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By Robert Marchand

_ North-western Quebec _ Bousquet mining district _ 21 M oz mining camp _ 100,000 t/yr (400 t/d) _ Shrinkage : 52 % _ Longhole : 40% _ Development : 8% _ 5 days/week _ 120 employees _Discovered in 1936. _12 changes of ownership. _Exploration & development : 1987 _Commercial production : 1990 _Shaft deepening : 1995 _Internal shaft : 1999

Oct 1, 2001

By Rob Kelso

What is a Lunabot? ?Robot Controlled Remotely or Autonomously ?Visual and Auditory Isolation from Operator ?Excavates Black Point 1 (BP-1) Simulant ?Weight Limit - 80 kg ?Dimension Limits - 1.5m width x .75m length x .75m height ?Designed, Built and Tested by University Student Teams

May 1, 2013

By L. Nichols, H. Keele, K. Czajewski

"The oil sands industry in northern Alberta disposes of the by-products of bitumen extraction into tailings ponds. The mix of tailings sand and mature fines, with gypsum added as a flocculent, constitute a waste stream called consolidated tailings or composite tails (CT). The two main challenges of the tailings disposal process are a chronic shortage of economical containment and the potential mobility of CT deposits in the event of breaching by, for example, long-term erosion, which makes abandonment of tailings ponds more complicated. Thick deposits of CT do not consolidate in adequate time to provide reassurance for abandonment of the CT containment facilities and that is why the containment must be adequately robust or the properties of CT improved for long-term stability. This paper addresses potential solutions to these problems by applying the gravity drainage of CT and quasi-homogeneous construction for containment facilities.Gravity drainage is suggested as a mechanism to facilitate the consolidation of CT in critical areas of the ponds by inserting internal sand layers within the body of the deposit. Critical CT areas are defined as areas around the perimeter of containment. The advantages of this solution are longterm stability of the disposal area by creation of stiff deposits around its perimeter and facilitation of reclamation efforts on its surface.The present solution for containment of CT deposits is not favourable from a materials balance point of view. In general, sand has become a valuable resource as it both forms the primary component of the CT matrix and becomes the containment building material. If sand is required for CT, less is available for dyke construction. This puts additional pressure on the industry as a whole to maximize the use of overburden and interburden for construction of tailings containment structures. Typically, only 50 percent of dry mine waste (overburden + interburden) is incorporated into dyke construction. The remainder of dry waste is stored either in waste dumps or in tailings ponds, consuming valuable in-pond space originally allocated to CT material. Quasi-homogeneous dykes are proposed to contain CT deposits rather than traditional minimum cross-section dykes built of overburden and/or cell sand. The quasi-homogeneous dyke may change the unfavourable material balance and lessen the construction constraints that are presently imposed on the mine/waste disposal process."

Jan 1, 2008

By Micheal Mayhew

"This presentation will provide information and facts on the drilling and blasting within a frozen shaft using a full face methodology. The Author will briefly discuss the history of the sinking of a frozen shaft in Saskatchewan, the challenges faced, and the solutions delivered. The case study will include a discussion describing the series of events that led to the decision to adopt a full face blasting methodology. The case study will describe key elements such as the drilling pattern, hole size and depth, explosive products and blast vibration monitoring systems. INTRODUCTION During the shaft sinking process at a project in Saskatchewan, it was recommended that a drill & blast benching method be utilized. Following an evaluation of the site, the geological conditions proved amenable to a ground freezing system that would stabilize the rock formations and prevent inflows of water into the shaft (Table 1).The shaft sinking process required that the ground surrounding the proposed shaft be frozen by drilling several near vertical freeze holes radially at a specified distance from the edge of the proposed shaft excavation. This would allow for the circulation of a chilled brine solution down the freeze holes. The circulation is achieved by means of stainless steel pipes encased within each other to provide a delivery and return conduit for the brine (Figure 1). As a result of this periphery freezing, the ground within the shaft excavation is also predominately frozen.The finished internal diameter of the shaft was 5.5 meters with varying overall excavation diameters to facilitate different liner designs used for various rock formations. For the most part, the fully unsupported/unfinished excavated diameter of the shaft varied from 6.9 meters to 8.5 meters depending on the zone excavated. The challenge was that the proposed drill and blast excavation method posed a significant risk due to the potential to damage the freeze wall and pipes during blasting. The decision was made to commence the excavation through mechanical excavation (i.e. chip and cut) methods then change to a drill and blast method utilizing a split benching design. Through this process, the team encountered five main challenges: 1. Sub-horizontal geological structures and slips that affected blasting performance and sometimes resulted in a sloping surface to the top of the bench; 2. Confined working area for the crews and equipment due to a limited work area as a result of utilizing the split (half) bench drill and blast methodology; 3. Mechanical chipping and cutting of the one-meter buffer zone to the final shaft diameter; 4. Crews working on extremely slippery uneven frozen rock and ice surfaces, which affected safety and efficiency; 5. Establishing a blast vibration monitoring system to determine if the drilling and blasting of the full shaft excavation diameter were acceptable."

