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By Robert MacPhail, Louis Racine, Bret G. Cousins

"Sodium Isopropyl Xanthate is employed in a wide variety of sulphide flotation procedures due to the specificity over other xanthates and the cost savings over other sulphide collectors such as thionocarbamates, xanthogen formates and dithiophosphates. An investigation carried out by Tennants Research Technology Applications (TRTA), has produced a chemically enhanced SIPX that has shown all the advantages of normal SIPX along with improved recoveries.Currently, FLEX 31 is manufactured at the Prospec Chemicals xanthate plant in Fort Saskatchewan, Alberta, Canada. Prospec, founded in 1983, produces approximately 4500 tonnes of xanthates per year and is presently undergoing expansion to produce approximately 6000 tonnes. The proprietary process for FLEX 31 is a new addition to plant operations that include the production of xanthates, thionocarbamates and xanthogen formates.This paper will cover the testwork on the zinc flotation circuit at the Louvicourt Mine in Val d'Or, Quebec. This testwork indicated that FLEX 31 could replace SIPX due to the ability of the FLEX to improve zinc flotation kinetics without adversely affecting final concentrate grade. FLEX has presently replaced SIPX in the zinc circuit at Louvicourt.BackgroundThe Louvicourt deposit is hosted by greenschist facies Archean volcanic rocks located in the southeast portion of the Abitibi Greenstone Belt. In the Val d'Or area, the volcanic rocks form an east-west striking, steeply dipping and south-facing sequence composed of thick mafic and ultramafic flows of the Lower Malartic Group.The joint venture has three partners-Aur Louvicourt Inc. (30%), Novicourt Inc. (45%) and Teck Corp (25%). Aur Resources holds 30% interest through Aur Louvicourt, a wholly owned subsidiary and is the operator of the project. Novicourt Inc. is 60.9% held by Noranda.Approximately 75% to 85% of Louvicourt's revenue comes from the sale of copper; 10% from gold (in the copper concentrate); 7% from zinc and 3% from silver (also in the copper concentrate) The copper at Louvicourt is fairly ""fast floating"". About 60% of the copper in the ore can be concentrated in the first flotation cells with feed grades of approximately 4% copper.The initial testwork and pilot trials had indicated that it would be difficult to achieve zinc grades over 52% and recoveries would have to be kept at a level amenable to salable concentrate grades. The actual mill results, with a fine regrind, achieved zinc grades of 55% to 56% at about 75% recovery."

Jan 1, 1996

By E. M. Hillary, D. F. Sangster

"Abstract - As a contribution to the World Minerals Geoscience Database Project, an index-level database comprising information on 142 SEDEX Pb-Zn deposits and occurrences was constructed in Microsoft Access 95 format. Although consisting of only basic information, the compilation, with full referencing of all data, is possibly the largest non-corporate SEDEX Pb-Zn database in the world. It thus provides an opportunity to make geoscience queries of SEDEX Pb-Zn deposits on a scale not previously available. This report provides a summary of the results of a few such queries both as a contribution to a better understanding of SEDEX lead-zinc deposits and as a demonstration of what even such a simple database can offer.During the compilation process, it became apparent that SEDEX Pb-Zn deposits could be assigned to one of the following three sub-types: Type 1 (53.5% of total) in which sulfide minerals are the only exhalative components; Type 2 (33.1%) in which the exhalative components are sulfides plus chert, siderite/ankerite, and/or barite; Type 3 (13.4%) which contains any components of Types 1 and 2 plus abundant iron-oxide minerals (commonly magnetite) and/or is associated with iron oxide-rich sedimentary rocks.Highlights of systematic queries of the database reveal that:1. a majority (84%) of the world’s SEDEX Pb-Zn deposits are not underlain by a discordant footwall feeder zone or vent complex;2. 63.6% of deposits occur in Proterozoic strata followed by a second peak in the Devonian (16.8%). Of those in the Proterozoic, 83% of Type 3 deposits occur in the Gondwanaland supercontinent, whereas only 42% of Type 2 and 39% of Type 1 are located in this large paleo-landmass;3. median (Pb+Zn) grades in Types 1 and 2 fall between 9% and 10% with a majority clustering within this range. Type 3 grades, however, range from ~1% to ~20% (Pb+Zn) with a median of 5%; 4. median size of Type 3 deposits is ~33 Mt, whereas Type 1 and 2 median sizes are ~7 Mt and ~8 Mt, respectively.The report concludes with a brief discussion of how an index-level database such as this can be used in conjunction with other, similar databases to reveal interesting and instructive relationships for guiding both research and exploration."

