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By D J. McDougall

Due to changes in the solid-state conditions of minerals in the rocks surrounding hydrothermal, contact meta-somatic or metamorphosed ore deposits, positive or negative thermoluminescent anomalies may be found in their vicinity. Several types of equipment which can be used for both quantitative and semi-quantitative measurements of the emission are discussed and compared. Methods of selecting, preparing and testing samples to ensure reproducible results are outlined, along with the kind of data which can be obtained from different types of equipment. The normal thermoluminescence of rock-forming min~ erals in the common rock types is noted, and variations in the solid-state conditions in the wall rocks which will cause marked increases or decreases of true thermoluminescence are outlined, as well as the thermolum-nescence-like emissions which are present in the vicinity of ore deposits. Introduction T HERMOLUMINESCENCE occurs when certain crystalline materials are heated to temperatures ranging between 180' and 400°C. The light emission can be ascribed to the release of electrons metastably trapped within the crystal lattice, with peak emissions taking place at temperatures which are more or less specific for a particular set of traps; the area under

Jan 1, 1965

By M. V. Rowley, D. D. Warkentin

With high commodity prices, tightening operating budgets and increasing disposal costs, there is an increasing interest in processing municipal waste streams, such as the ash derived from combustion processes. This paper describes laboratory development work for a process to recover metal values from the finer fractions of these wastes. The work was initially aimed at recovery of copper and the removal of lead and other heavy metals from the ash residue, but results showed gold and silver recovery to be an important objective as well. The process uses gravity and flotation as the principal separation steps, with gold recovery occurring mainly in the flotation stage. The particle size range and chemistry of the ash presented numerous challenges to developing a suitable flotation method, but an effective flowsheet was developed by separating the ash into a fine fraction and a separate slime fraction, each with a unique flotation strategy. Specialized conditioning stages and careful reagent selection have resulted in effective gold and silver recovery from a range of ash types and sources. The work has resulted in a patent being issued for a process based on this novel approach to the treatment of combustion wastes.

Jan 1, 2017

By A. Karzhau, B. Aldamzharov, Y. Tussupbekov

"This case study produced for the metal mine operating in Kazakhstan, which has ventilation controlled manually and working continuously at maximum speed to keep air flowing through each part of the mine 24 hours a day. The mine auxiliary ventilation has been designed with the maximum length of ducting systems with extra air volumes to compensate for potential leakage. The mine provided auxiliary ventilation to all production stopes even to those, which for various operating reasons become temporarily inactive. However, mine safety has to consider risks associated with the deficit of fresh air for mine gases dilution and removal. The mine needs additional investments for re-designing ventilation network and therefore needs to estimate financial risks associated with it. The aim of this study is to conduct risk analysis for ventilation network optimisation processes at the operational metal mine in Kazakhstan. Mine ventilation surveys were undertaken for validation of the computerised ventilation model. The mine ventilation network modelling has been undertaken to be compliant with Kazakhstan mining regulations. Results of the mine ventilation optimisation processes were analysed in terms of financial risks for investments into reconstruction of the ventilation shafts, extension of the ventilation network, variable speed fans and automated regulators. Savings were resulted from improving ventilation system efficiency by reduce airflow in areas without mining activity and redistributing existing ventilation capacity. Based on savings derived from the optimised ventilation network an additional NPV has been received. Stochastic modelling use Monte Carlo simulation producing representation of possible outcomes for the additional NPV considering electricity pricing as a major variability. As result probability distribution for the optimised ventilation network demonstrates the additional NPV in a range of 1.6 - 2.2 million US$ in the corresponding defined value of 90%. Moreover, there is only 5% chance that the additional NPV to be less than 1.5 million US$. Conducted risk analysis for ventilation network optimisation processes gives considerations on volume of investments into the mine ventilation system.INTRODUCTIONCosts for mine ventilation are typically increasing when mines expand to deep levels. Some underground mines in Kazakhstan still use continuously operating fans at its maximum capacity. This is not helping to save energy and costs that are in many cases very significant. This case study produced for the metal mine operating in Kazakhstan, which has ventilation controlled manually and working continuously at maximum speed to keep air flowing through each part of a mine 24 hours a day. The mine auxiliary ventilation has been designed with maximum length of ducting systems and extra air volumes to compensate for potential leakage. The mine provided auxiliary ventilation to all production stopes even to those, which for various operating reasons become temporarily inactive. However, mine safety has to consider risks associated with the deficit of fresh air for mine gases dilution and removal from the mine workings (Sabanov, 2019). The mine needs additional investments for re-designing ventilation networks and therefore need to estimate financial risks associated with it. The aim of this study is to conduct risk analysis for ventilation networks optimisation processes at the operational mine in Kazakhstan. Mine ventilation surveys for validation of the computerised ventilation model have been undertaken"

