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By F. Rivard, M. G. Drouet, W. S. Ruff, P. Carabin

"In both the hot rotary-salt-furnace (RSF) treatment of aluminum dross and the cold ball-millsieving (BMS) treatment of zinc dross, enormous amounts of energy and recoverable metal are wasted and large amounts of greenhouse gas are produced. Such wastes and the environment pollution could be avoided by using the DROSRITE PLUS™ technology. Economic and GHG production comparisons between the presently used industrial treatment processes and DROSRITE PLUS™, supported by over one hundred tests conducted in Canada, the US and Europe, are presented. The savings, with DROSRITE PLUS™, would be about $191 and $418 per MT of aluminum and zinc dross, respectively.INTRODUCTIONDross is a material, which forms on the surface of molten non-ferrous metal, such as aluminum or zinc, during melting, metal holding and handling operations when the molten metal is in contact with a reactive atmosphere. Dross normally consists of metal oxides entraining a considerable quantity of molten free (unreacted) metal, and for economic reasons it is desirable to extract the free metal before discarding the residue. In the case of aluminum dross, recovery of the metal is usually carried out by treating the dross in a furnace at a high temperature (Peterson et al., 2002). By contrast, cold zinc dross is crushed in a ball mill followed by sieving to separate the coarse metal portion from the fine oxide residue.The conventional aluminum dross treatment process, using gas or oil-heated rotary salt furnaces (RSF), is thermally inefficient and environmentally unacceptable because of the salt slag and CO2 produced from the combustion of fossil fuels. In the past several years, a number of salt-free processes have been developed and some of these have found limited commercial use (Gripenberg et al., 1994; Lavoie et al., 1990). In all these dross treatment processes, heating of the cold dross in a rotary furnace requires an external energy input that varies between 375 and 2,500 kWh per (metric) tonne of dross. Molten metal separated during processing of the heated dross is tapped, and the remaining solid residue is discharged from the furnace."

Jan 1, 2012

By Anoush Ebrahimi

Mining dilution is one of the most important factors affecting the economy of mining projects. Ore losses and dilution are present at all stages of mining and while several models can investigate the influence of dilution it is its quantification that poses the most serious challenge (Pakalnis et al., 1995). This paper introduces a methodology to quantify dilution factor in open pit mines. The proposed approach makes use of general mining software available in most mining companies. Therefore it is amenable to most mining operations and provides a readily available tool for understanding mining dilution for operators and consultants. To illustrate the methodology the author uses an example from a hypothetical mine. Dilution increases the operating costs in the mill by increasing the tonnage of material to be milled. In addition to its direct impact on short term income of a mine, dilution causes significant changes in other factors that on the long term reduce the overall value of the project. For example, it prolongs the mine life by reducing mill?s effective capacity. It also reduces the feed grade. In most cases, lower feed grade means lower mill recovery. Dilution also increases the cut-off grade which in turn reduces the overall ore utilization of a mine. In some cases, to take advantage of economies of scale, mining operations tend to plan for higher mining rates. Moving to a larger scale of operation means less selectivity, hence more dilution. This is true for all kind of deposits. High level and uncontrolled dilution may ultimately defeat the purpose of increasing production rates. If not impossible in theory it is nevertheless extremely difficult and expensive to eliminate dilution in practice. Some amount of dilution is practically unavoidable in most mining operations (Scoble and Moss, 1994). Although completely avoiding dilution may be impossible, we can measure and control it. By better understanding the root causes of the dilution and planning accordingly, it can be controlled and reduced. While we do our best to identify and calculate all the other cost items of a project, no matter how small, it is common to make general assumptions about dilution instead of quantifying it. It is common to assume a general dilution such as 5% for massive deposits and 10% for tabular shape deposits. While these figures may be a good starting point in early stages of mining studies, it does not take in to consideration the complexity of the problem. Dilution factor varies within a single mine for different benches and zones. This is due to changes we see in grade distribution and shape of ore body. It also varies based on commodity price. It is necessary to make sure that these variations are taken into account when designing or evaluating a mine. Dilution varies in different mines based on the deposit characteristics, operational aspects and economic cut-off grade. For instance, shape of the deposit, bench height, equipment size, and market conditions will have an impact on the level of dilution in a mine. It is paramount for mining projects to have a better understanding of dilution right from the beginning. In times like today, where mining operations must operate at peak efficiency, it is even more crucial to be able to calculate and have a better grasp of all the parameters influencing the economics of a project including dilution. In order to produce better project evaluations, dilution studies should be an integral part of any project.

