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By N. Katsikaros, R. Agius, B. Adaszynski

"With stricter limits globally being imposed on cyanide discharge levels in plant tailings, having a reliable and well designed cyanide detoxification circuit is critical for any operation. The INCO SO2/Air method is most commonly used for cyanide detoxification where sulfur dioxide and oxygen combine to oxidize cyanide to the much less toxic cyanate. This paper discusses the steps involved with the development, from laboratory through to design, of a robust cyanide detoxification circuit. The goal is to produce a cyanide detox circuit that reliably meets discharge target with the lowest operating and capital cost.Design begins in the laboratory where continuous tests simulate full scale operation of a detoxification circuit. Detailed design relies on several design parameters, based on theory and experience, to ensure the reactor can satisfy requirements of the cyanide destruction reaction. Selection of the correct reagents and their supply to the reactor is important to an operable process with minimal operator workload. The operability of the system relies on correct process control strategies to maintain optimal reagent addition and chemical conditions in the reactor. These design factors are critical to providing a robust plant that will consistently meet cyanide discharge levels, ensuring uninterrupted production and plant profitability."

Jan 1, 2014

"INTRODUCTIONThe Department of Energy, Mines and Resources has recently developed two automatic controls for use in the cyanidation circuits of Canadian gold mills. These controls are the alkalinity probe and the Continuous Cyanide Titrator both of which are discussed in the paper ""Instrumentation in the Cyanide Process"".(1)In May 1966 these process controls and associated recording equipment were installed in the Bralorne mill by two Scientific Officers from the Department. The object was to determine if the equipment was amenable and applicable to this particular circuit. The purpose of this paper is to discuss the performance of the equipment over a test period of 4 months.CONCLUSIONSThe results of test work to date indicates that both the Conductivity Probe and the Continuous Cyanide Titrator would be an as set to the operation of the Bralorne mill.The conclusions are generalized as follows:-(1) At a milling rate of 100, 000 tons per year the expected reductionin reagent consumption would give a direct saving of 37, 660 lbs of lime@$0.030 per lb and 15, 790 lbs of sodium cyanide@ $0.175 per lb. Such cost savings would pay for the equipment and installation costs within a 30 month period"

Jan 1, 1967

By E. Yalçin, A. Ünal

In transitional mineralization, the sill of the underlying variogram is relatively easily estimated, but the estimation of the nugget effect may be more difficult. This study aims to determine for massive copper deposits just how sensitive the estimations of ore reserve, economic copper content, total net value, and ultimate pit limits are to the errors of estimation in the nugget effect of the spherical variogram model. A massive copper deposit which has a log-normal distribution and a spherical variogram was simulated on a rectangular grid using the turning bands method. By using the three dimensional random kriging method, the mean block grades were estimated at four different sampling densities. The ultimate pit-limits were determined by using the two-dimensional dynamic programming algorithm. To calculate the estimation errors, the values obtained by using the actual nugget effect and structured component values were deducted from the estimations obtained using the incorrect nugget effect and structured component values. For isotropic massive copper deposits, it was found that the ratio of the nugget effect to the structured component of the spherical variogram has no significant effect on either the block grades estimated by the kriging method or on the ultimate pit limits.

Jan 1, 1992

By P. S. Daly

"The purpose of this paper is to describe the types of instrumentation used at Sherritt Gordon's Ruttan Mine to monitor the"" effects of mining. The aim of the monitoring program is to establish parameters to achieve a maximum ore recovery with minimum dilut ion in a safe environment.The copper-zinc Ruttan Mine is in northwestern Manitoba, approximately 760 km from Winnipeg (Fig . I) . Production commenced in 1973 from the open pit at the rate of 10,000 tons per day. Pit bottom lies 58 metres above the first mine level. Underground development has overlapped with open-pit mining throughout. The open pit was phased out in the latter part of 1980.Underground production began in early 1979. The mining method employed in the first lift is sub-level blasthole open stoping with delayed hydraulic filling. Large-diameter (6.5-in.) holes are being drilled in the crop stope area and are also planned for some other areas for vertical retreat mining into a slot. GeologyThe orebody consists of numerous sulphide lenses of variable size, geometry and geology . The lenses occur within volcanoclastic host rocks which exhibit a strong variability in intensity and type of alteration and structural characteristics. Stope design and sequencing were based on these complex geological factors."