Jan 1, 2019

By Antonio Laguna

During the past few years, synthetic efforts have been addressed to the synthesis of polymeric materials with optical or semiconducting properties or even with capabilities to act as sensors for volatile organic compounds (1). The most recent studies have been concerned with the synthesis of heterometallic systems with short metal-metal interactions and the determination of their optical properties. These complexes are undoubtly interesting from a theoretical viewpoint and include an emerging class of luminescent materials, in which the observed luminescence results mainly from the interactions among the different metal centers, as well as the environments of these atoms. The understanding of the photophysics and photochemistry of these products could permit the design of complexes with appropriate metals and ligands chosen for individual applications. Following a synthetic strategy consisting on reactions between acid heterometal complexes and basic gold precursors we have succeded in synthesizing supramolecular assemblies formed via acid-base stacking, which display tunable luminescence or selective and reversible sensitivity to Vapour Organic Compounds (VOC?s) (2-5). In this sense, in accord with previous Mössbauer studies carried out in the gold precursor complex [AuR2]- (R = C6F5), the basic properties of this and related materials have been employed as building blocks of supramolecular assemblies in their reactions with complexes in wich the metallic centers act as lewis acids as, for instance, Tl(I) or Ag(I) (6). For instance, in the case of silver, by reacting equimolecular quantities of [NBu4][Au(C6F5)2] with AgClO4, in acetone, a solid of stoichiometry [Au2Ag2(C6F5)4(OCMe2)2]n is obtained. The polymeric nature of this material in solid state is confirmed through X-ray analyses (figure 1). It consists of polymeric chains formed by repetition of Au2Ag2 moieties linked through short gold-gold interactions. Nevertheless, in solution, the nature of this polymeric materials was still a matter of controversy and, in this case, the optical properties have been shown as useful tools in order to obtain information. Thus, while dilute acetone solutions of this complex give a emission coming from pentafluorphenyl localized ??* excited states, in accord with the presence of isolated tetranuclear Au2Ag2 units, when the measurements are done at different and higher concentrations, a band that does not obey the Lambert-Beer law appears. This deviation is consistent with molecular aggregation in solution of these units through gold-gold interactions, and a clear correlation between the emssion wavelength and the structure in the solid state and in solution is shown (figure 2). DFT calculations accord with the observed experimental behaviour and show the nature of the orbitals involved in each transition.

Oct 1, 2003

By Fernando Canales

In the last decade a great dealt of attention has been paid to the chemistry of polynuclear gold derivatives with a central heteroatom, not only from the experimental and structural but also from the theoretical standpoint (1). Interesting hypercoordinated species of the type [C(AuPR3)5]+, [C(AuPR3)6]2+, [N(AuPR3)5]2+, [P(AuPR3)5]2+ and [P(AuPR3)6]3+ have been described (2), apart from other complexes with more conventional stoichiometry, and all have in common the presence of gold-gold interactions of ca. 3 Å. Usually, the chemistry of the first row elements of the p-block is known to follow classical rules of bonding, and only in cases of extreme electron-deficiency has the traditional electron count to be reconsidered to account for special types of molecular or solid state structures. Many of these heteroatomcentered complexes are electron-deficient and the gold-gold interactions provide a significant contribution to their stability. The attraction between these closed-shell systems has recently captured the attention in inorganic and organometallic chemistry (1). P. Pyykkö has summarized the available theoretical and experimental evidence of intra- and inter-molecular interactions between heavy metal atoms, such as Au(I) (3). In general, when heavy metals are involved in such interaction, the closed-shell attractions are of metallophilic or van der Waals type, as has been found for d8, d10 coinage-metals, and d10s2 centers. Experimental evidence for the aurophilic Au(I)???Au(I) attraction is available from crystallography, NMR and Raman spectroscopies, and optical spectroscopic measurements of the interaction strength (4). It is manifested in molecular conformations with relatively close Au???Au interactions of ca. 3 Å, whose strength has been estimated at ca. 25-50 kJ/mol (5). These attractions are weaker than most covalent or ionic bonds, but stronger than other van der Waals bonds and comparable in strength to typical hydrogen bonds. The phenomenon cannot be satisfactorily explained by classical theories such as hybridization (3). From a theoretical point of view, the attractions close to equilibrium distances are caused by dispersive interactions together with relativistic effects (6). The attractive contributions are dominated by pair-excitations from the heavy metal, such as gold, strengthened by relativistic effects. In all the cases where there are inter- or intramolecular metal-metal attractions, the correlation effects are essential; if they are omitted, at the Hartree-Fock (HF) level, repulsive interactions are obtained (6). Hence, it is necessary to use at least second-order Møller-Plesset level (MP2) methods for the dispersive forces, which are included in the electronic correlation effects (7).