Jan 1, 1998

Heap leaching accounts for around 20% of the global production of copper. In heap leaching coarsely crushed ore particles are contacted with an aqueous lixiviant percolating through large packed beds (heaps) of the ore particles, with the loaded solution collected at the foot of the heap for metal recovery, solution make-up and recycle. While dominant for copper oxide acid leaching, it finds application also in copper sulphide (bio) leaching. Despite their apparent simplicity, heaps are in fact complex chemical reactors in which a number of competing sub-processes determine the overall rate of metal extraction. Therefore, meaningful prediction of copper production rates from heaps, as well as operational optimisation and control, all hinge on a comprehensive and quantitative understanding of these complexities. The paper gives a comprehensive overview of the various sub-processes in copper heap leaching and their interactions, including those relating to mineral-particle-solution-gas phase interactions at the micro-scale bulk transport and distribution of solution in unsaturated dense beds at the macro-scale. On the basis of a more fundamental understanding and modelling, the potential for optimized operation offers new opportunities for the extraction of value from both low-grade and complex polymetallic ores. INTRODUCTION Heap leaching is one of the percolation leaching technologies, which includes in situ leaching, dump leaching, heap leaching and vat leaching (Bartlett, 1998). The common phenomenon in all these technologies is the transport of leach solution through a fixed bed of ore particles, which interacts with the solid as it migrates along its path. The target minerals are released into solution and with its flow are carried out of the bed. The targeted dissolved metal species are recovered from solution through various hydrometallurgical techniques, especially solvent extraction (SX), before the barren solution is returned to the leach process. Heap leaching is practised in three main types of applications: sulphuric acid leaching of copper oxide ores, cyanide leaching of gold ores as well as the oxidative acid leaching of secondary copper sulphide ores, usually with the assistance of certain microorganisms and, more recently, in the presence of chlorides (CaCl2, sea water). While gold heap leaching dates back to the 1960s, large scale copper heap leaching commenced in Chile from the 1980s, and sulphide heap leaching emerged essentially as a consequence of acid leaching from mixed oxide/sulphide ores from the 1990s (Watling, 2006).

Jan 1, 2019

By Sina Mostaghel, M. Guo, E. Klaffenbach

The current paper is focused on evaluating the behavior of copper in a copper production slag and the assessment of metal recovery possibilities. For this reason, thermodynamic calculations were carried out to determine the influence of parameters such as slag composition, temperature and oxygen partial pressure on recovery of non-ferrous elements. It was found that the formation of a solid face centered cubic (fcc) phase, mainly consisting of iron, limits the extent of reduction that can be achieved for an industrial process. However, it could be discovered that solubility of iron in the Cu-rich liquid alloy strongly depends on presence of other elements, particularly arsenic. As part of the results of this study, the solubility of Cu in the slag at onset of fcc phase formation is presented as function of temperature, oxygen partial pressure and concentration of MgO and CaO. It was concluded that lower temperature (1200°C) and higher CaO addition (16 mass-% in this study), would result in the highest achievable copper recovery from slag. INTRODUCTION Copper slag is produced as a by-product of the copper production process. The slag is still containing a high amount of copper and other valuable metals. This demands further treatment of the slag as a secondary source of metals. In addition, the physical/mechanical properties of slag allow its use in several applications such as hydraulic engineering, civil engineering, as an abrasive and many more. These applications help reducing the use of natural resources like rock and sand. By decreasing the content of heavy metals such as Cu, Pb, Zn and As in the slag, its application in abovementioned fields becomes more attractive as the probability of releasing these elements to the environment is further reduced. To develop a pyrometallurgical slag reduction process, it is necessary to understand what describes the end point of the reduction. This point is expected at maximum copper recovery without facing operational challenges. Existing literature provide some information on the effect of different parameters such as temperature or concentration of different slag formers, on solubility of copper and heavy metals in different phases. However, due to complexity of industrial slags, equilibrium experiments and accurate thermodynamic modelling are essential to be able to reliably predict the behavior of different elements in the system and determine the end-point of reduction.