Jan 1, 1981

By G. E. Norman, R. E. Michelutti

"One of the major objectives of Falconbridge has been and continues to be, to minimize the impact of its operations on the environment around its Sudbury operations.The approach to achieving this objective and the results, are discussed under the following topics .I . Sulphur Dioxide Emission Improvements.2. Effluent Treatment Improvements .3. Effect of the Envrionmental Improvements.Sulphur Dioxide Emission ImprovementsHistoryThe rejection of pyrrhotite in the concentrators is one method of reducing the amount of sulphur to the smelter and therefore the sulphur dioxide emissions. In the 1950s and 1960s, Falconbridge developed the capability to remove 54% of this pyrrhotite. Numerous attempts were made to increase this amount by treating the pyrrhotite for recovery of nickel and in some cases elemental sulphur. Processes tested and operated were: hydrometallurgical pilot plant, roast-leach commercial operation and nickel-iron refinery. None of these processes proved to be commerically successful for both technical and economic reasons. Falconbridge therefore redirected its efforts to reducing sulphur emissions by improving its smelting process under The Smelter Environmental Improvement Project. A major component of this project was the installation of a sulphuric acid plant to recover the S02 in the roaster off-gases."

Jan 1, 1984

By B. Skrotzki, E. Augenstein, C. Rockenhäuser

"The aluminium alloy 2618A is an Al-Cu-Mg alloy with additions of Fe and Ni, which is designed for long-term operation at elevated temperature in transportation and aerospace industries. The alloy degrades due to coarsening and dissolution of GPB zones and coarsening of S-Phase precipitates (Al2CuMg). This work investigates the overageing process at the application relevant temperatures of 190 °C for durations ranging up to 25,000 h. The nm-sized precipitates were selectively imaged along the <001>a direction of the Al matrix using dark-field transmission electron microscopy. The resulting images were then evaluated and allow the statistically relevant determination of radii distributions and the average radius of the precipitates dependent on temperature and heat treatment duration. Ostwald ripening of the S-phase does not describe the ageing process adequately. Instead, the evolution of GPB zones during the initial ageing regime needs to be considered in addition to the coarsening process of the S-phase. DSC measurements were performed to determine the presence of GPB zones in different ageing states. This will allow a quantitative description of the ageing process and a comparison to samples aged under external load.INTRODUCTION Alloy 2618A is an Al-Cu-Mg alloy, which is part of the 2XXX series of aluminium alloys. It is used in aerospace and transportation industry due to its desirable material properties for long term operation at elevated temperatures (Polmear, 2006). Fe and Ni are added to form intermetallic compounds, which retain microstructural stability at elevated temperatures (Bergsma, Li, & Kassner, 1996; Elgallad, Shen, Zhang, & Chen, 2014; Oguocha, Yannacopoulos, & Jin, 1996; Wang & Yi, 2006). The size and distribution of the nm-sized S-phase formed from the supersaturated solid solution (SSSS) during heat treatments mainly controls the strength (Khan, Starink, & Yan, 2008; Liu, Zhang, Ding, Sun, & Chen, 2003; Parel, Wang, & Starink, 2010; Shih, Ho, & Huang, 1996; Wang, Starink, & Gao, 2006). The S-phase has the composition Al2CuMg and is orthorhombic under equilibrium conditions (Heying, Hoffmann, & Pottgen, 2005; Perlitz & Westgren, 1943). Early experiments proposed the following decomposition of the S-phase sequence during heat treatments (Hardy & Heal, 1954; Silcock, 1961):"