Oct 1, 2003

By V. Thangadurai

The critical problems in commercialization of the present solid oxide fuel cells (SOFCs) and useful electrical and physical properties of IT-SOFCs electrolytes are discussed briefly. The conductivity behavior of the perovskite-type and perovskite-related structure metal oxides are summarized. Perovskite-type La0.9Sr0.10ao.8Mgo.203_8 (LSGM) exhibits high ionic conductivity of 0.1 S/cm at 800 °C. In order to improve the electrical conductivity of LSGM and the In analogue of LSGM, substitutions have been made at the A-and B-sites. Substitution of trivalent La3+ by isovalent Pr3+ in LSGM and La09Sro ilno.sMgo.202.85 (LSIM), yield new mixed conductors. Among the investigated compounds in the present study, La0APro.4Sro.21no sMgo.2028 exhibits the highest total electrical conductivity of 3.2 x 10-5 S/cm at 200 °C in air with the lowest activation energy of 0.44 eV (200-700 °C). LSIM and Pr-substituted LSGM and LSIM perovskites show p-type electronic conductivity in the oxygen partial pressure range 10-5 ? 0.21 atm at 750 °C. The perovskite-type solid solution SrSn1-xFex03 exhibits compositionally driven phase transitions. For example, compound of the composition x = 0 to 0.3 crystallizes in an orthorhombic GdFeO3-like perovskite structure, where as that of x = 0.4-0.9 crystallizes in simple cubic SrTiO3-like, and that of x = 1 in orthorhombic structure. The electrical conductivity of SrSn1_xFex03 increases with the increasing Fe content and reaches a value of about 10-1 S/cm at 25 °C in air for x = 1. Compounds with low Fe content exhibit both ionic and electronic conductivity, while the higher Fe content 0.5) perovskites are electronic conductors over a wide range of oxygen partial pressure of 10-22? 0.21 atm at 750 °C.

Jan 1, 2005

By R. V. Kumar, C. Schwandt, S. Subramanian, S. Sampath, S. Vimalnath

Waste waters and industrial effluents containing toxic heavy metals such as Pb, Hg, As and Cr, beyond permissible limits (0.1 ppm, 0.01 ppm, 0.02 ppm and 0.05 ppm, respectively) (WHO, 2011), pose serious threats to the ecosystem, when discharged into the environment. It thus becomes imperative to monitor the concentration of the toxic heavy metals and treat them before discharging. In this study with the above objective in mind, a microbial sensor has been developed using P. aeruginosa cells drop coated onto a carbon paste electrode (CPE) for the detection of Pb(II) ions in aqueous solutions. Detailed biosorption experiments confirmed that complete bioremoval of lead could be achieved using P. aeruginosa cells under optimum conditions of pH and biomass loading. For the electro-analytical determination, the Pb(II) ions were initially preconcentrated on the P. aeruginosa modified CPE, reduced and then anodically detected with oxidation of Pb0 using differential pulse anodic stripping voltammetry (DPASV) technique. All the experiments were carried out in the presence of 0.1 M sodium acetate buffer and KNO3 as background electrolyte. The parameters namely pH, scan rate, biomass loading, deposition potential, deposition time and concentration of Pb(II) ions were optimized for the biomass modified CPE, to obtain maximum voltammetric response. Under the optimized conditions of pH 5.6, scan rate of 0.008 V/s, biomass content of 1.7–1.8×109 cells/mL, deposition potential of –1.2 V, and deposition time of 500 s, the developed microbial sensor showed a linear response for the peak current between 0.5×10-5 M and 1×10-8 M of Pb(II) ions with the R2 value of 0.9979. The lowest limit of detection (LLOD) was determined as 2.4×10-7M of Pb(II) ions (0.05 ppm). The redox mechanisms of Pb(II) ions on the P. aeruginosa modified CPE surface are discussed.