Jan 1, 2019

By R. Dawes, B. R. Davis, T. LeBel, C. Pelow

Settling of copper, lead, and other metals in slags is an important part of overall recovery of metals, both in primary and secondary operations. The impact of slag reduction on settling is not well known. This is important since the cycle time for slag cleaning can be significantly changed with reduction of slag, through a decrease in slag viscosity and growth of metal/sulphide particles. KPM has worked with a number of primary and secondary recyclers to better understand how reduction of slag can improve overall operation. In this presentation, we describe the type of test work and analysis done and how this information can be used to guide improvements in metal recovery from slags. INTRODUCTION Typically, copper smelter slags contain about 1-2% copper, both in the form of copper and copper sulphide beads and dissolved copper oxide in the slag (converter slags are recycled to the smelter). With the average grade of world copper deposits close to 0.7%, it is practical to recovery copper from these slags. This is done routinely by one of two methods. The first method uses slow cooling in large ladles or beds to grow particles of copper sulphide. The slag is then crushed, milled, and floated for copper recovery. The slow cooling promotes precipitation and growing of the sulphide particles from the solution and then, the slag can then be milled to recover the copper through froth flotation. A consideration of this process includes the energy cost related to grinding, the use of water, capital cost for ladles, and the availability of tailing dumps. At the same time, the retreatment of the Cu concentrate produced in the mill will displace fresh concentrate, which ultimately results in a reduction of the throughput. The second method is the so-called “slag cleaning” and is based on the settling of copper and copper sulphides particles dispersed in the slag. This is achieved by decreasing its viscosity by increasing the slag temperature and reducing the iron spinel solid phases (magnetite) that increase the apparent viscosity of the slag. Typically, the reduction does not extend to the reduction of copper, and soluble copper (mainly as oxide) and fine matte particles remain at concentrations typically at or above 0.5% (Cardona et al., 2011; Cardona, 2011). This is an improvement on 20 years ago when slag was sent to dumps at 1% or less (Demetrio et al., 2000).

Jan 1, 2019

By F. Rosenblum, P. Spira

"Self-heating of broken sulphide ore or waste rock underground can lead to ignition and cause serious problems in a mine operation. An extensive study has demonstrated that it is caused by oxidation of certain species of pyrrhotite. Self-heating was found to take place in two stages, first progressing from ambient temperature to 1000C with exothermic formation of elemental sulphur in the presence of moisture and oxygen, and proceeding to a second stage in which the sulphur oxidizes to SO2. Ignition may eventually follow. An apparatus was developed to assess the propensity of sulphide rock to self-heating and the potential for hazardous behaviour. The effect of temperature, time, rock size, and surface treatment with chemicals is presented. On the basis of these results, possible approaches for preventing a runaway reaction are suggested for further investigation.IntroductionThe oxidation of sulphide ore, waste rock or tailings backfill, which may cause hazardous conditions underground through heating, oxygen depletion and, in some cases, fire, is well documented (Farnsworth, 1977; Ninteman, 1978; Rosenblum et al., 1982; Reimers, 1987). The self-heating mechanism, nevertheless, has not been satisfactorily explained. The onset of heating from ambient temperature, in particular, has not drawn the attention of researchers. In the course of a previous study (Rosenblum et al., 1982) aimed at finding a method to assess the hazard from using sulphide tailings as backfill, the authors developed an apparatus to measure self-heating rate under controlled, near ambient conditions. This has since provided decisive information in backfill investigations in the Noranda Group.The test methodology devised for tailings failed to predict self-heating in a broken waste rock known to be active. Upon repeated testing, however, a freshly crushed sample began to exhibit appreciable heating and oxygen uptake (Fig. I). After 14 tests, the self-heating rate reached a plateau of about 4O°C/h and then fell progressively to < 5°C/h after 30 tests. It was hypothesized that weathering in the presence of moisture and oxygen during testing had caused the onset of self-heating. Active tailings, on the other hand, are exposed to natural surface oxidation and need not be weathered to show heating. Based on this finding, a new apparatus was developed capable of simultaneously weathering a sample of rock and monitoring its self-heating behaviour."