Aug 1, 2013

By G. P. van Greunen

"The rehabilitation of underground excavations has been undertaken as long as mining has existed. The conventional methods have proven to be dangerous, time consuming and expensive. During the construction of Styldrift Mine at Royal Bafokeng Platinum several vertical and horizontal excavations had to be rehabilitated. With a zero tolerance attitude towards working in unsafe conditions, together with a non-negotiable timeline to construct and ramp up the mine, the project team could not rehabilitate these excavations using conventional methods. This paper details the approach used to rehabilitate excavations using remote methods and how it was adapted to suit varying operating conditions and requirements - ensuring that no person ever entered an unsafe excavation before it was fully rehabilitated and safe. INTRODUCTION Styldrift Mine is an expansion project of the Royal Bafokeng Platinum group. It is located in South Africa in the North West Province some 10 km south of the Sun City resort (Figure 1). The mine consists of a twin shaft system (Main Shaft and Services Shaft). Capital construction of the shaft and associated infrastructure started in 2009 and is projected to be completed during 2020. The designed steady state production capacity of the shaft is 230 ktpm of Merensky ore. Production ramp-up commenced in 2017 with steady state production set to be achieved 36 months later in the fourth quarter 2020.Merensky Reef was intersected at a depth of 600 m below surface and the shaft design has a mean mining depth of 630 m. The project was approved for execution in 2008 on the basis of a suitable mine design and bankable feasibility study. One of the main excavations that required rehabilitation was an underground reef silo, Silo 4. It is a key piece of infrastructure on the critical path to allow steady state production at Styldrift Mine to be achieved. During construction excessive scaling within the silo along its 64m length necessitated urgent rehabilitation to be able to meet the ramp-up schedule. The criticality and urgency of these excavations prompted the project team to reduce the risk of failure while ensuring safe and cost-effective rehabilitation. This paper will focus on the rehabilitation of Silo 4 while capturing the essence of the learnings of all the excavations that were rehabilitated remotely."

Jan 1, 2019

By M. B. McClenaghan

"The Matheson area of northeastern Ontario is of great interest for mineral exploration because of its geological similarities with the adjacent Timmins, Kirkland Lake and Larder Lake gold camps. Exploration in the area, however, is hampered by the presence of an extensive and thick cover of glacial sediments. For this reason, till geochemistry has become an important exploration method. The Ontario Geological Survey&apos;s Black River-Matheson (BRiM) reconnaissance till sampling project, carried out between 1984 and 1988, was designed to aid exploration in the Matheson area by documenting the Quaternary stratigraphy and providing a database of till geochemistry. Till samples were collected by backhoe trenching of surface pits in areas of thin drift (< 5 m) and by sonic overburden drilling in areas of thicker drift. This paper summarizes the results from the shallow till sampling component of the BRiM project.Two hundred and eighty-two surface samples were collected from Matheson Till, the youngest till sheet in the area, which was deposited during the Late Wisconsinan. For each sample, the non-magnetic heavy mineral (> 3.3 S.G.) and fine (- 250 mesh) till fractions were analyzed geochemically. In addition, visible gold grains were recovered by panning the heavy mineral fraction. Five areas that display multi-element anomalies in both till fractions warrant further investigation: (1) western Egan Township, overlying the Bradley Lake Syenite; (2) around and south of the Munro and Croesus mines, in Munro and Guibord townships; (3) central Garrison Township overlying the Garrison Stock; (4) south of Ore Car Lake in northern Thackeray Township; and (5) southeastern Harker Township. In these areas, Matheson Till has a local provenance and the bedrock sources for these anomalies are nearby in the up-ice (north northwest) direction. Due to the fine-grained nature of gold in bedrock deposits in the BRiM area, the small size of most gold grains recovered from till samples (< 50 um) and the variable distribution of gold in the heavy mineral concentrate and fine fractions, the fine fraction in addition to the more commonly used heavy mineral concentrate should be used for gold exploration.Surface till samples have been affected by weathering to a depth of = 3 m. Weathered till is characterized by a very low carbonate content in the matrix fraction, the absence of sulfide grains in the heavy mineral concentrate and higher concentrations of copper, nickel and zinc in the fine fraction than in the heavy mineral concentrate. As a result of the weathered/ oxidized nature of the surface till samples, the fine fraction is a better sampling medium for base metal exploration."