Jan 1, 2018

"Adopted in 2003, theses guidelines are intended to assist the Qualified Person(s) in the planning, supervision, preparation and reporting of mineral resources and mineral reserve estimates. The report is divided into two broad sections “General Guidelines” appropriate for most metals and “Guidelines Specific to Particular Commodities” for potash, industrial minerals, uranium and coal. “These guidelines are reference in the Companion Policy of NI 43-101 with a statement that a qualified person classifying a mineral deposit as a mineral resource or mineral reserve should follow them.”Executive SummaryThe following is a summary, in table format, of the main elements of the Estimation Best Practice Committee's report ""Estimation of Mineral Resource and Mineral Reserve Best Practice Guidelines"". While the summary table is provided for convenience, the Committee recommends that the report be read in its entirety and the table summary not be used as a stand alone document. These guidelines are not intended to be either prescriptive or exhaustive. They do not preclude innovation. Preamble: These guidelines have been prepared by the Canadian Institute of Mining and Metallurgy and Petroleum (CIM) led Estimation Best Practices Committee. They are intended to assist the Qualified Person(s) (QP) in the planning, supervision, preparation and reporting of Mineral Resource and Mineral Reserve (MRMR) estimates. All MRMR estimation work from which public reporting will ensue must be designed and carried out under the direction of a QP in accordance with National Instrument 43-101 (NI 43-101) and related forms. A QP is defined in NI 43-101 as ""an individual who is an engineer or geoscientist with at least five (5) years of experience in mineral exploration, mine development, mine operation, project assessment or any combination of these; has experience relevant to the subject matter of the mineral project and technical report; and is a member in good standing of a professional association"". Disclosure of MRMR estimates is to be made in accordance with industry standard definitions approved by the CIM (the CIM Standards adopted by the CIM Council in August 2002) which have been incorporated by reference into NI 43-101."

Jan 1, 2003

By A. Boucher, D. Marcotte

"Abstract- The accepted practice for the estimation of thin (2D) vein deposits recommends the use of the grade x thickness service variable (i.e., the accumulation). Grade estimates are obtained indirectly by the estimated accumulation/estimated thickness ratio. This practice stems from the varying support (thickness) problem and the resulting non-additive nature of the grade variable. We compare the actual performance of the direct grade estimation approach used by some practitioners to that of the indirect approach using accumulation. Our simulated and real data indicate that the direct approach is more accurate for point grade estimation where the grade-thickness correlation coefficient is positive (and vice-versa). Moreover, the relative gain of the direct method increases with the (positive) correlation coefficient. This finding contradicts common thinking that the indirect approach should be the preferred method where grade-thickness correlation is strongly positive. Also, for a given positive grade-thickness correlation, the relative gain of the direct method increases with the coefficient of variation of the grade and thickness. Résumé - L’estimation de veines minces (2D) minéralisées est normalement réalisée à l’aide de la variable auxiliaire épaisseur x teneur (i.e., accumulation). Les estimés des teneurs sont alors obtenus indirectement du rapport accumulation estimée/épaisseur estimée. Cette pratique découle du problème du support variable pour la teneur, l’épaisseur de la veine, qui fait que la teneur est alors une variable non-additive. Nous comparons les performances de l’estimation directe de la teneur, utilisée par certains praticiens, à celle de l’approche indirecte. Pour l’estimation ponctuelle des teneurs, nos données simulées et réelles indiquent une meilleure justesse de l’approche directe quand la corrélation teneur-épaisseur est positive (et vice-versa). De plus, le gain relatif de la méthode directe s’accroît avec la valeur de cette corrélation (positive). Ce résultat contredit l’opinion répandue que l’approche indirecte devrait être utilisée en présence de forte corrélation positive de la teneur et de l’épaisseur. Nous montrons également que pour une corrélation positive teneur-épaisseur donnée, le gain de l’approche directe augmente avec les coefficients de variation de l’épaisseur et de la teneur."