Jan 1, 1995

By Debasis Deb, Ankur Narang, G. P. Raja Sekhar, Kamal C. Das

"The stability of underground structures made (especially) in jointed rock mass is of the utmost important to designers, engineers and operators. Rock bolting is generally being practised to reinforce excavation walls and roofs by minimizing the movement of rock joints. This study proposes a new analytical model for the prediction of displacements, axial load and shear stress of a fully grouted rock bolt intersected by single or multiple rock joint(s). The governing equations for rock bolt are derived based on the assumption that shear force generated by slippage along the bolt-grout interface has a linear relationship with respect to the relative displacement at the interface. The proposed model is novel compared to the existing literature in the sense that it accounts for rock kinematics as an enriched sum of a continuous part (elastic) and a discontunous (jump) part due to the presence of joint planes. The complete solution of the bolt displacement, axial load and shear stress is derived as a function of the rock displacements, and, it comprises of Heaviside and/or Dirac delta functions in the solution space due to the jump at the intersection point(s) of bolt(s) and joint(s).Our novel solution is able to predict the elastic behavior of a grouted bolt, subjected to opening of the joint planes in pull-out test and also in excavated tunnel. These two cases also serve as benchmarks for grouted rock bolts with multiple joints and can be used for validation of numerical and experimental models. Further, the solution is applied to analyze the behaviour of grouted rock bolt as well as cable bolt installed in circular tunnel, with variations in rock joint(s) opening displacement, shear stiffness in the boltgrout interface and locations of the joint planes. Results indicate that for a specified grout stiffness, the maximum axial load developed in the bolt for discontinuous distribution of rock mass displacement is significantly higher than that of the continuous distribution of rock displacement. This paper provides an analytical interpretation of bolt beaviour considering multiple joint planes with variable joint opening displacements. The results obtained via the analytical method are in line with the corresponding investigation done by other authors. Thus, our analytical treatment can be used to understand bolt behaviour in realistic situations."

Jan 1, 2015

By R. Augsten, K. K. Feng, K. C. Cheng

"Methane emission and the desorption characteristics of vertical borehole coal samples of Greenhills multiple seams in the vicinity of Elkford, British Columbia were investigated. Three vertical boreholes at depths of 335 m, 380 m and 395 m were drilled along the west wall of an active open pit mine and 44 core samples from various coal seams were obtained for desorption studies.The results are presented in three tables and illustrated with six figures. The methane content was found to increase significantly at depths greater than 200 m. The adsorption equation by Kim and the properties of the coal have been used to predict the methane content and these were compared with the measured values by the direct method.IntroductionThe coalification of plant material in the sedimentary rock produces different ranks of co al and gases of various composition. The gas mixtures are retained in coal beds and rock strata under pressure. The gas content in a coal bed varies considerably depending on the rank of coal and the depth of overburden. The gas content is defined as the volume of gases adsorbed in a unit weight of coal materials. The net amount of gases remaining in a coal seam depends on the difference in the rate of generation of methane by geo-thermal heat and the rate of diffusion of the gases out of the coal bed through sedimentary cover. The major factor that causes the change in rank and the liberation of gases is the temperature, which is governed by the thermal conductivity of sedimentary cover and the heat flux through the underlying rocks. The rate of diffusion of the gases out of the coal is governed by the porosity of the overlying rocks and the degree of fracture. The measured values of gas content in major North American coal beds are from zero up to 24 cm3 per gram of coal(1,2). A gas content of 2.5 to 9 cm3/g is considered ""gassy"" and that of 9 cm3/g or higher is considered ""very gassy"" for underground mining(3). Normally, the gas mixtures released from bituminous coal in Canada contain 90 to 98% methane and the remainder is made up of CO2, N2, 02, H2, He and other hydrocarbons(4)."