Jan 1, 1992

By X. T. Zeng, X. Z. Ding, W. R. Chen, L. R. Zhao

The effect of deposition defects in metal nitride coatings on erosion behavior is presented. AlCrN, AlTiSiN, TiSiN and TiSiN:MoS2 coatings were fabricated by reactive cathodic arc evaporation and unbalanced magnetron sputtering processes. The microstructure and mechanical properties of the coatings were characterized by X-ray diffraction, scanning electron microscopy and nanoindentation, and the erosion property was assessed by solid particle impingement testing. The erosion characteristics on the coating surface and cross section were analyzed and correlated to the defects, which played an important role by promoting formation of erosion pits.

Jan 1, 2006

By James Ball

The paper describes a new type of pump and related apparatus specifically designed to dredge sea-floor minerals from very great depths. The principle of operation of the dredging system is explained, and the factors which determine the equipment size and the power required for varying mining rates and depths of water are discussed. An estimate of the capital and operating cost of a commercial-scale mining system is presented.

Jan 1, 1968

By C. Peña, C. A. Zaror, M. Bustamante, M. Vega

The goal of this paper is to present the carbon footprint of Chilean cathodic copper, as a function of the electricity generation mix, ore grade, and process efficiency. Ore grades and acid leach-solvent extraction-electrowinning (LX-SX-EW) yields in the range 0.5-1.0% and 60-80%, respectively, were considered here. Additionally, electricity supply from two Chilean suppliers, SING (1.0 ton CO2/MWh) and SIC (0.4 ton CO2/MWh) was assessed. Simulation results show that oxide plants located in northern Chile, served by the SING network, feature carbon footprint within 4.3 - 11.4 MWh I ton cathodic copper, as compared with 3.5 - 9.3 MWh I ton copper in plants served by SIC. The footprint was found to increase as ore wade and process yields decrease.

Jan 1, 2011

By V. ?. Chanturiya, I. Zh. Bunin, M. V. Ryazantseva

"Surface changes in naturally occurring metal sulfides (pyrite, arsenopyrite, sphalerite, chalcopyrite, galena, and molybdenite) due to treatments from high-power electromagnetic pulses (HPEMP) at varying times were studied using X-ray photoelectron spectroscopy. Analysis of the obtained results revealed common patterns and differences in surface transformations. The transformations were found to include two main stages. The first stage was observed at low treatment intensities (up to n ~103 pulses). At this stage, formation and accumulation in the surface layer of the nonstoichiometric sulfide phase, oxides and hydroxides, as well as elemental (polysulfide) sulfur and/or metastable sulfur species (thiosulfate, sulfite) was observed. The second stage (N = 3·103 pulses) is characterized by thermal removal of sulfur species and renewal of the mineral surface (sulfidization).It was found that the chemical changes of the sulfur on the surface layer of pyrite, arsenopyrite, chalcopyrite included accumulation/formation of S0 (Sn 2-) at the first stage of the transformation, followed by its removal. An increase in the surface concentration of elemental sulfur for pyrite and chalcopyrite was 10–12% while arsenopyrite was approximately 5%. A more complicated chemical change was observed for sphalerite and galena, that included formation/accumulation of metastable sulfur species (thiosulfate, sulfate) at the initial stage (up to 103 pulses) and during longer treatment times, the sulphur is reduced back to its initial state (sulfide or disulfide). The increased concentration of S0 sulfur, compared to the original sample, was only 3% in galena and sphalerite.The application of HPEMP treatment to improve flotation selectivity is supported by singlemineral flotation tests. Changes in flotability as a result of HPEMP treatment are principally explained by surface phase changes that resulted in higher adsorption activity."