Jan 1, 2001

By P. Amelunxen, E. Herrera, L. Hill, J. Vanderbeek

It is well-known that the particle breakage mode that occurs in a semi-autogenous mill is significantly different from that which occurs in a ball mill, enough so to warrant very different lab scale breakage tests and scale-up models. This phenomenon, at the plant scale, leads to changing process constraints depending on the particular properties of the ore. For example, hard or coarse ores might bottleneck the SAG/AG mill throughput, while the ball mills might constrain throughput when milling soft or fine ores. Because the power draw of a ball mill is fairly constant compared to that of a SAG/AG mill, periods of SAG/AG mill bottlenecks lead to increased ball mill circuit specific energy consumption and a finer grinding circuit product size, which together constitute a loss of efficiency of the overall process. Because HPGR circuit throughput is less sensitive to changes in ore hardness, this situation is less common or severe in those circuits. Furthermore, the ball mills are somewhat decoupled from the upstream crushing circuits through the use of surge bins, which also helps to mitigate this effect. As a result, ore hardness variability constitutes a significant cost penalty that should be applied to SAG mill/ball mill/pebble crusher SABC circuits when evaluating the relative economics of different grinding circuit configurations. Nevertheless, the issue is by-and-large ignored, because most current HPGR trade-off studies are based on averaged ore hardness parameters. This paper provides an estimate of the magnitude of this penalty for a typical operation by using Monte Carlo simulation methods and plant data from Freeport-McMoRan’s autogenous, semi-autogenous, and HPGR-based comminution circuits. A case study from an in-house 96K t/d sulphide project is presented to illustrate the concept. It was found that by including variability effects, the trade-off NPV of the HPGR circuit over that of the SAG mill circuit increases to $84 million from negative $33 million, effectively reversing the original decision of selecting a SAG mill circuit for the fixed hardness case.

Jan 1, 2012

By H. M. Howard

"The subject of Grinding Media is perhaps still quite controversial. Many people will state today that the spherical ball is the only way, and also will state that this was settled long ago others are not so sure, and I am in this second category. Over the years there have been many changes in grinding and in our evaluation of grinding media we should always be ready to review the overall situation at regular intervals. There is quite a lot of inertia in all branches of Hineral Dressing and this is particularly true of grinding or, more correctly, comminution.In order to understand some of the things which are done today, and to understand some of the thinking on grinding, it is desirable to go back some sixty years and trace the developments in comminution. In doing this we should consider perhaps three general categories with respect to daily tonnage. An arbitrary division would be under 500 T.P.D., 500-5000 T.P.D., and large operations which might have several parallel units, each with a capacity of about 5000 T.P.D. For the future, in the broad picture, this last category will be the important one because we will be more and more dependent on large-tonnage low-grade operations for most of the metals of the future. Some such division is necessary because the daily tonnage will have a very great effect on what we can and cannot do. It is also necessary to consider the entire flowsheet, and to remember in particular that grinding is not merely to make small particles but is also a conditioning process, to prepare the materi.al for the process which follows. The relationship of crushing to grinding must also be considered. It is logical to include the liners as part of the grinding media and we should try to distinguish between the wear of grinding media and the consumption of media by corrosion. Finally, the whole question of grinding media should be examined in relation to the possibilities of using some form of autogenous grinding."

Jan 1, 1970

By R. G. L. McCready, Marco Zentilli

"Brogan coal containing mean averages of 6,93% w/w total sulphur (1.1% sulphate sulphur, 1.6% organic sulphur and 4.24% pyritic sulphur) and 16.34% mineral ash was used as the test coal in these studies.Shake-culture leaching studies on seven size fractions of the coal, ranging in size from 0.045 mm to 1.7 mm, showed that the rate of pyrite leaching was inversely proportional to the particle size. As the use of continuous stirred tank reactors would be uneconomical for the reclamation of waste coal, column leaching studies were initiated on coal sieved to >6.35 mm to < 12.7 mm, The leachate was monitored for all metal ions and the release of iron was used to monitor the rate of pyrite oxidation . Measurable amounts of Zn, Co, Ni, Mn, and V, all of which can replace iron in the pyrite crystal structure, were released during the initial stages of leaching. Further, high concentrations of AI, Ca and Sr and trace amounts of Sn, As. Se, B, Ca, Cr, Si, and Ti were detected in the column effluents.These studies suggest that bacterial leaching will remove the pyrite and qualitative amounts of environmentally undesirable trace metals from the coal if such a process is developed for waste coal reclamation.IntroductionIn 1947, Colmer and Hinkle showed that acid mine drainage was the result of an oxidative attack of the iron-mineral sulphides by Thiobacillus ferrooxidans to produce sulphuric acid . Ashmead (1955) proposed the use of these organisms to desulphurize coal , but the subsequent research effort was concentrated on the metabolism and physiology of these extreme acidophiles.A limited number of studies have been reported on the bacterial leaching of pyrite from coal (Silverman et al., 1963; Dugan, 1977). However, the bacterial oxidation of pyrite is a self-limiting first-order reaction, probably due to the precipitation of Fe(OH)J and elemental sulphur , which progressively inhibits intimate contact between the bacterial cells and the sulphide-mineral surface. Consequently, the best pyrite oxidation experiments with coal reported to date required five days to remove 97070 of the pyrite (Dugan &Apel , 1978) using mixed cultures of T. ferrooxidans and T. thiooxidans."

