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By V. G. Minenko, A. L. Samusev

For efficient leaching of gold from sulfide ores and the oxidation of sulfides into sulfates, it is necessary to introduce a substantial amount of oxidant - atomic oxygen, hydrogen peroxide, ozone, potassium or sodium hypochlorite. In this paper, hypochlorite electrochemical treatment is proposed as an effective method of intensification of refractory ore leaching. The objective of the study is to increase the gold recovery from refractory ores with the application of chlorinated solutions that are modified by chemical or electrochemical treatment. The advantages of the use of the modified chlorinated solutions instead of cyanide for gold-sulfide leaching are considered to be the following: a high concentration of the oxidizing agent (hypochlorite ions) in solution helps to achieve a high speed of gold dissolution in the form of chloride complexes; the possibility to produce hydrochloric acid solutions that may be used for the recovering gold by electrolysis; the possibility of the processing of carbonaceous, arsenic-bearing and others kinds of refractory gold ores. The mechanism of the interaction of the modified chlorine solutions with gold-bearing sulfides (pyrite and arsenopyrite) was investigated. The efficiency of the application of chemical and electrochemical leaching method was proved. The main factor that limits the rate of gold dissolution is a high concentration of sulfate ions (SO4 2-) formed in the solution by the oxidation of sulfide minerals. The optimal parameters of chemical and electrochemical leaching method and effective ways to remove sulfate ions were determined. The method helps to increase the level of the gold recovery up to 87% from refractory ores. The intensification of leaching by chemical and electrochemical treatment opens up opportunities for the creation of modern efficient and environmentally-friendly technologies that may be applied for gold extraction from refractory gold ores.

Jan 1, 2016

In this paper, the use of electrochemical AC impedance spectroscopy (EIS) for in-situ diagnosis of high temperature PEM fuel cells is described. EIS was performed on single cell hardware, built in-house, using a fuel cell test station modified to operate up to 120°C and 30 psig. Electrode kinetics, membrane resistance, porous-layer and mass transport impedance of MEAs, assembled in-house, were measured at various current densities for the purpose of validating the test station and cell hardware, and for optimizing MEA and flow-field designs. The effect of temperature and humidity on the various impedances was investigated. Impedances obtained at lower and higher temperatures were compared to obtain insight into the issue of water management in operating fuel cells.

Jan 1, 2005

By Marcel Laflamme

Underground mine workers are continuously exposed to many sources of noise and vibration in carrying out their tasks. Even though the vibration and sound level of some mining equipment is documented, it was found that this information is not readily available for many pieces of equipment and that it is not considered when choosing equipment or planning tasks. As well, health problems related to vibration and noise are among the main causes of work-related compensations in mines. Considering those facts and although no Canadian regulations could force employers to limit the exposure of their workers to vibration, the Société de recherche et développement minier (SOREDEM) decided to address the problem by including vibration among their research priorities. In order to characterize noise and vibration sources of different mining equipment used for underground mining, the authors have performed a study in 8 different mines in Québec on 28 different mining equipments. The paper will present the procedure that was followed to select the list of equipment to be evaluated and the low cost acquisition system developed to perform multichannel time domain acquisition in harsh environments. It will also present the results obtained for the noise level produced by the equipment as well as for the hand-arm and whole-body vibration exposures.

Aug 1, 2013

By L. C. Bird, A. I. A. Salama, M. W. Mikhail

"A program of sampling in Canadian washeries was carried out by the Coal Research Laboratory (CRL) over a period of two years to evaluate the performance of various density separators in these operations(1). Operating data on quality and recovery of products, separation efficiencies and cutpoints for each separator type are compared with data predicted using computer models developed from previously existing data.The Heavy Medium Cyclone (HMC), Heavy Medium Bath (HMB) and jig models give predictions of recoveries and qualities of products for sizes coarser than 10mm with absolute value of error mean less than 7 per cent, whereas predictions of the Water Only Cyclone (WOC) model for -10mmfine coal are with absolute value of error mean less than 8 per cent. Although less accurate for -10 mm fine coal, the HMC and jig models produce predictions of recoveries with absolute value of error mean less than 2 per cent and quality of products predictions with absolute value of error mean less than 28per cent which are generally acceptable because more liberation in the fine fractions appears to compensate for any difference in the accuracy of separation, resulting in good prediction of the clean product. However, ash content of the actual fine refuse was lower than predicted. The difference is attributed to the highly friable nature of Canadian coals which tends to cause an excessive loss of coal fines to the fine refuse.A comparison study based on quality and recovery of products is made between the Dutch State Mines (DSM) models and the recently modified models, both produced comparable predictions."