Jan 1, 2016

By M. Pearson, J. Irvine, L. Mercer, P. Lauzon

Severe service ball valve performance has a direct impact on the mechanical availability of high pressure oxidation and leach plants. The primary function of these valves is to isolate the autoclave for maintenance. Valve failure has historically contributed significantly to the maintenance downtime of these process plants. An understanding of the root causes of valve failures is the first step in increasing the reliability of the severe service ball valves. A rigorous failure analysis practice has been developed by Hatch for investigating refractory failures in furnaces and kilns based on Kepner-Tregoe methodology. The method has also been modified for valve failure analysis. From a root-cause failure tree analysis of each of the valve components, a list of all potential valve failure mechanisms was developed for ranking and evaluation, along with recommendations for further valve development. The purpose of this paper is to present the reader with the modified Kepner-Tregoe method and examples of this analysis.

Jan 1, 2004

By M. Pearson, L. Zunti

Pressure vessels in hydrometallurgy and many chemical process industries throughout the world require the use of specific lining systems to protect the parent vessel materials from corrosion and/or abrasion. Typically, these lining systems are required for processes that are corrosive and operate at elevated temperatures. Two basic lining system types arc: 1) refractory linings in conjunction with an impervious corrosion resistant membrane, and 2) metal-clad vessels consisting of titanium or other corrosion resistant metal bonded to a carbon steel shell. Operating conditions play a key role in selecting the type of lining system to be utilized for a specific application. Lining systems in turn have an impact on the process vessel size, which ultimately affects the capital cost of the vessel. The capital costs for each type of lining system must be considered for each application, as well as maintenance concerns, quality control, and delivery. It is clear that operating conditions play a key role in the final selection of lining systems utilized for specific projects based on initial capital cost, life-cycle maintenance costs, and overall service life. This paper will examine the fundamentals of each lining system, their inherent technical strengths and weaknesses, and present relative cost comparisons for each lining system as they apply to specific operating conditions and vessel sizes.

Jan 1, 2004

By Troy Boisjoli, Sean Hillacre, Matt Batty

Discovered in 2014, the Arrow Uranium Deposit has undergone rapid development through successive drilling programs informed by geostatistical drill hole spacing studies. An independent third-party geostatistical drill hole spacing study was completed by Clayton V. Deutsch Consultants Inc. in 2017, evaluating the uncertainty of thickness, grade, density, and areal limit of the Arrow Deposit to determine a drill hole spacing suitable for Indicated Mineral Resources. Indicated spacing was defined to be 24 to 32 metres (“m”) to be within 15% of predicted for a nominal quarterly production volume with a probabilistic tolerance between 75 and 90%. The study not only defined the Indicated Mineral Resource resultant from the 2016 drill program, but also provided drill hole targeting criteria for subsequent drill programs aiming to convert Inferred to Indicated Mineral Resources. A subsequent geostatistical drill hole spacing study on the Arrow Deposit was completed by Clayton V. Deutsch with Resource Modeling Solutions in December 2018 to define Measured Mineral Resource spacing in support of the 2019 drill campaign. The study resulted in a drill hole spacing of 9 to 17 m to be within 15% of predicted for monthly and quarterly production volumes using an accepted industry probabilistic tolerance of 90%.