Jan 1, 2016

By T. A. Chepushtanova, I. U. Motovilov, V. A. Luganov, G. D. Guseynova

"Kazakhstan has significant reserves of oxidized nickel ore with low nickel content. Processing of such ores is effective via pyrometallurgical technologies such as smelting to ferronickel or matte with consecutive extraction of metals from melting products. There are also many options for direct hydrometallurgical processing. However, they do not provide high nickel recovery.We studied extraction of nickel and cobalt by using combined technology of low-grade ore processing from the Shevchenko Kazakhstan deposit, which includes smelting to poor matte followed by hydrometallurgical processing of the matte. The nickel content in the initial ore was 0.75–0.9%. Pyrite concentrate of Sokolov-Sarybai Kazakhstan deposit containing 0.17% cobalt and 0.06% nickel had been used as a sulphidizer at smelting.As a result of the study, we were able to extract up to 98% of the nickel and up to 80% of the available cobalt. Obtained matte contained 3.1% nickel, 0.29% cobalt, 71.6% iron and 24.3% sulfur. Matte leaching was carried out by hydrochloric acid. The iron content in the pregnant solution was 170–185 g/dm3. Nickel and cobalt precipitation from the pregnant solution was carried out by hydrogen sulfide. Initial matte or cast iron was used as neutralizers during the process. Practically complete precipitation of nickel was achieved at 3–5% excess of neutralizer at 333–343°K in 4 hours; up to 95% of cobalt precipitation was achieved in 50% excess of neutralizer at 363°K in 8 hours. The purified solution contained 200–600 mg/dm3 of nickel, 80-100 mg/dm3of cobalt and 175–185 g/dm3 of iron. The purified solution allows to obtain iron oxides by high-temperature hydrolysis, and to regenerate hydrochloric acid.As a result of non-ferrous metals precipitation we obtained sulfide concentrate with concentration of Ni up to 25% and concentration of Co up to 5.5%. The technology is recommended for semi-industrial testing."

Jan 1, 2016

By G. E. Agar, W. B. Kipkie, R. A. Tenbergen

"This paper describes the successful pilot plant testing of a flowsheet, developed at Inca Ltd.&apos;s J. Roy Gordon Research Laboratory (JRGRL), for improved separation of the constituent minerals in a bulk copper-nickel concentrate.The flowsheet utilized the fact that xanthate concentration on the mineral surfaces is the most important variable affecting the high lime separation and that reduction of the xanthate concentration could be affected by an oxidative conditioning stage.When the adsorbed xanthate was reduced to a suitably low level, chalcopyrite and pyrrhotite could be floated to leave a high-grade nickel concentrate as the residue. Sodium cyanide was added to the froth product 10 depress pyrrhotite and any remaining pentlandite. The temporary depression of chalcopyrite by the sodium cyanide was overcome by introducing a conditioning step following cyanide addition . The advantages of the flowsheet; improved copper-nickel separation, operation without heating of the pulp, production of both high-grade nickel concentrate (29.5% Ni, .26% Cu) and copper concentrate (<0.5% Ni, 30% Cu) were all realized.IntroductionLaboratory test work undertaken at Inco and described previously(1) has demonstrated the importance of low xanthate concentrations on the mineral surfaces in the copper-nickel separation at the Copper Cliff mill.It was determined that oxidation of the separator feed (a cleaned bulk concentrate, grading about 11% Cu and 11% Ni) to lower the xanthate concentration on the mineral surfaces prior to separation significantly improved the quality of the separation (CI 4)*, particularly when no cyanide is used, but also when cyanide is added. As well, it was found that the oxidation permitted a pentlandite-pyrrhotite separation - the oxidation step brought about a decrease in the flotation-rate of pentlandite leaving pyrrhotite and chalcopyrite largely unaffected. Consequently, it was possible to produce a highgrade pentlandite product (30% Ni, 0.3% Cu) at ~83% Ni recovery , as well as a high-grade copper concentrate (0.5% Ni, 30% Cu at -96% Cu recovery)."