Jan 1, 1985

By Robert B. Parsons

"Recent amendments to the so-called ""flow-through share"" rules in our income tax legislation have given a new lease on life to this once popular investment vehicle.The original ""Dome funds"" for exploration in the Beaufort sea recognized the usefulness of flow-through share arrangements, but, by and large, flow-through shares have until recently not been able to compete effectively with other tax shelter instruments-mainly , limited partnership interests.This month, Tax Notes examines the new tax rules for investing in flow-through shares.IntroductionRecent income tax amendments have helped to rekindle investor and industry interest in so-called ""flow-through shares"" as a mechanism for mobilizing funds toward mineral exploration and development activities. In competing for investment funds, flow-through shares have historically placed a distant second in terms of investor popularity, being unable to compete with the more attractive features of limited partnership arrangements. However, primarily as a result of favourable changes to the income tax law, flow-through shares may be well poised for a comeback.This article examines the income tax considerations pertaining to issuing and investing in flow-through shares, and compares the former and the new income tax rules relating thereto. Of course, inasmuch as a flow-through share issue can involve various contractual arrangements between the investor and the issuer, there is a wide range of legal and business factors that generally have to be taken into account. A discussion of these matter s is beyond the scope of this paper, which is restricted to a summary of the special income tax factors attributable to flow-through shares."

Jan 1, 1983

By H. A. Steane

"This presentation will cover some crushing and grinding considerations in plant design and operation. There is nothing basically new in the subject matter presented here; however, the viewpoint and interpretation may be novel and some of it controversial.The presentation is based on experience gained in the metallurgical design of plants for Placer Development Limited from 1960 to the present. The plants are listed on SLIDE No. 1.These plants have conventional crushing, followed by rod mill and ball mill grinding. I would like to think that autogenous grinding plants are the answer for eliminating secondary crushing plants and reducing the capital and operating costs; however, the autogenous plants seem to have some problems of their own and I have had no experience with them.Fineness of Grinding RequiredFirst let us consider the fineness of grinding required.The fineness of grind will affect the recovery obtained, and the relationship must be determined. If the grinding is too coarse, some of the valuable mineral, either free or locked in middling grains, will not be floated. However, as grinding is probably the greatest single operating cost, we do not want to grind any finer than is justified economically.Finer grinding should not be carried beyond the point where the net smelter return for the increment saved becomes less than the added operating cost. Even this fineness will be beyond the economic limit because of the additional capital cost for the equipment needed to achieve it. These factors will be discussed in more detail later.The Bond work index is the commonest method of expressing a measure of the effort that has to be expended in grinding. It is the kilowatt-hours required to reduce a ton of rock from an infinite size to 80% passing 100 microns. The formula allows a calculation to be made for other size ranges."

Jan 1, 1976

By Y. Wang, S. Liu, K. Jiang

China is the world’s largest producer and consumer of non-ferrous metals. In recent years, significant progress has been made in research and commercial application of pressure leaching technology to produce some nonferrous metals. Research and development activities include a low-pressure, low-temperature leaching technology for complex copper concentrates, a hematite process to treat pressure leach solutions from zinc concentrates, and a pressure leaching - solvent extraction process for molybdenum and rhenium. Examples of commercial practices include a two-stage pressure leach technology in a zinc plant in Danxia, a direct leaching technology in a zinc plant in Zhuzhou, a high pressure acid leaching process for a nickel laterite project in RAMU, and a pressure leaching technology to treat arsenic sulfides in a plant in Guixi. The development of pressure leaching technology for the recovery of lead, vanadium, gallium, germanium, and precious metals in China will also be described in this paper.