Jan 1, 2020

By Laura Steeves, Inoka Eranda Perera, Jhon Silva, Arend Spaeth

"This paper reviews the main concepts ruling the design, implementation and performance of active explosion barrier systems for underground coal mines in the USA. The document examines the parameters needed for the design of this type of system, such as overpressure and flame speed, suppressant deployment speed and duration time of the active barrier. The document also discusses the available solutions for gassy mines and the concepts behind the selection of machined mounted and roadway active explosion barrier systems. The paper includes a review of active explosion barrier systems developed in the USA and South Africa. The paper provides an overview of the current status of this technology in the underground coal mining industry in the USA, and the possible implementation of such systems. The paper concludes with the identification of the future work needed for the successful implementation of the systems in several conditions to enhance the safety of underground coal mine operations and to provide more tools to the miners for their safety.INTRODUCTIONUnderground coal mine explosions can be addressed in two ways, through prevention or mitigation. The United States mining industry practices have emphasized the prevention of explosions in underground coal mines, with little being done to mitigate the explosion once it occurs. Despite successful practices for the prevention of underground coal mine explosions in the US, the threat of an explosion does still exist. Other mining countries around the world have developed and implemented mitigation systems such as active explosion barriers in their underground mines. In the 1970s, the United States Bureau of Mines (USBM) carried out extensive research on the use of active or triggered explosion barrier systems in the US; however, they are not currently used.Explosion barrier systems do not prevent explosions but mitigate the propagation of the shockwave and flame to other areas of the mine. The use of active explosion barrier systems enhance the safety of underground coal mines. Active or triggered explosion barriers will detect the arrival of the shockwave or flame front of the explosion and release inert material to suppress or extinguish the flame. The technology required for active barrier systems is readily available based on their presence in other mining countries and must be adapted for their use in US coal mines. The adaption must consider the coal dust and methane concentrations in US coal mines, the mine geometry, and the mining equipment used."

Jan 1, 2019

By Eugene Sun, Qiang Xu

This presentation will review current applications of composite materials in aircraft design and discuss the challenges involved in maximizing the potential of the composites in aircraft applications.

Jan 1, 2010

By M. Medraj, M. A. Arafin, D. P. Turner, P. Bocher

In this study, the kinetics of isothermal solidification during TLP bonding of Inconel 625 and 718 superalloys with nickel based multi-component filler alloy BNi-2 have been studied through a combination of direct experimentation and mathematical modeling in order to determine the effect of process variables and, thus, to optimize the process. The predicted isothermal solidification completion times from the boron diffusion model and migrating solid/liquid interface model were in good agreement with the experimentally determined values and, they were found to be significantly less than those of other nickel superalloys with different nickel based filler alloys. Further, significant reduction of holding time was observed with increasing bonding temperature and with decreasing joint gap. No significant grain growth in the base metals has been observed in the temperature range being investigated (1325 - 1394K).

Jan 1, 2006

By P. Karimi, A. Amini, H. Abdollahi, A. Mirzaeian

Due to detrimental effects of cyanidation process on the environment in gold industry, application of cyanide alternatives such as thiourea are being highly considered in recent projects. In this study, the extraction of gold from poly metal ore, using thiourea was precisely investigated. The representative sample contained 2.5 ppm Au and 160 ppm Ag. The sample was categorized as oxide gold ore, because the main part of its pyrite was changed to iron oxide and hydroxide such as goethite and lepidocrosite. According to mineralogy, the main gangue mineral was quartz. It shows that gold grains are free and interlocked with quartz matrix. Also, the presence of a high amount of silver in free and sulfide forms and also the low content of other sulfide minerals such as base metal sulfides is a remarkable point of these studies. In the gold extraction process with thiourea, there are several parameters that affect the rate of gold dissolution such as concentrations of thiourea as a gold lixiviant and ferric sulfate as an oxidant, pH and size fraction of feed and leaching time. The influence of these parameters on the gold dissolution rate were investigated and optimized. The results showed that the concentration of thiourea is the most influential parameter which could increase the gold recovery to 71% (at 10 g/kg sample). The optimal amount of ferric sulfate was determined as 6 g/kg sample. Also, the optimal levels of size fraction of feed and pH of solution were 88% under 75 micron and 2 as acidity; respectively. Leaching time was 6 hrs. In addition, this sample was leached with cyanide and the optimized values of size fraction of d88=75 microns, pH=10.45, 2500 ppm of sodium cyanide with time leaching of 24 hours were concluded for maximum recovery of gold in solution. Finally, the results of cyanidation tests were compared to the thiourea dissolution after optimizing all affected parameters in cyanidation process.