Jan 1, 1985

By B. R. McEwen, A. N. Samis, J. H. Gray

"The Quebrada Blanca Copper Mine is an open pit copper project located at 4200 m elevation in the Chilean Andes. The deposit is a world-class supergene enriched porphyry with a mining reserve of 89 Mt of 1.3% copper. The supergene zone will produce 6.3 Mt of ore per year for more than 15 years to produce 75 000 t of cathode copper per year.This paper will review the project geology the computer model ling of the deposit, the mine planning process and the property production scheduling. With copper production .fixed by the onsite electrowinning capacity, the planning and scheduling aspects to optimize the project NPV will be discussed. Some of the logistics, equipment. personnel and process difficulties of operating a mine at 42()() m elevation will also be described. IntroductionProject DescriptionThe Quebrada Blanca Copper Mine is an open pit copper project located in the northern part of the Chilean Andes. The project is based on the supergene enriched portion of a porphyry deposit which has a mining reserve of 89 million tonnes at 1.3070 copper. Below the enrichment blanket, the protore represents a further resource.All ore will be leached, the higher grade material will be crushed and stacked for thin layer heap leaching, and the lower grade and oxide material will be dump leached as run-of-mine material. Solvent extraction and electrowinning will be used to produce 75 000 tonnes of cathode copper per year. Power will be supplied by a 40 MW on-site diesel generating plant with the waste heat being used for process heating. The leaching operation will operate in excess of 15 years and may be followed by open pit mining and conventional ore concentration of the underlying primary sulphide mineralization.Because of the dry location water will be piped 37 km from a ground water source at the Salar de Michincha. Water requirements will be approximately 1.5 million m3/year. The operation will eventually employ approximately 550 people on a 7-day-on, 7-day-off rotation with two twelve-hour shifts per day with an on-site camp. Busing will be provided to the port city of Iquique."

Jan 1, 1994

By M. M. Hajdasinski

"Any study in the field of theoretical investment analysis, or practical project evaluation, must be based on the assumption of one of the commonly applied models of time and value representation, and the corresponding interest convention. The lack of clear assumptions in this regard may easily confuse both the author of the study and the reader, and sometimes even lead to inaccurate results.In the first part of this paper formal definitions of the continuous and discrete models of time and value representation are introduced; in the second part, the importance of transparent assumptions is shown using a concrete example of inflation calculation in project evaluation.IntroductionIn his recent paper, L.D. Smith(10) discusses an important aspect of mine investment analysis: inflation in project evaluation. His interesting presentation of current practices of handling inflation by Canadian and international mining companies in their feasibility studies ends, in the conclusions, with a call for clear assumptions made when evaluating mining projects.With this call, an important issue has been addressed. The lack of clearly stated assumptions, or their incompleteness, are relatively frequent with regard to the interest convention or the model of time and value representation assumed when evaluating mining projects. It is, for example, fairly common that the quantitative evaluation of a project is performed according to the discrete model of time and value representation and the discrete interest convention, and the graphical presentation, supporting this evaluation, reflects the continuous model.Sometimes such an ambivalency means nothing more than just a formal inconsistency; there are cases , however, that may well lead to a serious confusion and /or to inaccurate results . Before this can be demonstrated on a specific example, a concise description of the models of time and value representation used in the evaluation of industrial projects seems to be in order. Although these models (especially the discrete one) have been in common use in the research and practice of financial analysis for considerable period of time, they have been applied rather intuitively, and their forma] definitions are difficult to find even in the standard textbooks on Engineering Economy."