Jan 1, 2019

By G. M. Hughes, J. Ellis

"IntroductionHydraulics has been considered a science for thousands of years. Applications included water wheels, dam; and slice gates but today the term ""hydraulics"" commonly refers to ""power hydraulics"" in which fluid is used under controlled pressure to do work. The mining industry employee’s hydraulics in practically every aspect of its operations, on such equipment as trucks, front end loaders, LHD units, crushers and hydrostroke feeders. The many benefits of hydraulic drives such as ease of variable speed control, high power: weight ratio low envelope size of hydraulic actuators and ease of maintenance have been extensively developed in the pulp and paper, chemical, steel, and other basic industries, but until quite recently, concentrator operators have not taken advantage of the versatile methods of power transmission offered by hydraulic systems.After some preliminary testing at the Lynn Lake concentrator Sherritt Cordon Mines Limited installed several hydraulic system; at their Fox Mine concentrator to drive slurry pumps and conveyors. The systems used at Fox and the systems which will be installed at the Ruttan concentrator are briefly described in this paper. Since acceptance is based mainly on understanding a few basic principles of hydraulics system; including pumps, motors, circuitry and central methods are also presented in the paper.PRINCIPLES OF POWER HYDRAULICSBasicsHydraulic systems provide an uncomplicated, versatile means of power transmission. The basic system is quite simple. It consists of five main elements, as shown in Figure 1, which are:1. Prime Mover - generally an electric motor.2. Hydraulic Pump - fixed or variable positive displacement type such as a piston or vane p.mp which displaces a constant amount of fluid for ead1 revolution - this unit converts mechanical energy of the prime mover into fluid energy.3. Oil Reservoir - a standard tank to operate under atmospheric pressure which has a capacity of 2 - 3 times the maximum gpm outp.it of the hydraulic pump."

Jan 1, 1972

By Sunil Koppalkar, A. R. Laplante

"A Pilot Plant was installed in one of the grinding circuits of the Dome Mill, Ontario, Canada. Tests were conducted on a KCCD12 Knelson Concentrator to systematically analyze the effect of operating variables on the performance. The main objective of this test work was to explore the effect of very coarse (6 mm) feed size on GRG recovery and performance of the Knelson concentrator. Preliminary results have shown that coarse feed size is deleterious to maximizing recovery. Size-by-size GRG recoveries have shown an unusual trend in the sense that recoveries do not decrease monotonically with decreasing gold particle size. Similar trend observed in one other full-scale unit has been compared. The observed trend is probably linked to the feed size distribution and top size.INTRODUCTIONKnelson Concentrators are the most widely used semi-continues centrifuge gravity separation units for the recovery of gold and platinum group minerals (Xiao and Laplante 2004). The main advantages over the competing centrifugal gravity units are their design, operating simplicity and good separation efficiency that make them the most common choice for gravity circuit. Despite some characterization at bench scale and a limited characterization at full scale, there is a lack of a comprehensive study of impact of operating variables on efficiency that reflects realistic full scale operating practice. GRG characterization studies, at bench scale, on these units provide information about the occurrence of gold in ore samples but fail to provide adequate information about the effect of operating variables for full-scale units. It is well known fact that such work is difficult to undertake on full-scale units owing to the difficulties involved in sampling and systematic variation of operating parameters. To tackle these problems, a pilot facility consisting of a 12-in CD Knelson Concentrator, a feed screen and tailing pump was installed in one of the grinding circuits of Dome Mill. This paper describes the work carried out and the results.Porcupine Joint Venture (PJV)Porcupine Joint Venture was formed in 2002 between Placer Dome Inc. and Kinross Gold Corporation. The JV utilizes the milling capacity of the Dome Mill to process ore from Placer Dome and Kinross mines. The Dome Mill circuit has gravity concentration, cyanidation and CIP circuits to recover gold. Gravity recovery has been used at Dome Mill since inception to recover gold: first by using conventional jigs, which were eventually replaced by two Knelson Concentrators in 1993, one in each of the grinding circuits. Three more Knelsons joined the circuit in 2002 to improve gravity recovery. An Acacia Intensive Leach Reactor was also added in 2002 to process gold concentrates replacing shaking tables (Chong et al 2004)."