Jan 1, 2015

"In 1989, improved milling and recoveries resulted in a record production of 7 722 200 Ib Nb2P5 at the Niobec Mine. At the beginning of 1990, proven and probable reserves stood at 11 952 000 tons grading 0.65% Nb205. This year, the world steel market has slowed and as a result, niobium consumption is down. Consequently, production at the mine dropped 14% for the second quarter.IntroductionThe Niobec Mine is located approximately nine miles northwest of Chicoutimi, Quebec, (Fig. I). After ten years of production, the Niobec Mine is the only operating niobium mine in North America and the second largest producer of niobium in the world after Brazil.The Niobec Mine is a joint venture held equally by Teck Corporation and Carnbior. Teck Corporation operates the mine while Cambior is responsible for marketing and selling the niobium pentoxide concentrate (Nb2P5).The deposit was discovered in 1964 following an airborne radiometric survey in search of uranium. However, mining of the deposit did not begin until 1976. The initial production capacity of 1500 tons per day was increased several times to the present capacity of 2450 tons . The deposit is considered as an ample source of pyrochlore concentrate for at least 15 years at the present production rate.Niobium, also known as columbium, is a greyish-white metal used primarily as an alloy in various steels and other metals. The use of niobium in alloys heightens tensile strength and improves corrosion resistance while at the same time reducing the weight of the alloy. Niobium is also used in the fabrication of steel for the aerospace, automobile, and oil and gas pipeline industries. Although substitutes for niobium exist, their use often results in diminished performance and net cost increases."

Jan 1, 1990

By Brad Gunn

Our program has been dictated by the necessity of broadening our markets through converting potash chemicals to fertilizers more in demand than potassium chloride. To this end, we have undertaken studies on methods of preparing potassium sulphate or potassium magnesium sulphate by rather simple conversion. When mixtures of sodium and magnesium sulphates are reacted with potassium chloride, the ratio of sodium to magnesium is quite pertinent. A range of ratios has been selected and the conversions of potassium chloride to sulphate have been assessed by cyclic operations. Results from experimental procedures for producing potassium nitrate are also avail-able. The influence on recovery for systems including organic solvents is assessed for a variety of situations.

Jan 1, 1971

By K. Ski

Three series of membranes were prepared by blending polyvinylidene difluoride (PVDF) or poly(vinylidene difluoride-co-hexafluoropropylene) [P(VDF/HFP)] with sulfonated poly(ether ether ketone) (S-PEEK), sulfonated polysulfone (SPU), or sulfonated polystyrene, in order to develop novel PEM membranes suitable for melt processing. By means of incorporating fluorine-containing block copolymer into the blended system, acceptable proton conductivity and low water containing PEM materials were prepared. The effect of sulfonated poly([vinylidene difluoride-co¬hexafluoropropylene]-b-styrene) block copolymers as compatiblizers was investigated: conductivity was enhanced for S-PEEK/fluoropolymer blends. Microstructures for the blends were examined by transmission electron microscope (TEM) and scanning electron microscope (SEM).

Jan 1, 2005

By Alastair J. Sinclair, Asger Bentzen

Abstract - A simple linear model for interpreting duplicate analytical data for a range of values is described. The conceptual model can be quantified using a reduced major axis model rather than the inappropriate traditional least squares approach. The reduced major axis parameters (slope, intercept) and associated errors provide a quantitative basis for: (1) defining bias, and (2) comparing results within laboratories, between laboratories and between sampling/analytical methods. Application of the linear model must be preceded by extraction of obvious outliers or small groups of values that have an inordinately large impact on values of the statistical parameters defining the model. In addition, consideration must be directed to the presence of more than one population of sampling error, as in the case of multiple styles of mineralization being sampled.