Jan 1, 1990

By L. Avramovic, B. Han, R. Jonovic, Z. Stevanovic, K. Haga, S. Dimitrijevic, A. Shibayama, L. Godirilwe, Y. Takasaki

Pyrite (FeS2) and chalcopyrite (CuFeS2) which could not be recovered, are present in mine tailings discharged from Bor copper mine. About 0.3% of copper (Cu) is contained in these tailings, therefore it can be considered as a useful copper resource. At the same time, pyrite is known to promote acid mine drainage, therefore it is necessary to detoxify the tailings. An efficient method has been proposed to recover copper and reduce the environmental impacts of tailings by a combined hydrometallurgical process of high pressure leaching (HPL), solvent extraction (SX) and electrowinning (EW) process. HPL process yielded high copper leaching rate (>95%) when water (H2O) was used as a leaching media. On the other hand, pyrite contained in the tailings was converted to iron oxide by high pressure leaching. Stability evaluation of the residue obtained from high pressure leaching, confirmed almost no elution of metal ions. The percentage of copper extracted and stripped from the Cu-loaded organic phase reached 97.4%. Moreover, it was confirmed that a low iron concentration in the electrolyte of copper electrowinning, resulted in the improvement of current efficiency of copper electrodeposition. INTRODUCTION Mine tailings are a by-product obtained from the flotation process and there are billions of tons of already existing tailings all over the world. Moreover, several billion tons of additional copper tailings will inevitably be produced in the coming years (Onuaguluchi & Eren, 2012). With the declining average copper ore grade, alternative sources of copper are highly sought, mine tailings have a comparable copper grade to that of low-grade ores (0.2–0.3%) therefore, it can be said that copper mine tailings present a very important source of secondary raw material for copper production (Antonijevi´c et. al., 2008). Consequently, it is necessary to confirm the possibility of the valuable metals (such as copper) recovery from mine tailing in which copper content is about 0.2–0.3%. Our group investigates the flotation and high pressure leaching processes for recovering of copper from mine tailings. Previous work on related samples include copper and nickel flotation tailings from Botswana, and a low–grade copper ore sample from Chile. Impurities, which are associated metals in mine tailing such as iron (Fe) and aluminum (Al) co-dissolve into the solution during the HPL treatment process. Therefore, separation of copper from impurities is necessary for selective extraction of copper from mine tailings and upgrading the copper concentration in the leachate to enable further purifying of Cu by electrowinning. Thus, it is important to investigate the possibility of applying solvent extraction for concentration of copper from leachate solutions of mine tailings (Onuaguluchi & Eren, 2012; Antonijevi´c et. al., 2008). In this study, LIX-84I was used as an extractant of copper from the leachate of high pressure leaching of mine tailing. As a fundamental solvent extraction investigation, the extraction behavior of copper was studied using a simulated solution of the leachate from the high-pressure oxidative leaching of mine tailing. The effects of; solution pH, extraction time, extractant concentration and phase ratio on copper extraction and sulfuric acid (H2SO4) concentration on stripping were investigated.

Jan 1, 2019

By J. M. Toye

The Canadian nuclear fuel business relates primarily to supplying fuel for Candu reactors. The product development, manufacturing and marketing activities of our nuclear fuel industry will therefore be determined by the rate of growth of the Candu nuclear reactor business. The nature of this relationship between the Canadian nuclear fuel industry and the Canadian Heavy Water Reactor program will be considered, with particular reference to the factors influencing future economic growth.

Jan 1, 1969

By J. Vanderlaan

A proposed method of applying high-speed, non-destructive part scanning technologies to the production process of UO2 (Uranium Dioxide) fuel pellets is presented. The scanning process examines each pellet individually for correct surface roughness and any surface imperfection or defects. Using optical scanning methods, which consist of laser and CCD camera technology, the complete circumferential surface of the pellet is examined to meet the reactor fuel bundle specifications.