Jan 1, 1985

By K. K. Mamyrbayeva, G. Z. Guseinova, V. A. Luganov

This paper presents the results of theoretical and technological studies on leaching ore of Bozshakol deposit. The major oxidized copper mineral of Bozshakol ore is atacamite, and the content of malachite, azurite, cuprite, copper sulfide minerals is insignificant. The chemical composition of the test sample, with mass, %: 0.9 Cu; 55.00 SiO2; 22.40 AI2O3; 9.00 Fe2O3; 2.24 CaO; 3.56 MgO. Technological studies on leaching the ore were carried out with sulfuric acid 5, 15, 25 and 50 g/dm3 in one and multistage (up to 4 stages) countercurrents. Size of ore 1 mm, the ratio S : L 1: 4. It was found that with increasing of acid concentration from 5 to 50 at one-mode leaching the copper extraction reaches 55 %. At multistage leaching the copper recovery reaches 73 %, a productive solution containing g/dm3: Cu - 4; Fe – 1.2; Al – 1.1 at a pH of 2.8. When leaching the cinders obtained by ore roasting in an atmosphere with limited value of oxygen, with increasing of the roasting temperature from 650 up to 750 °? the extraction of copper from the Bozshakol ores reduced from 31.5% down to 24.4%. The rate of filtration is markedly increased compared with the crude ore. When leaching the cinders obtained by roasting in a nitrogen atmosphere a marked increase up to 87.2-88 % of copper extraction from Bozshakol ores is observed. As an extractant for the extraction of copper from the pregnant solution used LIX984N extractant in kerosene. On the basis of the calculations it was built extraction isotherm and modeled processes of copper recovery with the circuit 2E + 1P, allowing up to 93 and 96% extraction. The through extraction of copper from the ore in the electrolyte was 67 %. Technology is recommended for semi test.

Jan 1, 2016

By B. K. Gorain

"The Air-Metabisulfite (AMBS) process has been developed to treat complex copper, molybdenum and precious metals (gold, silver) bearing ores using brackish and sea water. This process was able to produce significantly better copper concentrate grades and higher recovery than possible using conventional lime and/or cyanide based processes for certain complex ores with high sulphide gangue minerals. The AMBS technology in now used commercially following a rigorous evaluation process emanating from laboratory curiosity to plant applications. In addition, many projects have embraced this technology at a feasibility level involving various complex ore bodies.This paper presents a summary of the findings from testing the AMBS technology on various ores types. Experiences gained from this successful development from bench to piloting and then to plant applications are also presented in this paper.INTRODUCTIONProcessing of sulphide copper, molybdenum and precious metal ores typically involves flotation of copper bearing minerals while depressing non-sulphide and other non-copper sulphide minerals such as pyrite, pyrrhotite and sometimes arsenopyrite. Though in many precious metal ores, the values are in these sulphide minerals as well. The conventional method employed in the industry to produce a saleable copper concentrate requires depression of these non-copper sulphide minerals using an alkaline pH above 9.5, though 10.5 to 11.5 is more common. Alkalinity is controlled by the addition of a pH modifier. Lime is normally employed as a pH modifier due to its relatively low cost; however, the costs associated with adding lime can be significant and the effectiveness of lime as a depressant could be drastically reduced when the process water supplied to the flotation circuit contains high level of dissolved salts. In some cases, a small amount of cyanide is used along with lime to depress non-copper sulphide minerals especially when its proportion to copper sulphide minerals in the ore is higher than three."

Jan 1, 2016

By F. Rashchi, A. Ghahreman, J. Moghaddam, S. Karimi

"In this work, the kinetics of ferric-ferrous redox reaction on a passivated sphalerite electrode is studied by the application of chronoamperometry method in 0.5 mol L-1 H2SO4 solution. The electrochemical behaviour of ferric and ferrous couple was investigated on two sphalerite minerals acquired from two different geological regions (Iran and Canada). The carbon paste electrode (CPE) of sphalerite was initially prepared and passivated in 0.5 mol L-1 H2SO4 solution and after reaching the steady-state condition, ferric and ferrous sulfate solution was added. The registered current density differences between the current density of the passivated electrode and that of the electrode in the iron salts added solution were interpreted for kinetics of ferric-ferrous reduction-oxidation reaction. The equilibrium potential for ferric-ferrous couple on sphalerites surface was about 500 mV vs. Ag/AgCl. The exchange current density (i0) for ferric-ferrous redox reactions on the sphalerite samples had the same magnitude of 10–6 A cm–2. The electrochemical impedance spectroscopy (EIS) measurements at 550 mV vs. Ag/AgCl illustrated that sphalerite oxidation was controlled by two charge transfer process in both of Fe-free and Fe-containing solutions. The charge transfer resistance remarkably decreased when Fe salts added to the 0.5 mol L-1 H2SO4 solution, owing to the ferric-ferrous redox reactions being the predominant reactions on sphalerite surface.INTRODUCTION Direct atmospheric leaching processes for sphalerite concentrates are performed by reduction of an oxidant (e.g. ferric ion and oxygen) and oxidation of sphalerite to zinc ions. Oxygen could be an oxidizing agent as well, but the concentration of oxygen is very low in the solution. The role of ferric ions is predominate in this oxidation (Babu, Sahu, & Pandey, 2002). The oxidation reaction for sphalerite in the presence of ferric in can be presented as follows (Lampinen, Laari, & Turunen, 2015):"