Jan 1, 1984

By D. R. Boyle

"The Murray Brook precious metal (Au, Ag) gossan deposit is one of nine supergene mineral deposits that have developed over polymetallic massive sulfide deposits in the Bathurst Camp of New Brunswick. Reserves consist of 1.9 Mt of gossan containing 1.53 g/t Au and 65.9 g/t Ag. The supergene zone at Murray Brook consists of six distinct units: a) altered massive sulfide, b) pyrite-quartz sand, c) massive sulfide gossan, d) disseminated sulfide gossan, e) ferruginized wallrock, and f) leached bedrock. The massive sulfide gossan constitutes the main body of economic mineralization and consists of goethite, primary quartz, secondary amorphous silica, K-Fe-Pb-As- Sb-Ag hydrated sulfate and oxide minerals (beudantite, plumbojarosite, jarosite, bindheimite, scorodite), trace cinnabar, and cassiterite of primary origin. The order of stability of sulfide minerals during oxidation was pyrite _ arsenopyrite _ galena _ Bi-Sb sulfosalts _ chalcopyrite _ tetrahedrite group _ sphalerite.Two mass balance techniques (isovolumetric and conservative element) indicate elemental depletion in gossan relative to primary ore in the order: S>Cd=Zn>Cu=Mn>Na=Ca>Fe=Al=V> Mg= K=Se=Hg>Pb>Ti>Ag>Mo=P>Ba; and enrichment in the order Au>As>Sb>Si>Bi. Tin, as cassiterite, is conservative.Mineralogic and paleomagnetic data suggest that the gossan deposit formed under a warm temperate climate throughout the Pliocene. The rate of oxidation and the compositional nature of the ore was significantly influenced by: a) the position of the primary sulfide zone in a downward moving hydrological system, b) primary sulfide composition and zonation that promoted the development of a strong electrochemical oxidation system, and c) pronounced intercalation of carbonate and sphalerite mineralization in sulfide zones, which promoted rapid development of secondary porosity, permeability and downward flow of oxygenated groundwater.During progressive oxidation and physico-chemical erosion of the gossan zone, Au was transported downward in groundwaters, probably as a Auo colloid complex, to be concentrated in the lower horizons of the gossan profile. The precipitation of Au was greatly increased in the pyrite/arsenopyrite zones. Gold concentration was multi-cyclic with continual dissolution and nucleation of the metal until final concentration in the void structures of the gossan. Leaching experiments and microprobe analyses indicate that Au is present in the gossan as submicron composite sols of Au-Ag-silica."