Jan 1, 2007

By P. Babak, E. Insalaco, O. Babak, P. Henriquel

"Detection of outliers is an important primary step in geomodelling. Geological datasets often contain errors due to problems such as measurement device limitations and recording glitches. In this paper, a robust Mahalanobis distance–based approach is successfully applied to particle size distributions of the Joslyn lease, an oil sands development site near Fort McMurray, Alberta, Canada. Automatic outlier detection and geological review were combined. Outliers from the Mahalanobis distance approach were compared with their respective core photographs to determine if laboratory measurements and facies classification were correct. A checking procedure for laboratory technicians is proposed.RÉSUMÉ La détection des valeurs aberrantes constitue une étape primaire importante de la modélisation géologique. Les ensembles de données géologiques contiennent souvent des erreurs en raison de problèmes tels que les limitations des instruments de mesure et des pépins d’enregistrement. Dans le présent article, une méthode robuste basée sur la distance de Mahalanobis a été mise en oeuvre avec succès aux granulométries du bail minier Joslyn, un site de sables bitumineux en développement à proximité de Fort McMurray, Alberta, Canada. La détection automatique des valeurs aberrantes a été combinée à une étude géologique. Les valeurs aberrantes de la méthode de la distance de Manalanobis ont été comparées avec leurs photographies respectives de carottes de forage afin de déterminer si les mesures en laboratoire et la classification des faciès étaient correctes. Une procédure de vérification pour les techniciens de laboratoire est proposée."

Jan 1, 2015

By Corrie Pitzer

Safety performance in many companies and even industries has stalled in the last decade. Accidents rates are improving but more slowly and yet, serious accidents and fatality rates are not: in fact, these rates of more serious incidents have plateaued or deteriorating. In more dramatic cases, organizations that had ?exemplary? safety statistics, suddenly had a catastrophic or multi-fatality event occurring. Popular books such as In Search of Excellence of Peters and Waterman or Good to Great of Jim Collins created many delusional ?secrets? or recipes, to achieve great performance and yet, most of those ?great? or ?excellent? companies have now declined in their performance, some dramatically and some have fallen from grace, like Fanny Mae, a company Jim Collins identified as a ?great? one, that had to be bailed out in the latest financial crisis. These organizations share several ?delusions? about managing a business, and often, these delusions are created by academics, consultants and authors of popular books about (safety) management. More damaging however is that safety management is defined by a set of key practices or axioms: a focus on human error, compliance to rules and regulations, creating consistency of systems and actions, reacting to accidents and incidents and on prevention of those, measuring safety with performance metrics and rewarding employees on those measurements. All of these axioms contain the seeds of destruction, and they create conditions and dynamics that precede the disaster - and ironically, they are all 'good things' we do in safety management. For example, a mere focus on safety in the business, and the measurement of it, create a culture of repression of mistakes and incidents, leading to a situation where the organization knows incrementally less and less of risks and risktaking behaviour - until the day it all coincides destructively! This paper discusses seven such safety delusions and how they impact on the views and safety management approaches. They are: ??The delusion of (risk) control, that risks can actually be identified, analyzed and mitigated. ??The delusion of compliance, that mere compliance to rules will safeguard the business ??The delusion of consistency, that standard operating procedures will ensure safe operations ??The delusion of predictability, that accidents and trends tell us everything about the risks of the future ??The delusion of human error, that people are the weakest link and that it can actually be dealt with through discipline and enforcement. ??The delusion of statistical trends, that safety is measurable with statistics ??The delusion of invulnerability, that we have achieved high levels integrity and capability to deal wit risks. Organizations that suffered catastrophes all share some or all of these delusions and this paper traces how these played a role in many disasters, ranging from the Titanic, Challenger, Piper Alpha, Bhopal, BP Deepwater Horizon and many more.