Jan 1, 1998

By R. S. Boorman, L. J. Surges

"Ore reserves at Brunswick Mining and Smelting totalled 101(106) tonnes grading 97 grams tonne-1 silver at the end of 1981. Silver is recovered as a byproduct in the copper, lead and bulk concentrates and is lost to the zinc concentrate and plant tailings. Recovery of silver for the year 1981 was 58,5%. The loss of potential revenue is in excess of Cdn $50(106) year-1 at current silver prices and milling rates.Silver is known to occur in solid solution in galena and as tetrshedrite. Evidence indicating the identity and significance of a third silverbearing mineral assemblage is reviewed and the distribution of silver losses is detailed. Characteristics of each the carrier which may be responsible for observed distribution, including size, surface quality, liberation and composition are considered, and the third carrier is shown to have possible implications for improved silver recovery by other Canadian producers as well as by Brunswick Mining.Silver tetrahedrite tailings distributed tailings and losses occur as coarse, free and as very fine unliberated galena in losses of the third carrier are between the zinc concentrate and are concentrated in the finest size fractions. Process alternatives for enhanced silver recovery by physical and chemical means are reviewed."

Jan 1, 1983

By Jef M. van der Zel

The idea of veneering rich yellow gold alloys with a low-melting, high expansion porcelain, has been a driving force for metal-ceramic research in dentistry. Since 1990, with the advent of the Carrara System (1), a low firing temperature veneering ceramic allowed the use of gold alloys with a lower content of platinum and palladium than the traditional gold-platinum ceramic alloys and thus a richer yellow color. The earlier high gold-platinum ceramic alloys contained at least 12 percent platinum group metals to allow for dimensional stability when firing the relatively high melting traditional ceramics. This amount of platinum group metals gave the alloys a greyish tinge which made them unattractively steel colored and difficult to present as ?high gold alloys?. In the 90?s the new generation universal low-melting, high expansion gold alloys gained market share at a steady rate, starting in Germany, and have now for a great part replaced the traditional gold-platinum alloys (Fig. 1). The first alloys contained 75 wt-% gold and up to 9 wt-% platinum or palladium. They provided a metal-ceramic system in which a gold alloy having an aesthetic yellow colour is fired on with a dental porcelain tailored thereto. The tailoring is meant in respect of compatible coefficients of thermal expansion, and compatible melting or softening points of metal and porcelain. The major concerns in the beginning were with thermal stability during firing, the thermal expansion coefficient of the alloy against the porcelain (compatibility) and the yellow color. In a later stage, alloys with gold contents in the range of 50-60 wt-% gold were developed in analogy with the pale yellow gold-reduced alloys that were formerly used for crown & bridge work and the white gold-reduced alloys used for metal veneering where gold was substituted in part by palladium.

Oct 1, 2003

By R. D. Clark, Heinz Althoff

"The advantages and disadvantages of a mobile crushing plant with a belt conveyor system compared to a stationary crushing plant with haulage by heavy trucks are described. The capital expenditure and operational expenses are compared and the individual cost factors are explained.Every open-pit mine or quarry, however, is different and has to be assessed and analyzed individually in order to find the optimum and most economical solution. Due to the high truck haulage costs, many open-pit mines tend to the more economical solution with mobile crushing plants and continuous belt conveyor operation. IntroductionFor the material haulage within an open-pit mine the heavy truck is the most flexible means of transportation. However, due to the high fuel consumption, wear and labour costs, it is also the most expensive one. Therefore, many open-pit mines are looking for alternative solutions for the material transport A mobile belt conveyor system is most likely to meet the requirements because it is efficient and economical with regard to the power consumption. The material , however, has to be crushed at the mining face to lumps which are suitable for conveyor operation. For this purpose a mobile crushing plant is required.In some open-pit mines the crusher is also used to crush the material to lumps of smaller particle sizes in order to save crushing costs in the preparation plant.Although a cost comparison between a system with stationary crusher and truck haulage and a system with mobile crusher and belt conveyors favours in most cases the belt conveyor system, up till now only a few open-pit mines and quarries benefit world-wide from the advantages of the mobile crushing plant with continuous belt conveyor operation."