Jan 1, 1995

By D. R. Boyle

"Abstract - The Murray Brook precious metal (Au, Ag) gossan deposit is one of nine supergene mineral deposits that have developed over polymetallic massive sulfide deposits in the Bathurst Camp of New Brunswick. Reserves consist of 1.9 Mt of gossan containing 1.53 g/t Au and 65.9 g/t Ag. The supergene zone at Murray Brook consists of six distinct units: a) altered massive sulfide, b) pyrite-quartz sand, c) massive sulfide gossan, d) disseminated sulfide gossan, e) ferruginized wallrock, and f) leached bedrock. The massive sulfide gossan constitutes the main body of economic mineralization and consists of goethite, primary quartz, secondary amorphous silica, K-Fe-Pb-As- Sb-Ag hydrated sulfate and oxide minerals (beudantite, plumbojarosite, jarosite, bindheimite, scorodite), trace cinnabar, and cassiterite of primary origin. The order of stability of sulfide minerals during oxidation was pyrite > arsenopyrite > galena > Bi-Sb sulfosalts > chalcopyrite > tetrahedrite group > sphalerite.Two mass balance techniques (isovolumetric and conservative element) indicate elemental depletion in gossan relative to primary ore in the order: S>Cd=Zn>Cu=Mn>Na=Ca>Fe=Al=V> Mg= K=Se=Hg>Pb>Ti>Ag>Mo=P>Ba; and enrichment in the order Au>As>Sb>Si>Bi. Tin, as cassiterite, is conservative.Mineralogic and paleomagnetic data suggest that the gossan deposit formed under a warm temperate climate throughout the Pliocene. The rate of oxidation and the compositional nature of the ore was significantly influenced by: a) the position of the primary sulfide zone in a downward moving hydrological system, b) primary sulfide composition and zonation that promoted the development of a strong electrochemical oxidation system, and c) pronounced intercalation of carbonate and sphalerite mineralization in sulfide zones, which promoted rapid development of secondary porosity, permeability and downward flow of oxygenated groundwater.During progressive oxidation and physico-chemical erosion of the gossan zone, Au was trans- ported downward in groundwaters, probably as a Auo colloid complex, to be concentrated in the lower horizons of the gossan profile. The precipitation of Au was greatly increased in the pyrite/arsenopyrite zones. Gold concentration was multi-cyclic with continual dissolution and nucleation of the metal until final concentration in the void structures of the gossan. Leaching experiments and microprobe analyses indicate that Au is present in the gossan as submicron composite sols of Au-Ag-silica."

Jan 1, 1995

By Catharine E. G. Farrow, David H. Watkinson

The Ni-rich Epidote Zone of the Fraser Mine at Sudbury is characterized by alteration assemblages rich in epidote, actinolite and magnetite which are roughly concentrically zoned and up to hundreds of meters in length. These zones form discrete areas within the igneous Footwall Breccia below the Sudbury Igneous Complex. The assemblages grade from epidotized Footwall Breccia matrix,through actinolite-bearing rocks with few vestiges of original Footwall Breccia textures, into the most intensely altered rock, a magnetite-rich facies. Pentlandite, pyrrhotite and pyrite are the dominant sulfdeminerals; chalcopyrite is minor. The strongest geochemical differences compared to original Footwall Breccia occur where epidote and actinolite dominate the alteration assemblage (EPDZact facies). The most intense alteration facies are typifed by low Cu/(Cu+Ni) values of <0.1. Strong positive correlations exist between Ni and Co, Cu and Au, and among Ir, Os, Ru and Rh. Chondrite-normalized Ni - PGE -Au - Cu plots display an overall increase in all elements from the least to the most intensely altered facies except for Au and Cu. As the proportion of actinolite increases in the EPDZact facies, the chondrite-normalized profles fatten considerably due to low Au and Cu contents. Abundances of Ni, Ir, Ru and Rhalmost ubiquitously increase with degree of alteration. In contrast, Cu, Zn, Ag and Au contents are low in almost all facies. It is suggested that saline hydrothermal fluids moved through the Epidote Zone formation and became more oxidized as they reacted with the Footwall Breccia and with footwall units, including ultramafic rocks. As this trend continued, magnetite-rich zones eventually developed. The occurrence of calcite locally in these magnetite-rich zones suggests that the pH of the fluid increased, causing Cu solubility to drop as activity in the hydrothermal system waned. Sudden changes in fluid composition such as this may have been due to an influx of fluid from bounding ultramafic rocks. Hydrothermal activity was focused along the margins of footwall ultramafic units. Individual hydrother-mal cells may have formed in the Footwall Breccia along pre-existing structures or sulfde concentrations. Metals may have been scavenged from sulfide-bearing Footwall Breccia and possibly Sublayer rocks. Nickel and some PGE were deposited in the Epidote Zone, whereas Cu, Au and Ag continued out-ward into the footwall where they were deposited as Cu-rich footwall veins. Copyright © 1996 Canadian Institute of Mining, Metallurgy and Petroleum.

