very beneficial for the flotation of ultrafines. For coarse particles, the reasons for the drop-off in recovery can be related to the highly turbulent nature of the pulp in a conventional flotation cell. To improve recovery, it is necessary to find a way of bringing particles and bubbles into contact in a quiescent environment. A new process for coarse particle flotation will be described, in which a fluidized bed is created in the flotation cell. The flow conditions are very gentle, and the high solids concentration leads to rapid rates of capture of the particles. Experimental results are presented. It is apparent that the maximum floatable size for coal and minerals can be increased by a factor often over the limits found in current practice. Advances

Jan 1, 2009

By J. M. Harrison

FIELD PARTIES of the Geological Survey of Canada have, in recent years, mapped several areas east and northeast of Flin Flon (9, 10, 12, 13, 14, 15, 118) (1). In all these areas the mineral deposits show a close geographical relation to faults although the immediate control of most of the known ore deposits appears to be folded structures in the rocks (8, 11, 18, 19). Air photographs and published geological maps were studied in order to ascertain whether or not the mineral deposits in the Flin Flon-Snow Lake mineral belt as a whole show any spatial relationship to faults. It was found that most of them do show a distinct geographical association with faults, or with topographic features that may be faults. It is further apparent on air photographs that a major lineament, for composite of lineaments, follows approximately the contact between gneisses and schists of the Kisseynew complex on the north, and Amisk volcanic rocks on the south. However, this Kisseynew lineament is not continuous and, especially in the eastern half of the district, it is exceedingly irregular in plan. Parts of this lineament have been mapped as faults (9, 10, 12, 14), other parts have been shown as stratigraphic contacts (1, 2, 21, 27), and still other parts have not been studied except by reconnaissance methods. It is suggested that the Kisseynew lineament may be due to a deformed fault that extends easterly from a point about 25 miles west-northwest of Flin Flon to a point at least 10 miles northeast of Wekusko lake and that the mineral deposits.

Jan 1, 1951

By A. Warczok

The computer simulator of slag cleaning in an electric furnace is based on a set of mathematical models describing heat liberation, the rate of slag reduction and gravitational coalescence of matte and metallic copper inclusions. The simulator consists of a mass and heat balances as a frame, in which the kinetic equations of the slag reduction and phase separation are incorporated. Friendly user-screen interface allows for dynamic control of process parameters in real time together with simulation of typical operations, such as furnace charging with liquid slag, solid reverts and coke, tapping out of the slag and matte, motion of electrodes and control of the transformer. The batch, semi-continuous or continuous process can be simulated. The process parameters are displayed and continuously recorded for further analysis. The computer simulator is a powerful tool for process optimization, training of furnace operators and the design of new slag cleaning installations.

Jan 1, 2007