Jan 1, 1986

By G. Heinrich

Size by size recoveries and pentlandite liberation data from image analysis were used to evaluate plant tests of pentlandite flotation and pentlandite/pyrrhotite middlings flotation from pyrrhotite in a circuit treating low grade magnetic intermediates at the Strathcona Mill. Raising the pH with lime gave improved selectivity for liberated pentlandite but there were major losses of middlings unless the circuit was pulled hard indicating the need for finer regrinding. However, size by size recoveries at the normal regrind showed that, while the pentlandite recovery improved in the finer ranges, the pyrrhotite recovery also increased leading to little selectivity in any size range. A much finer regrind gave greatly improved selectivity in all size classes suggesting that 'surface polishing' as well as liberation effects were being seen. Using liberation data from image analysis of the products, surface polishing accounted for over 99 % of the improvement in selectivity. Unreasonable recoveries of more highly liberated particles were observed for liberation analysis by volume fraction but showed the expected trend in recovery when the analysis was based on surface liberation. The surface liberation data showed that the improved pentlandite recoveries after finer regrinding were to more than 90 % caused by improved liberation with the remainder contributed by surface polishing effects. The practical implications of this work are that i) the comparison of particle grades with product grades allows to estimate the effect of regrinding on co-flotation/entrainment, ii) surface liberation models should be used for evaluating recoverabilities by flotation and iii) regrind circuits may be set up to ensure that all particles go through the mill or that only the coarse fraction is reground depending on the surface scrubbing found to be required.

Jan 1, 1993

By M. Saghaeian, R. Hejazi, Sheikhzadeh. V., A. Abazarpoor

"The morphology of iron ore concentrate particles is known to influence several properties of iron ore pellets, including drop number, compression strength, and reduction properties. This study aimed to determine why grinding methods and particle morphology influence green pellet properties. Particle morphology can be directly related to the grinding method: ball mill (BM) or high pressure grinding roll (HPGR). In this research, two iron ore concentrates, namely Gol-e-gohar and Ardakan, were ground by a BM and HPGR to produce a suitable Blaine value for pelletizing (nearly 1,900 cm2/g). Morphological features of particles were subsequently investigated by sieving and cyclone sizing for particle size distribution, the Brunauer-Emmett-Teller (BET) and Blaine methods for specific surface area (SSA) and scanning electron microscopy (SEM) for the evaluation of shape descriptors. It was concluded that HPGR resulted in generating a greater proportion of fine particles compared to ball milling. The BET results showed that the SSA of HPGR pellet feed was higher than that from the BM due to a greater proportion of cracks, indents and bulges and the creation of new fracture surfaces. SEM images processed by ImageJ software indicated that particles ground by the BM tended to be more rounded and less elongated than HPGR particles because of the breakage model of each method. To show the effect of morphology on the green pellet quality, green pelletizing was performed in a pilot pelletizing disc. The HPGR pellet feed had a higher drop number in comparison to the BM pellet feed, due to the angular and rough shape of the particles.INTRODUCTIONThe demand for fine or ultra-fine particles as feed material for ceramics, composite materials, and mineral processing is on the rise (Kotake, Kuboki, Kiya, & Kanda, 2011). The final concentrate of iron ore processing plants (filter cake) has a size range roughly between 0 and 300 µm, depending on the liberation of Fe containing particles (Meyer, 1980). In general, to produce a suitable pellet feed of appropriate quality, Blaine numbers of 1800–2000 cm2/g should be achieved (Van der meer, 2015)."

Jan 1, 2016

By P. H. Oliver

The current theory of how explosives fracture rock offers no explanation for the improvements in fragmentation and the reduction of fly rock and back break which occurs when an adequate inter-row delay time is used. Nor does it explain why fragmentation improves with rectangular drill patterns where the hole spacing may be as much as four times the hole burden. A new theory of rock fragmentation is proposed which builds on the existing theory. It offers rational explanations for the many blast effects which the current theory does not address and the need for adequate timing delays between rows. It improves on the design of inter-row delay timing by defining the length of free-blasted face needed in front of a hole when selecting the delay element which will provide an adequate inter-row delay time.