Jan 1, 1996

By Catharine E. G. Farrow, David H. Watkinson

Abstract-The Ni-rich Epidote Zone of the Fraser Mine at Sudbury is characterized by alteration assemblages rich in epidote, actinolite and magnetite which are roughly concentrically zoned and up to hundreds of meters in length. These zones form discrete areas within the igneous Footwall Breccia below the Sudbury Igneous Complex. The assemblages grade from epidotized Footwall Breccia matrix, through actinolite-bearing rocks with few vestiges of original Footwall Breccia textures, into the most intensely altered rock, a magnetite-rich facies. Pentlandite, pyrrhotite and pyrite are the dominant sulfide minerals; chalcopyrite is minor. The strongest geochemical differences compared to original Footwall Breccia occur where epidote and actinolite dominate the alteration assemblage (EPDZact facies). The most intense alteration facies are typified by low Cu/(Cu+Ni) values of <0.1. Strong positive correlations exist between Ni and Co, Cu and Au, and among Ir, Os, Ru and Rh. Chondrite-normalized Ni - PGE - Au - Cu plots display an overall increase in all elements from the least to the most intensely altered facies except for Au and Cu. As the proportion of actinolite increases in the EPDZact facies, the chondrite-normalized profiles flatten considerably due to low Au and Cu contents. Abundances of Ni, Ir, Ru and Rh almost ubiquitously increase with degree of alteration. In contrast, Cu, Zn, Ag and Au contents are low in almost all facies. It is suggested that saline hydrothermal fluids moved through the Epidote Zone formation and became more oxidized as they reacted with the Footwall Breccia and with footwall units, including ultramafic rocks. As this trend continued, magnetite-rich zones eventually developed. The occurrence of calcite locally in these magnetite-rich zones suggests that the pH of the fluid increased, causing Cu solubility to drop as activity in the hydrothermal system waned. Sudden changes in fluid composition such as this may have been due to an influx of fluid from bounding ultramafic rocks. Hydrothermal activity was focused along the margins of footwall ultramafic units. Individual hydrothermal cells may have formed in the Footwall Breccia along pre-existing structures or sulfide concentrations. Metals may have been scavenged from sulfide-bearing Footwall Breccia and possibly Sublayer rocks. Nickel and some PGE were deposited in the Epidote Zone, whereas Cu, Au and Ag continued outward into the footwall where they were deposited as Cu-rich footwall veins.

Jan 1, 1996

By B. S. Fee, R. R. Klimpel

"Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores such as cobalt, platinum, gold, silver, etc. Most of the initial developments made with the process of froth flotation in the early 20th century were based on the processing of metal sulfide ores. It is the relative technical and economic ease of floating metal sulfide minerals with small amounts of appropriate frothers and sulfur containing (thio) collectors that has been a major economic driving force behind the rapid expansion in the supply of metals such as copper, lead, zinc, and nickel that has occurred in the world's economy in the last 50 years. This author estimates that 1.6 billion tons of sulfide ores are processed each year (1980) in the free world based on the froth flotation process.While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very large. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic locations having suitable grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 98% were known and manufactured in some form 25 years ago (1). This is clearly not typical of reagent development in other process related industries. In addition, the industrial scale practice of froth flotation applied to sulfide ores has proceeded since the 1920 s with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully at an industrial scale for many years."

Jan 1, 1987

By Terry A. Kirk

"This paper describes Derrick High Frequency Vibrating Screens and their process performance in grinding circuit applications. The benefits of screen sizing in closed circuit grinding are illustrated by case histories from a wide spectrum of mineral industries.INTRODUCTIONHigh capacity, reliable wet screening equipment is increasingly in demand by the mining and allied industries for separations finer than 1/.7 5mm (I/ mesh). The continuing search for improved energy efficiency and reduced steel consumption in comminution has Jed to greater scrutiny of every aspect of circuit design and operation, including the classifier efficiency (selectivity) function.Derrick Manufacturing Corporation has been developing high frequency screening machines designed specifically for grinding circuit applications since 1973. Successful installations are currently operating in iron ore, copper, coal, uranium, phosphate and other extractive industries."