Jan 1, 2017

By Ming’an Wei, Qifu Duan, Chengtie Wang, Jianmin Wu, Erdenet Mining Corporation, Lusheng Zhou, Wenjian Wu

"The development and industrial applications of magnetic liners have been around for decades. A Swedish company SKEGA first introduced the rubber magnetic liner, and then followed by the Chinese metal magnetic liners. Traditionally, both liners were used in secondary and regrind ball mills, especially for ores containing magnetic materials. Mostly steel balls or equivalent forged balls, usually below 60 mm in diameter, were used as grinding media. With the development of mineral resources over the years, SAG mills and larger ball mills have become the dominating components for comminution circuit design globally. Due to the larger ball mills and thus larger grinding balls (up to 80 mm) especially in non-ferrous metal industry, the application of magnetic liners has been very limited. Based on the new development of materials, the HMR Group Limited has been exploring new designs of magnetic liners which led to the invention of the Nano Steel Magnetic Liner (NSML). Recently, NSML had a successful application in a 5.5m x 6.5m ball mill with 80 mm balls at the Erdenet Copper Mine in Mongolia. Although the first set of liner had a service life of 12 months, the design of the second set was modified and improved to accommodate the operating conditions, which is expected to reach least 24 months service life. The original manganese steel liner only had about 8 months liner life.1. IntroductionMineral processing is an essential operation for the development and utilization of mineral resources. The effective concentration of various useful minerals depends on sufficient liberation of the mineral particles in the ore. The crushing and grinding process such as ball milling are the most common way to achieve proper liberation size. Mill liner is an eternal topic for the operations, not only because it protects the equipment, but also affects the mill grinding efficiency and availability. In recent years, magnetic mill liners have been developed that can easily meet the production targets with numerous benefits, such as zero maintenance, reducing specific energy consumption and improving grinding efficiency."

Jan 1, 2016

By Dave M. Jakelski, Gerhard W. Heinrich, Peter F. Wells, Daniel F. Suarez, Jamie B. Beauchamp

"A need for increased copper concentrate production and for less nickel contamination in that concentrate at Falconbridge's Strathcona Concentrator during 1993 required the installation of a third flotation column and improvements in the operation of the existing circuit. Innovations in wash water distributor and launder design were fundamental in achieving objectives on a new 3 meter diameter column. Minnovex plug-type spargers were tested on a 2.1 meter diameter column in order to reduce maintenance requirements. The paper details the measures taken to enhance column flotation circuit performance and the results obtained.IntroductionFalconbridge's Strathcona Mill receives 10% of the ore feed from a rich copper zone in the Strathcona mine. This ore typically grades 7% copper and 1 % nickel and, because of its high copper content, it is processed in a separate circuit ('C' circuit) which consists of conventional grinding, rougher flotation with the concentrate being treated with lime to pH 12.3 and cyanide followed by column flotation. The columns are operated with mechanical flotation scavengers and the scavenger concentrate is returned to the column feed. Testwork conducted in 1991 showed that the original flotation column (Column 2) could only produce 5.5 mtph of concentrate with less than 0.5% nickel (Figure 1).Between 1991 and spring of 1993, production was increased at times by open circuiting the scavengers. Metallurgical improvements were obtained by improving the wash water distributor of Column 2 and by wash water sprays in conventional cells (1 ). However, the scavenger concentrate assayed 0. 7 -0.8% nickel and was not acceptable for long term operation. When Copper ore production targets were significantly increased for 1993, it became clear that an additional flotation column was required. The new flowsheet employs mechanical scavenger cells on the column tails as in Column 2 circuit (Figure 2). A -parallel configuration was chosen as pulp solids fraction has been shown to be a critical variable and the second column in series would receive a lower density feed."