Jan 1, 2016

By M. Li, T. H. Dubaniewicz, H. N. Dougherty

"INTRODUCTION Underground mines have the risk of accumulation of gases if not ventilated properly or during and after emergency situations when ventilation controls become damaged or stop working. Common gases that are monitored in underground coal are carbon monoxide (CO), oxygen (O2), and methane (CH4). Methane is a gas found in coal and released through mining; its explosive range is between 5% and 15%. Methane is commonly monitored on all mining equipment and with handheld gas sensors carried by most underground coal miners. Fiber optics (FO) is a popular technology applied in underground mining, particularly for its electric isolation. Many miners are familiar with FO cables hung throughout the mine, which are primarily used in the communication of voice and data (Taylor et al. 2010). Advancements in laser and fiber optics have made this technology well suited for gas sensing the harsh environment of mining applications. The work of the U.S. Bureau of Mines on the use of fiber optic monitoring of gases in mines started in 1992. Dubaniewicz et al. (1992, 1993, 1995, 1999, 2003) developed prototype gas sensors, including methane gas sensors, for research purposes. Previous research (Dubaniewicz & Chilton 1992) identified potential fiber optic gas detection cross-sensitivities associated with the index of refraction of various gases. Index of refraction effects was noted for carbon dioxide (CO2), which is produced in significant quantities by mine fires or explosions. NIOSH released a broad agency announcement (BAA) for the development and demonstration of mine safety and health technology (CDC NIOSH 2017). The objective of this announcement was to validate a sensor, specifically fiber optic, for the mine environment and test it in an actual working mine. The RSL FO sensor “OptoSniff” provided from this work is capable of hundreds of sensing points over distances of up to 20 km, has a measurement range of 0.05% to 100%, senses in an oxygen- deficient environment, and has no gas cross-sensitivity (RSL, 2014). The RSL system is totally electrically passive outside the central control unit (CCU), which can be housed in an MSHA-approved explosion-proof (XP) enclosure or in a control room outside of the permissive area. The sensor was tested in dusty and signal loss situations, showing it is tolerant of dust, has functioned at signal losses up to 90%, and was successfully tested in a mine environment for a five-month period (Walsh et al. 2014). This paper describes how the RSL FO sensor is used for testing using smoke and its contaminants for applicability in emergency situations. The testing investigates individual contaminants such as CO2, temperature, and smoke. Each test was performed separately and methane concentration, laser power, and response time data were collected."

Jan 1, 2019

"INTRODUCTION Secondary mine ventilation has been hamstrung by the availability of fans that satisfy many competing operational factors. The sometimes legislated requirement of adequate volume to the working face, the minimisation of excavation required to fit the selected secondary fan and the power requirements of various operational phases has, in the past, not been satisfied with just one system. The new system being introduced overcomes all these competing requirements in one simple unit with an added bonus of being very durable without loss of performance when wear occurs on the impeller. Existing Operational Situation Several studies (e.g., CIPEC, 2005; LCG Energy Management Group, 2009) have found that between 42% and 49% of mining energy costs are associated with mine ventilation, depending on the type of mine and the equipment used. If we then look at the general breakdown of installed power of the secondary fan to primary vent fans, the ratio is around 2:1. Thus the total energy costs of the secondary ventilation represents about 25% of the overall mining energy costs. It is of interest to note that the cost of running a typical two stage axial vs the cost of purchasing and general maintenance cost of the same fan over a 10 year period, the breakdown will be conservatively in the order of 40:1.The currently available secondary ventilation axial fan have one duty curve at any given time. Generally this can be changed manually within the motor power parameters but requires a minor outage and removal of the fan to do the pitch adjustments. As the bag vent ducts grow longer, the supplied air reduces to a point where inadequate air is supplied to the working face. This is exacerbated by underground equipment putting additional resistance in the heading and further reducing the supplied air at the working face. This shortfall is normally corrected by the installation of an additional fan in the system to overcome losses in the bag system but comes at a cost of production as the heading needs to be shutdown to install the second fan. This also means that the heading is supplied with excessive air on the shorter runs with a wastage of fan power."