Aug 1, 2013

In Canada and other countries, legislation dealing with accurate factual disclosure for mining projects is being expanded to include more metallurgical design information. Grinding characteristics, metallurgical recovery of contained metal and environmental hazards associated with tailings disposal will be part of the Canadian package. This paper will examine only that portion of the plant design standard practice that relates to grinding ? in SAG mills and ball mills that grind ore from as received crushed material to final ground product that is fed to downstream plant recovery processes. Since grinding plants represent a significant portion of the overall concentrator cost, accountability in grinding design should no longer be left to chance, opinion, or hidden proprietary information. This is the 21st century. It is reasonable to expect that a factual basis will be needed so that a Qualified Person (QP) reviewing the grinding mill design can understand it. That will also involve how many and what quality of samples are taken to define the characteristics of the deposit, the accuracy of the tests used to measure ore hardness and the method by which the dimensions and power of the mills are determined. If these criteria are not met, the QP may have no choice but to call attention to the lack of convincing information regarding the grinding mill design for the project. At the same time clients are expected to demand getting the correct design information at a cheaper price, results in a more timely way, and accurate design to the point where the risk of tonnage shortfalls are reduced or eliminated. Tables will be presented to show how this can be done to current standards using existing technologies. At the same time it is acknowledged that current standards are ill-defined and that to achieve consensus among those involved in SAG mill design work, it may require that a consortium funded project be set up to establish what tests and methods for grinding design are suitable to meet the new standards for metallurgical disclosure. By running parallel design methods on several major projects over a period of 5 years, the process of how to design grinding circuits will be better understood because the projects will start at the design stage, include procurement of the mills and will end with a benchmark test on the plant. Selecting the samples and the best schedule for doing test work on core samples submitted and mill design, will be discussed in this paper. The issue of completing the grinding design in time for early procurement decisions is on the critical path for most major projects. Discussions between mining engineers, geologists and process engineers are needed early in the development cycle, at the scoping study level (SS), preliminary engineering assessment (PEA), prefeasibility study (PFS), and at the feasibility study stage (FS), to ensure that proper planning for design sample selection is done. Using adequate sampling and testing methods in a timely way allows early procurement of the grinding mills and minimizes project risk.

Jan 1, 2012

By V. Vorontsov

Catalysts Mo-Ni-S were prepared by combining MoS2 and NiS in weight ratios 1:1, 1:2, 1:4 and 2:1, each of which was admixed with 5%wt Ag and 5%wt YSZ to increase three-phase boundary surface area of the electrode. For comparison catalysts comprising Co-Mo-S in weight ratio of 1:1 of CoS to MoS2 admixed with 5%wt Ag and YSZ and pre-sulfided NiMoO4 were made. The catalysts were examined for electrochemical oxidation of hydrogen sulfide in solid oxide fuel cells (SOFC) at 850°C temperature. The catalyst prepared from 1:1 (wt. %) NiS:MoS2 demonstrated highest performance. Current density up to 1,330 mA/cm2 and power density up to 300 mW/cm2 were achieved using pure H25 gas, which are the highest among the reported H2S-powered fuel cells ever published in the literature. Anode materials of different NiS:MoS2 ratios have been characterized by means of X-ray diffraction (XRD), simultaneous TG/DSC analysis and Neutron Activation Analysis (NAA). TG/DSC and NAA results show that addition of NiS and Ag increases fuel cell performance due to improved structure/integrity of the anode. It has also been demonstrated that MoS2 is a suitable anode material in H2S-powered SOFC, with high activity towards H25 oxidation and thermally and chemically stable up to 850°C.