Jan 1, 1984

By Edward R. T. Tiekink

Metal complexes continue to find therapeutic applications for the treatment of a wide variety of human ailments. Amongst these, gold(I) thiolates are used to treat rheutmatoid arthritis but, usually in combination with other treatment or as a last resort. New therapeutic uses of gold, both as gold(I) and gold(III), are being sought, including against cancer, the focus of the present contribution. It will be demonstrated that certain gold(I) complexes show exciting cytotoxicity profiles, rivalling those of cisplatin and standard chemotherapeutic agents. More importantly, anti-tumour activity is demonstrated in vivo, albeit in a mouse model. The presentation will give a brief overview of ?metals in medicine? with a focus on gold and will summarise some recent advances in the development of goldbased anti-tumour agents.

Oct 1, 2003

By Larry T. Yano, Gerald T. Handford

"Tailings Engineering has the responsibility for planning and coordinating tailings disposal and for ensuring the stability of the tailings dyke. The inclusion of both the planning and geotechnical functions in one team has resulted in greater efficiency in the design and planning process. Major challenges in tailings disposal facing Syncrude are in the areas of: (1) sludge disposal; (2) instrumentation and monitoring; and (3) liquefaction. Meeting these challenges will have a significant effect on the cost effectiveness and reliability of the mining operation.IntroductionThe size of the tailings disposal operation has led to the establishment of a large engineering team devoted solely to the disposal of tailings. This paper discusses: the tailings disposal operation, the role of tailings engineering, the organization, inter faces, and the present and future challenges with respect to tailings disposal."

Jan 1, 1990

Evolution of Maintenance Practices. Mechanization processes in underground mining operation became stronger in the 70?s - 80?s. Did maintenance practices changed over the years ? What are the the major differences of today?s practices? Speaker, Robert Hamilton Working at different positions into maintenance organization since over 25 years, Robert has accumulated a various and rich experience of ways to maintain equipment. He started as mechanic on construction projects and open pit mines. Early in his career Robert had supervision&apos;s responsabilities. He started in open pit operation (QCM) and after in underground operation with Lac Minerals, working at Bousquet, Doyon and Louvicourt mines, as U/G mechanical foreman and mechanical general foreman, and Maintenance superintendant with Cambior . He also experienced the underground mines equipments supplier&apos;s side working with Sandvik Canada.

Feb 1, 2008

By M. Mai Manga, M. Renaud, A. Faucher, C. Bazin

ABSTRACT Grinding circuits are operated in the presence of disturbances resulting from ore changes and equipment malfunctions, such as hydrocyclone overloading, worn or partly plugged apexes, surging, and internal liner detachment that may cause a loss of classification performance. Metal recovery losses in downstream leaching or flotation circuits justify the development of methods for rapid detection of hydrocyclone malfunction. Two methods are presented in this paper. The first method uses measurements from sensors available in a grinding circuit and the second uses vibration sensors installed on the individual hydrocyclones of a nest.

Jan 1, 2011

By J. S. Redpath

There appears to a definite trend toward the hoisting of greater tonnages from greater depths. There is every indication that this trend will continue and we feel that improvements in the design, fabrication and installation of mine guides will be of significant importance in the future. The present state of the art in mine-shaft guide systems is reviewed and the parameters required for improvement are derived in this paper. Finally, a proposal is made for a new steel guide system which meets the parameters and is economical to fabricate and install.

Jan 1, 1977

By E. Burdukova

Long-chain primary amines are utilized as collectors in potash ore flotation. Krafft temperature of long-chain amines in brine is about 100 °C, and this makes such amines practically insoluble in brine at room temperature. When dodecylamine, melted at 70 °C and dispersed in hydrochloric acid (or acetic acid) aqueous solution, is added to a KCl-NaCl saturated brine at room temperature, the amine precipitates. The experimental results demonstrate that the precipitating amine hydrochloride particles are selectively abstracted by KCl particles, but not by NaCl particles.

Jan 1, 2009