Jan 1, 1994

By Ming’an Wei, Qifu Duan, Jianmin Wu, Lusheng Zhou, HMR Group Limited / HMR Technology Canada Ltd., Wenjian Wu

"The development and industrial applications of magnetic liners have been around for decades. A Swedish company Skega first introduced rubber magnetic liners, and then followed by the Chinese invention of metal magnetic liners. Traditionally, both liners were used in secondary ball mills and regrind ball mills for industrial applications, especially for ores containing magnetic materials. Mostly steel balls or equivalent forged steel balls, usually below 60 mm in diameter, were used as grinding media. With the development of mineral resources over the years, SAG mills and larger ball mills have become the dominating components for comminution circuit design. Due to the development of larger diameter ball mills and thus larger ball charge diameter (up to 80mm) especially in non-ferrous metal mining operations, the application of magnetic liners have been very limited. Based on “Magnetism theory” and the development of new materials, the HMR Group Limited has been conducting researches on examining the structures of magnetic liners, associated materials and magnetic circuits which led to the invention of the Nano Alloy Steel Magnetic Liner (NSML). Recently, NSML had a successful application in a 5.5 m diameter × 6.5 m ball mill at the Erdenet Copper Mine in Mongolia. The NSML successfully handled a ball size up of to 80 mm, while achieving a service life of nearly 24 months, which is almost 3 times wear life of the original manganese steel liners.INTRODUCTIONDuring the development and utilization of mineral resources, mineral processing is an indispensably operation. The effective separation of various useful minerals depends on sufficient monomer dissociation of minerals in the ores. The crushing and grinding operation is prior operation of mineral processing. Ball mills are used in different concentrators. As the main consumption parts of mill, liners have been studied constantly. As a new kind of liners, magnetic liners besides can meet the requirement of production, and also have characteristics with maintenance-free, saving energy and increasing efficiency. So the magnetic liners are worth to be spread and applied world widely."

Jan 1, 2016

By R. Kaibyshev, I. Zuiko

"The effect of thermomechanical processing (TMP) including cold working by tension or rolling before aging on mechanical properties and a dispersion of secondary phase particles of an AA2519 alloy (Al–5.64Cu–0.33Mn–0.23Mg–0.15Zr–0.11Ti–0.09V–0.08Fe–0.01Si (wt. %)) was examined. It was shown that the use of tension to impart pre-strain =10% in the alloy provides the best combination of properties in peak aging condition of the material. The AA2519T87 alloy processed by tension exhibits the yield stress, ultimate tensile strength and elongation to failure of 425.4±2.4, 475±2.4, 12.1±0.4, respectively. The one pass rolling leads to -5% decreases in yield stress and –20% decreases in ductility. Dislocations serve as heterogeneous nucleation sites for ?'-phase precipitation. The positive effect of tension is attributed to the formation of more homogeneous dislocation structure, uniform distribution of ?'-phase particles with higher aspect ratio. Precipitation of the O–phase plates on {111}a habit plane was detected. In addition it was shown that the low angle boundaries are heterogeneous nucleation sites for ?' and O particles.INTRODUCTION Precipitation-hardened alloys are a commercially important group of the materials because their mechanical properties can be modified by treating and changing the microstructure. Age-hardenable aluminum alloys belonging to the Al–Cu–Mg system are widely used due to their attractive combination of high specific strength, good fracture toughness and enhanced creep resistance attributed to the highly efficient strengthening by the ?'-phase (Palmer, StJohn, Nie, & Qian, 2017). An AA2519 alloy exhibiting superior mechanical properties among high-strength Al–Cu alloys (2XXX-serie) due to minor Mg additives. It is known that, in alloys with high Cu/Mg ratio (=5.6) the classic precipitation sequence (Palmer et al., 2017; Wang & Starink, 2005; Wang et al., 2010) of supersaturated solid solution (SSSS) decomposition occurs: SSSS ? GP zones (Al3Cu) ? ?? (Al3Cu) ? ?' (Al2Cu) ? ? (Al2Cu) (1)"

Jan 1, 2018

By H. A. Kramer

The Flin Flon mine is at present hoisting 5,000 tons of ore daily. In 1943, a peak production of 6,188 tons daily was maintained. Before discussing secondary blasting, it is felt that a short review of stoping procedure will be of value. A list of papers dealing with various phases of the mining operation is appended. Stoping Practice Stopes vary in width from 30 to 150 feet, in length from 200 to 450 feet, and in height from 200 to 500 feet. A typical stope would be 300 feet long, 300 feet high, and 90 feet wide. The dip of the orebodies varies from 60° to vertical, with an average of 82°. The ore is heavy, having a specific gravity of 4. Both sub-level benching and radial drilling are practised. Blast-hole diamond drills are responsible for 85 per cent of the total production. Two sizes of hole are drilled at present, 11 7 6 -inch and 1%-inch. For bench drilling, the burden on the 1 7/16-inch hole is six feet, with holes on 8-foot staggered centres; in radial drilling, the burden is six feet, with nine feet between the toes. The burden on the l 5/8-inch hole is eight feet with holes on 11-foot centres, or 13 feet on the toe of fanned holes. Holes are blasted with two No. 8 instantaneous electric caps per hole, using 403 Polar Forcite. A powder charge of 0.15 pounds per ton is used, with a tonnage factor of 5 and 6.4 per foot drilled for the two sizes. Two rows of holes are blasted together. Fragmentation is satisfactory with the above drilling pattern and powder charge. However, some overbreak and fall-off locally results in oversize material. It is estimated that, after primary breaking, 75 per cent of the ore will pass through the grizzly openings and 15 per cent will pass the draw points, leaving 10 per cent that must be broken further to pass the draw points.