Jan 1, 2005

By Jun Wang, Bo Zhang, Wenqing Qin, Guanzhou Qiu, Xiaozhen Ma, Key Lab of Biohydrometallurgy of Ministry of Education, Fen Jiao, Central South University, Yansheng Zhang

"Arsenic contained in minerals sulphides such as enargite and tennantite can represent not only a significant penalty element in base metals production, but also a disadvantage to the environment through producing the toxic emissions to the atmosphere. The achieving separation of arsenic bearing minerals either at an early stage in floatation or from the floatation concentration is significantly beneficial to the economic. However, there was no feasible commercial method of separation at the flotation stage because of its similar flotation properties with many other copper sulphides. In this work, the selective floatation of enargite from copper sulphide concentrate containing maily covllite, chalcoctie, bornite and enargite was investigated. A low arsenic concentrate with As 0.2%, Cu 22% and a high arsenic bulk concentrate with As 7.65%,Cu 33.65% were obtained. It was indicated that Ca(ClO)2 is the effective depressor and the optimum floatation condition was obtained. Most enargite could be floated at pH 11 and controlling the pulp potential at 290 mV SHE. Furthermore, the arsenic removal from the high arsenic bulk concentrate by hydrometallurgy methods was study. The results showed that the reaction temperature and the concentrations of NaOH were the most important variables affecting the rate of arsenic removal. A low arsenic concentrate with As 0.24%, Cu 37.80% was obtained in the optimum leaching condition.INTRODUCTIONArsenic is mainly produced as a by-product of base metals smelting, in particular copper concentrate. China has imposed an upper limit of 0.50% arsenic for concentrates entering (Long et al., 2012). Arsenic amount of copper concentrate within 0.40% has also been imposed in China, and above which will be introduced about 50 RMB penalties per ton for each 0.10% increasing by smelters now. Therefore, it is necessary to remove arsenic from the copper concentrate.In copper concentrate, arsenic is mainly in the form of arsenopyrite (FeAsS), enargite (Cu3AsS4) or tennantite (Cu12As4S13). There are two types of arsenic in the separation of copper and arsenic: arsenopyrite and arsenic bearing copper mineral. It is easier to separate arsenopyrite from the copper sulphide by flotation in comparsion with the separation of enargite or tennantite from non-arsenic copper minerals (Pineda et al., 2015). The main difficulty in the separation of arsenic copper mineral from non-arsenic copper sulphides is their extremely similar floatability, which depends on the fact that tennantite and enargite contain high copper content (Pineda et al., 2015)."

Jan 1, 2016

By A. Elsayed

Lightweight metals such as magnesium show potential in replacing aluminum alloy castings for engine cradles, cylinder heads and transmission cases. Unfortunately, magnesium alloys are susceptible to entrapping oxides and other impurities during casting that hinder strength and ductility. The current study examined various filter sizes and configurations with the aim of improving the mechanical properties of permanent mould cast AZ91E. The unfiltered casting had a yield strength, tensile strength and ductility of 93.3 MPa, 153.3 MPa and 2.2 % respectively and increased to 96.5 MPa, 167.0 MPa and 3.1 % with the use of a fine mesh filter within the well of the mould. The increased strength and ductility was likely due to the filtered castings containing fewer inclusions on their fracture surfaces as compared to the unfiltered castings.

Jan 1, 2011

By E. Yan, M. Mankosa, E. Wasmund

"Rougher flotation in sulfide ore applications is typically performed using conventional flotation machines. The trend over the last 100 years has been for these cells to become larger in volume with cells reaching sizes of 600 m3 and in some cases even larger. The associated foundation loadings, transport and installation requirements along with building size have also increased. In an economic environment where projects must be executed with dwindling capital supply, bigger is not always better. More recently, industry is being challenged to identify new technology that allows for improved flowsheets and financial returns through efficiency gains. To meet this challenge, a novel high-intensity two-stage flotation system known as the StackCell™ was tested to potentially achieve installation volume and required retention time gains when compared to conventional flotation circuits. The results from on-site pilot scale work conducted on typical copper concentrator rougher flotation feed showed retention time gains on the order of 6 to 9 times when compared with conventional flotation equipment. This finding has the potential of reducing capital demands for rougher flotation circuit designs when incorporating the StackCell. The results from test work will be discussed in this paper showing comparative test data generated with a Denver bench batch test, pilot StackCell and large conventional cells, all receiving the same feed.INTRODUCTION Over the last few years, a number of flotation devices have been described as two-stage or multi-stage flotation units. The idea behind multi-stage flotation devices is that flotation performance can be improved by utilizing cells with multiple specialized units in series, but not all units of the same type, as is practiced in trains of multiple tank cells or columns. Instead, each unit is specially designed to provide the optimized fluid environment for each stage of the sequential flotation process. This idea has been proposed based on a study of the fundamentals of the flotation process, and the realization that the process consists of discrete sequential processes, that are each optimized with dissimilar process conditions (Finch et al, 1995), (Zhou, 1996). Instead of executing each step simultaneously in a homogeneous fluid environment, where the conditions would be a compromise between particle collection and froth recovery, the multi-stage flotation unit operates in a set of segregated process environments, each with a set of conditions optimized for that step."