May 1, 2003

By J. G. Whellock, J. W. Matousek

"In spite of their importance to the economics of many metallurgical projects, the values and technical aspects of precious metals losses in smelting operations are relatively undeveloped subjects. This paper attempts to provide a condensed overview of precious metals losses in pyrometallurgy, using some of the limited production and experimental data available. It is intended for three general groups of metallurgists - operators attempting to improve recoveries, cost-accounting engineers making estimates of metal losses in new or modified smelting plants, and research engineers and students.Distribution coefficients of precious metals, the ratios of metal or matte assays to slag assays, are convenient means of comparing smelting processes and the tendencies of the different metals to enter the slag. For silver in lead blast furnaces this ratio is approximately 250;for gold and the platinum group metals in submerged-arc, electric furnaces it is 900. The ratio for gold and PGMs in converters varies with the degree of oxidation of the charge. The over-all cycle-average for industrial installations is tentatively estimated to be 35.IntroductionThe magnitude of precious metal s losses in pyrometallurgical operation s is difficult to assess. The primary producers are generally reluctant to discuss the subject. The by-product producers (the large sulphide smelters o f copper , nickel, and lead) might be more willing to disclose information, but these have precious metals in their feeds in such dilute quantities that accurate accounting is extremely difficult. It is not until numerous processing steps have been taken t hat the precious metals become sufficiently concentrated to permit acceptably accurate sampling and analytical procedures to be followed. Thus, knowledge of the precious metals contents o f the intermediate mattes and slags may simply be unavailable."

Jan 1, 1990

By Lluís Fontboté, François Lieben, Robert Moritz, Massimo Chiaradia

"Abstract - The Las Cañas stratabound Pb-Zn deposit, in northern Chile, is hosted by the Lower Cretaceous back-arc marine carbonate and subordinate andesitic volcanic and volcaniclastic rocks of the Chañarcillo Group, one hundred meters above a thick pile of magmatic arc-related basaltic to andesitic lavas of the Bandurrias Formation. The deposit is not obviously related to local intrusive rocks, which include an Upper Cretaceous granodiorite batholith and dacite dikes emplaced along normal faults. Stratabound Pb-Zn-Ba mineralization is constrained within several porous stratigraphic units, including a 50 cm thick lapilli-tuff and less well mineralized algal-laminated limestone. Veins that locally carry some Pb-Zn occur along normal faults that displace the stratabound orebodies, and this study shows that they represent a different type of mineralization.Fluid inclusion microthermometric data for ore-stage quartz and barite indicate saline fluids (6 to 30 wt% NaCleq) and a formation temperature in the range of 80°C to 160°C for the stratabound mineralization. Two types of fluid inclusions occur in the veins. The type I vein fluid inclusions indicate less saline (average 7 wt% NaCleq) and slightly higher temperature fluids, in the range of 110°C to 220°C. The type II vein fluid inclusions indicate highly saline fluids (25 to 40 wt% NaCleq) which might be related to minor copper mineralization.206Pb/204Pb, 207Pb/204Pb and 208Pb/204Pb ratios for galenas from the stratabound mineralization, with mean values of 18.42, 15.59, and 38.25, respectively, fall within the range of stratabound polymetallic deposits of northern and central Chile, and indicate a Lower Cretaceous magmatic rock source. Galenas from veins have slightly more radiogenic 206Pb/204Pb, 207Pb/204Pb and 208Pb/204Pb ratios, with mean values of 18.47, 15.61, and 38.34, suggesting two distinct episodes of mineralization.?34S values for galena and barite range from negative values (-31.2‰ to -12.5‰ and -7.7‰ to +12‰, respectively) in the mineralized tuff to higher values in the algal-laminated limestone (-2.5‰ to +1.2‰ and +13.2‰ to +17.4‰, respectively), close to Lower Cretaceous marine sulfate values for barite, and homogeneous values for the galena from veins (-12.0‰ ±1‰). Changing redox conditions can partly account for ?34S variations in the tuff. Sulfur sources include magmatic rock sulfur (-1.5‰ to +9.7‰) as well as sedimentary sulfides and marine sulfate."

Jan 1, 1999

By Nadarajah Nanthananthan

Mine backfill has been widely used as a means of ground support. While proper design of backfill is important to ensure ground stability, understanding the failure mechanism and the various influencing factors in a filled block is essential. This paper presents a modified sliding failure model of a confined backfill block. The model considers both the gravity load of the backfill itself and the overlying load from fragmented ore or backfill on the top. A series of new equations were derived and the relationships between factors such as factor of safety, friction angle, cohesion and the exposed stope height were developed. This modified model will be a very useful tool in backfill design, for example, to optimize the design of backfill for a given stope geometry or to determine the maximum backfill height for a particular type of backfill.

May 1, 2007