Jan 1, 2019

By Richard K. Komarek, Antonios Zavaliangos, Roman T. Dec

Current industrial compacting and briquetting practices are largely based on trial-and¬error techniques. This paper introduces a two dimensional finite element model of compacting/briquetting process in the roller press. Compared with other modeling techniques used in the past, finite element method offers the most versatile approach because it incorporates adequate information about powder behavior, geometry and frictional conditions and offers predictions for the density distribution, velocity field, stress distribution, roll force, roll torque and other important process variables. More over, when the model includes elasticity, it offers predictions on the internal stress state in the briquet after its exit from the pressing zone. This enables to perform realistic computer experiments which are helpful for establishing the pressing system design, forming cavity geometry and process conditions necessary to produce compacted or briquetted product with the desired properties.

Jan 1, 2004

By J. Warner, K. A. Hughes, P. E. R. Blanchard, J. Esslifie-Dughan, J. W. Reid

Tailings management plays an important role in reducing the environmental impact of Orano Canada’s uranium mining operations in Northern Saskatchewan. Uranium ores mined from Northern Saskatchewan are often associated with large amounts (10-20 wt%) of other harmful elements, such as arsenic and nickel. The JEB Tailings Management Facility (TMF) located at McClean Lake, Saskatchewan is designed to prevent potentially harmful water-soluble elements, particularly arsenic, from contaminating the surrounding environment by promoting the formation of water-insoluble minerals. It is therefore important to determine which arsenic-containing mineral species are present in the TMF and how they evolve over time. Researchers at the Canadian Light Source (Saskatoon, Saskatchewan) were successful in identifying the arsenic mineral species present in tailings samples collected from the TMF in 2013 using X-ray absorption near-edge structure (XANES) spectroscopy. XANES analysis verified that the majority of arsenic species in the TMF existed as amorphous and crystalline ferric arsenates with minor amounts of arsenite and arsenide species. Herein, XANES is again used to analyze a new set of tailings samples collected from the TMF in 2018, specifically focusing on the difference in the distribution of arsenic species in old and new tailings.

Jan 1, 2020

By B. Holaday, H. McLaughlin

Orano Canada Inc. (Orano) operates the McClean Lake mill, located in the Athabasca Basin region of northern Saskatchewan. It processes high concentration uranium ore, delivered in slurry form from the Cameco Corporation operated Cigar Lake mine. After the ore slurry is leached at the McClean Lake mill, the leach residue solids are washed in a process known as counter-current decantation (CCD). This process uses six thickeners to facilitate counter-current washing of solids with diluted acid, flocculation and settling, as well as pumping the washed leach residue to the tailings neutralization circuit (TN), and the uranium laden aqueous to the solvent extraction circuit (SX). Challenges in settling the slurry in CCD started in early 2017. This is largely due to the variation in fines content of the slurry (D80≤75μm), and clay content. There are several changes that have allowed the mill to adapt to the evolving challenges of settling leached Cigar Lake leach residue. Flocculant selection and mixing, temperature effects, bed level management philosophy, and centre-well feed density are key parameters that have been explored as part of adapting to changes in Cigar Lake ore quality.

Jan 1, 2020