Jan 1, 2018

By V. Fiedler, H. Lachmund, R. Bruckhaus

"After producing some 50 million tonnes of crude steel the two BOF vessels at Dillinger Hüitte steel plant had to be replaced in 2000/01 due to large deformations. The new vessel design was optimised after a number of theoretical evaluations.Refractory investigations were carried out as well as temperature measurements of the shell to provide a basis for selecting a suitable steel grade. This was found in a DH-GTS development - P420M HT - with high strength at elevated temperatures and excellent weldability and made it also possible to build the vessel without any cooling system.After putting the first vessel into operation the metallurgical results showed no remarkable improvement compared to the old vessel, and were therefore not satisfying. With the help of cold model investigations the arrangement of the tuyeres was modified to improve the flow conditions and the interactions between oxygen jet and the heterogeneous plume of the bottom stirring gas. Further optimisations were obtained by a variation of lance tips and a newly developed lime addition model."

Jan 1, 2004

By R. G. Achter, E. Wilson

"This paper describes the work undertaken to renovate the Pipe Open Pit drainage system, incorporating measures to improve effluent quality. It describes the technical feat of drilling an l l-inch-diameter hole from the bottom of the open pit for a distance of 1100feet to intersect an underground drift. An innovative method of checking the inclination of this hole is also described. The paper outlines the construction required on the 1500-foot level of the underground mine to facilitate water impoundment and slime settling, as well as the instrumentation for control of water flows. IntroductionThe Manitoba Division of Inco Metals Company operates two mines-an underground mine at Thompson and an open pit at Pipe. The Pipe Open Pit is located 20 miles southwest of the City of Thompson, approximately 400 miles due north of Win nipeg(Fig. 1).Surface diamond drilling of the Pipe ore body was sta rted in 1959. In 1961, a shaft was sunk to a depth o f 1,570 feet to carry out further exploration work and obtain bulk ore samples for metallurgical testing. Site preparation to mine the upper portion of the orebody as an open pit was initiated in 1966 and by late 1967, the removal of silt and clay overburden was in progress. The contractor completed the dredging in the fall of 1969. Over 12 million cubic yards of muskeg, clay and silt was removed, which in some places was 150 feet deep over the orebody. Production began in early 1971, by which time the top three benches of rock had been excavated and the orebody exposed.The Pipe No. 2 Mine main shaft was started in 1970 and completed in 1973 at a depth of 3,160 feet. Development for exploration purposes was carried out on the 1500 and 2460 levels, and permanent pumping facilities were installed.The surface dimensions of the open pit are 2,200 by 1,600 feet, with the long axis striking northeast. The walls are designed with a 45- to 50-degree slope and the ultimate design depth is 720 feet. The present depth of the pit is 520 feet and production capacity is 6,800 tons of ore per mine day."