Jan 1, 1946

By The Staff

"IntroductionThe difficulties to be overcome in shaft sinking and development at depth are centred at McIntyre in and around No. 12 shaft. Generally they have been due to increasing pressures and temperatures and to changes in geological structure. This paper describes the difficulties and the measures taken to meet them.No. 12 ShaftEarly in 1932 McIntyre decided to sink an internal shaft from the deepest level, the 3875, to explore and develop known veins extending below that horizon. The location selected was a point 1,500 feet south and 1,000 feet east of No. 11 shaft, the main operating shaft through which the men, ore, and most of the materials are handled Shaft and auxiliary equipment we1e laid out for 3,000 feet, that is to 7,000 feet below the surface. The shaft was to be capable of handling 1,500 tons of ore daily from the lowest level, and was to provide ventilation, compressed air, water, and electric power to service all levels.No. 12 is an internal, vertical shaft extending from the 3875 to the 6825 level and bottoming at 7,007 feet. It is a rectangular, four-compartment shaft, 21 ft. 9¾, in. by 6 ft. 3¾ in. outside the timber (Figure 1).There are two hoisting compartments and one airway compartment, each 4 ft. 10¾ in. by 5 ft. ¾ in. inside dimensions, and a combined manway, pipe, and cable compartment 4 ft. ¼ in. by 5 ft. ¾ in. Levels have been established at 150-foot intervals, pump stations and sumps at 1,000-foot intervals, and. there are two loading pockets, one at the 5525 and the other at the 6900 level.The shaft was sunk in two stages. The first stage, consisting of preparation above the 3875 level and sinking to 5585 feet, was completed by March 31st, 1935. The second stage of deepening the shaft from 5,585 to 7,007 feet was started August lst, 1946, and completed three years later."

Jan 1, 1950

By Kéiko H. Hattori, Neil T. Pettigrew

"Abstract- The Legris Lake Complex is a northeast-trending, 7.3 km long by 3.5 km wide, maficultramafic intrusive complex located in the western Wabigoon Subprovince of the Archean Superior Province. It is one of a series of mafic-ultramafic igneous complexes, the most notable of which is the Lac des Iles Complex, host to Canada’s only producing Pd mine with reserves of 93.5 Mt grading 1.53 g/t Pd as of December 2001 (North American Palladium Ltd., Annual Report 2001). Cu-Ni- PGE mineralization was first discovered in the Legris Lake Complex by a local prospector late in 1999. Shortly thereafter, the property was jointly optioned by Avalon Ventures Ltd. and Starcore Resources Ltd., who continue to discover significant Cu-Ni-PGE mineralization.One of the most notable features of the Legris Lake Complex is the occurrence of extensive brecciation caused by multiple injections of volatile-rich magma. The Cu-Ni-PGE mineralization is hosted by leucogabbro within a 2 km long by 600 m wide, highly brecciated area in the northwestern area of the Complex. The mineralized rocks contain disseminated to blebby sulfides (1 to 5 vol.%), comprising chalcopyrite + pyrite ± pyrrhotite + millerite ± pentlandite, typically surrounded by epidote. The majority of the Cu-Ni-PGE mineralization displays low ratios of Pt/Pd (~0.20) and high ratios of Cu/Ni (~2.9), which are similar to those of the Lac des Iles and the River Valley mafic-ultramafic complexes. The lithologies in Legris Lake Complex bear similarities to the heterolithic gabbro of the Twilight and Roby Zone deposits at the nearby Lac des Iles mine. However, the mineralization at Legris Lake, which is restricted to leucogabbro overlying unmineralized clinopyroxenite, is similar to that of stratiform deposits such as the Stillwater and Munni Munni complexes. The mineralization at Legris Lake is best explained by the late-stage, immiscible separation of a sulfide melt from volatile-rich parental magmas and the subsequent minor redistribution of metals by deuteric fluids."

Jan 1, 2001