Aug 1, 2010

By M. C. Bétournay

? Prioritization of this research area identified through a CANMET-MMSL industry survey ? Possible technological breakthrough ? Prioritization of this research area identified through a CANMET-MMSL industry survey ? Possible technological breakthrough Disadvantages ? High cost and long time for equipment development ? Hard rock mining condition limitations ? Size of equipment ? Equipment set-up time, down time ? Performance does not meet financial return targets ? Ground control issues ? Drill and blast retains overall advantage

Feb 1, 2008

By F. Moraw

The use of redox fuel cells in which oxygen is replaced by other oxidants such as ferric ions can have significant advantages [1-3]. For example, the cathodic reaction rate (e.g., exchange current density) can be increased by several orders of magnitude compared to the oxygen reduction reaction, and platinum group metal (PGM) catalysts are not required. As a result of the higher aqueous solubility of the oxidant in a redox fuel cell compared to that of oxygen the rate of mass transfer (e.g., limiting current density) is greater by several orders of magnitude. The redox fuel cell can achieve high efficiencies and has other fuel cell advantages. The main issue with the redox fuel cell cathode is the regeneration of the oxidant where low conversion efficiencies are achieved even with platinum group metal (PGM) catalysts. In the work discussed here the Fe3+/Fe2+ redox fuel cell cathode is characterized over a range of electrolyte concentrations and operating conditions, and electrode materials. Cyclic voltammetry and electrochemical impedance spectroscopy are mainly used for electrochemical characterization and mechanistic determination of the cathode redox couple. Further characterization with selected compositions and materials is done in a redox fuel cell (hydrogen - Fe3+/Fe2+). The mechanism and the advantages and disadvantages of the fuel cell redox cathode are discussed with respect to other approaches and other redox couples.

Jan 1, 2005

By Mohammad Mehdi Salarirad

Fouling of activated carbon by different organic species is a great concern in carbon in pulp and carbon in leach circuits. The rate of gold adsorption onto activated carbon from synthetic gold solutions which contained different amounts of potassium ethyl xanthate (PEX), potassium isobutyl xanthate (PIBX), and methyl isobutyl carbinol (MIBC) were evaluated through batch adsorption tests. The results of these experiments showed that the gold sorption rate rapidly decreases by increasing the concentration of collectors including PIBX and PEX up to 20 mg/L and at higher concentrations, the kinetic rate constant (i.e., k value) gradually leveled off to 53 h-1 for PIBX, which is 46% lower than the 99 h-1 for PEX. At low concentrations of up to 15 mg/L of MIBC, a gradual decrease of the rate constant was observed, while at higher concentrations, the decrease was found to be more pronounced. The gold loading capacity and the rate of gold adsorption onto activated carbon from a gold cyanidation pulp in the presence of flotation reagents were also evaluated. The results of these experiments showed that the flotation reagents have no significant effect on the gold loading capacity of activated carbon at prolong adsorption period, meanwhile the rate of gold adsorption continuously decreases by increasing the concentration of flotation reagents in the leaching solution.

Oct 1, 2011

By A. L. Williamson

Bio-mining, the use of micro-organisms to recover elements of interest, is potentially useful for the economic extraction and recovery of ore metals from a variety of ore deposits, including the lowergrade uranium deposits found in the uraniferous quartz-pebble conglomerate beds of the Quirke Syncline, Elliot Lake, Ontario. The metabolism of chemolithotropic bacterium Acidithiobacillus ferrooxidans depends on an ability to derive energy and reducing power from the oxidation of ferrous iron found in a varirty of iron sulphide minerals associated with base, precious metal and uraniferous ores. The characteristics of this bacterium, especially the ability to oxidize both iron and sulphur with an associated high tolerance of low acidity, allow the organism to contribute significantly to bio-mining processes. Under ideal conditions, A. ferrooxidans promotes the oxidation of ferrous iron in low sulphide ore materials, breaking the crystal structure of the ore minerals .releasing protons to the surrounding solutions, thus promoting the dissolution of uranium and rare earth elements, among others. This paper provides an overview of elemental dissolution and release of uranium, rare earth elements and investigates the acid generating potential of the final residues, from a series of environmentally controlled, biologically-mediated uranium ore extraction experiments. The findings will be used to guide in the design of larger scale bio-leaching experiments to further assess the potential for successful bio-mining operations along the Quirke Syncline.

Aug 1, 2010