Jan 1, 1980

By S. Huang, K. Xia

"During crushing, rock particle is subjected to complicated loading and the particle shape influences the fragmentation outcome. Thus, a highly controlled laboratory test is desired to establish the relation between the loading and the fragmentation parameters for better understanding the rock crushing mechanism. In this work, a split Hopkinson pressure bar system in combination with high speed cameras is utilized to carry out a dynamic ball compression test, in which the spherical rock sample is adopted to avoid the shape effect. Using the elasticity theory, the history of the axial symmetric tensile stress perpendicular to the diametrical direction along the loading axis is determined, from which the loading rate and the dynamic indirect tensile strength are calculated. With the aid of the moment-trap technique, the total energy absorbed by samples is accurately determined from the one-dimensional wave analysis. Furthermore by considering the kinetic energy of the sample fragments using the high speed camera snapshots, the surface energy is determined for each sample. The relations between the loading rate and the fragmentation parameters, i.e., the number of fragments and the surface energy are established. The application of this method to a granitic rock shows that it is flexible and can be applied to the crushing study of generic brittle solids. INTRODUCTIONCrushers are widely used to break rocks to extract specific minerals in the mining industry, and to obtain rock fragments with desired size distribution in construction industry. During crushing, the breakage of particles can be in either the single-layer mode or the multi-layer mode. In the single-layer mode, the particle breakage can be assumed to be isolated, in which the interaction between particles can be ignored. It is thus suggested by many researchers that highly controlled single particle test should be performed to establish the relation between the input mechanical energy and size reduction for better understanding the crushing mechanism (Yashima et al., 1987; Tavares, 2007; Pitchumani et al., 2004)."

Jan 1, 2015

By Christoph Mueller

Mining operations today are characterized by a high degree of mechanization and single machine automation. However, their potential for cost cuttings and increase of productivity in the future is quite limited. Enabled by newest IT and networking technology, the operation of underground mines more and more can be seen from a process related point of view: Standardized Ethernet and Wireless LAN are the basis for a seamless and unified communication backbones providing an integrated infrastructure for data, telephony and video. All the components are available as off-the-shelf devices leading to an excellent cost efficiency. Basing on this standardized infrastructure, the equipment is able to become part of the infrastructure. To enable machines to be integrated smoothly into IT networks, the machine also has to have enhanced networking capabilities. This results in a new system architecture for the on board machine electronics separating real time control from IT and networking level functionality which can be implemented by a Machine Server. This machine server communicates with all kinds of above ground systems using mainly standardized formats like html or the IREDES standard. This information is made available mainly in 3D mine visualization tools. The paper explains this technology and shows implementations by two working examples.

May 1, 2006

By L. Vinnett, I. Panire, J. Yianatos, F. Ruiz

"The effect of operating variables on the flotation performance at steady-state has been extensively used for modelling and simulation of separation processes. However, the dynamic characterization of key variables such as gas dispersion (Sauter mean diameter of bubbles, D32, and bubble surface area flux, SB) and froth discharge rate is rather scarce. In this paper, an experimental design allowed identifying the dynamic response of the flotation process in a laboratory scale flotation cell, using a two-phase system, airwater. The installation included automatic control for the superficial gas rate, JG, and the pulp level. Gas dispersion was evaluated in terms of the Sauter mean diameter of a bubble population and the bubble surface area flux, SB = 6·JG/D32. The D32 was measured using the McGill Bubble Size Analyzer (MBSA) along with a semiautomated algorithm. In addition, the froth discharge velocity was measured in real-time using a Visiofroth system. The control variables were the superficial gas rate and the frother concentration (Teefroth). The dynamic characterization of the flotation process is useful to design algorithms that allow the direct or indirect control of metallurgical variables such as flotation rate, concentrate grade and mineral recovery.INTRODUCTIONGas dispersion indicates the way the gas entering a flotation machine is dispersed throughout the volume of the equipment (Schwarz & Alexander, 2006). The gas dispersion is typically characterized in terms of the bubble size (dB), gas holdup (eG), superficial gas rate (JG), and bubble surface area flux (SB =6·JG/dB) (Finch et al., 2000). All these variables have significant impact on the metallurgical performance of flotation machines. For example, Perez-Garibay et al. (2014) investigated the effect of the particle size, bubble size distribution, superficial gas rate and collector dosage on the flotation performance, for the separation of a sphalerite mineral. The authors observed that recovery increased with the superficial gas rate ranging from JG = 0.04 to JG = 0.08 cm/s and also a maximum recovery was obtained for bubble size close to 300 µm. Gorain et al. (1998) reported a strong correlation between the flotation rate constant and the bubble surface area flux in a zinc cleaner circuit. Both variables were directly proportional at three froth depths in the range of hF =7 - 45 cm. Dobby & Finch (1986) incorporated the effect of the superficial gas rate and the bubble size on the flotation efficiency by means of a kinetic model for the rate constants. Yianatos & Contreras (2010) developed a model to estimate the carrying capacity in large flotation cells by including the bubble surface area flux."

Jan 1, 2016

By Horatio Sam-Aggrey, Margot Hurlbert

Small Modular Reactors (SMRs) have been touted as having the potential to spearhead a nuclear renaissance. Proponents of this view cite the relative advantages of SMRs over larger nuclear plants as reason for their optimism. Some of these merits include potential economies of mass production deriving from modular factory production, enhanced safety features, the ability to locate reactors in remote areas, and reduced requirements for technical workforces to install and maintain SMRs. Despite the optimism surrounding SMRs, nuclear power is still a contentious issue for many Canadians. The growth of nuclear power remains adversely affected by concerns over issues such as: the health and safety consequences of nuclear power generation, environmental impacts, proliferation of nuclear materials, terrorism, and radioactive waste disposal (Organisation for Economic Co-operation and Development (OECD) and Nuclear Energy Agency (NEA), 2010). Personal values, past nuclear accidents, and lack of trust in the industry are other factors influencing attitudes toward nuclear energy (IAEA, 2005). In Canada, three main areas where SMRs could be used have been identified. These include: a. Traditional, on-grid power generation, especially in provinces looking for zero-emissions replacements for CO2-emitting coal plants. B. Remote off grid communities that currently rely on diesel or fuel oil power generation. There are 279 off-grid communities in Canada, many of them being Aboriginal communities (NRCan, ND). Notably, all communities in each of Canada’s three territories qualify as remote. C. Resource extraction sites, such as mining and oil and gas extraction sites. Many of Canada’s principal producing mines and active mineral exploration projects are located within close proximity of Aboriginal communities. Any proposed implementation of small nuclear plants would, require the support of the local population, which is largely Indigenous in the northern regions. The possible deployment of SMRs raises important issue relating to consultation and accommodation of Aboriginal concerns regarding potential SMR deployment. This is especially pertinent given Canada’s pending implementation of the provisions of UNDRIP. One of the provisions of UNDRIP that has attracted a lot of attention is the one on Free Prior and Informed Consent (FPIC). Drawing on previously published information, this paper discusses some of the challenges posed by FPIC for the deployment of SMRs. The next section outlines the provisions of FPIC and its implementation in the Canadian context. Section 3, details some of the challenges posed by FPIC to the deployment of SMRs, while Section 4 provides some concluding comments.

Jan 1, 2020

By H. Jung

The microstructures and grain sizes of nanocrystalline Co and Co-1.1 wt % P alloys were characterized by X-ray diffraction. The corrosion properties of as-deposited and annealed Co-l.l wt % P alloys and pure nanocrystalline Co were studied in a deaerated 0.1M H2SO4 solution using electrochemical measurements and XPS surface analysis. Potentiodynamic polarization testing and impedance measurements revealed that an addition of P to pure nanocrystalline Co improved the anodic-corrosion resistance, however, this enhanced corrosion resistance was compromised by the heat treatment process at 350 and 800°C. The corroded surface of annealed nanocrystalline Co-1.1 wt % P at 350°C exhibited a more porous P-enriched layer compared with the as-deposited case. A sample annealed at 800°C corroded uniformly showing second-phase precipitates. The enhanced corrosion resistance of as-deposited nanocrystalline Co-1.1 wt % P was attributed to the elemental P on the surface that acted to hinder the anodic dissolution kinetics of surface Co following an initial selective dissolution of Co. However, at a higher anodic overpotential region, the superior corrosion resistance of Co-l.l wt% P alloys did not last because of the formation of a non-protective surface film consisting of mainly hypophosphite and phosphate as well as elemental P.

Jan 1, 2005

By G. Belzil, F. Robichaud, M. Fullam

ABSTRACT The Goldex mining and milling complex is operated by Agnico-Eagle and is located in the Abitibi greenstone region of northwestern Quebec, Canada. Commercial production started in 2008. Average gold recovery is 93.5% with approximately 68% of gold recovery coming from the gravity circuit. During the first three years of production, three gravity audits were carried out to optimize performance of the gravity circuit. This paper presents results of the audits and the development of a model linking gravity and overall gold recovery, and outlines improvements made to increase gravity performance and reduce operating costs.

Jan 1, 2013