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By Liu Liching, Yan Hongjiu

"In China, the electromagnetic induction initiating system f or blasting engineering has been greatly developed in recent years. With this system, the total number of detonators that may be blasted in each round has been increased from 420 to 940 as a result of the invention of a powerful exploder, and the length of leg wires of a detonator has reached 12 metres. These developments make the new system suitable for blasting engineering of medium scale. Another attractive feature of this system is a set of testing instruments which have been invented that enable a final check of the system to be made before firing as in normal electric blasting. This paper reports on these developments.IntroductionThe initiating system for blasting explosives is one of the key factors to the success of blasting engineering, along with the progress of explosives and blasting technology. Today, we have initiating systems which may be divided into two groups: electric and non-electric. The earliest initiating system was non-electric. Many mines, as well as many projects where rock excavation was involved, were blasting non-electrically before the new nonelectric initiating system came into use in the 1970s. Electric and non-electric initiating systems are the two main initiating systems which are now being widely used, each having its advantages and disadvantages. In electric blasting, the blasting supplies and the blasting circuit can be checked before it is used or fired and, therefore, is regarded to be more reliable. In many large-scale blasting engineering where reliability of the initiation is of great concern, a checkable initiating system is preferable. The electric initiating system is not considered to be as safe as the non-electric initiating system due to its inability to choose proper ignition energy. Therefore, a non-electrical initiating system would be preferable where a hazard of stray currents or static electric energy exists. At this stage of development, one would say it would be advantageous to develop a new initiating system."

Jan 1, 1988

By Scott Hean

"Environment and the economy are interwoven irrevocably. They need to be publicly discussed, and discussed often so that new approaches can be understood and developed - approaches which are complimentary, not confrontational.Discussion on the environment is timely, from the perspective of mining engineers and developers, from that of bankers and from their shared perspective as people concerned with enhancing the quality of life, the environment and the economy.This paper will touch on how what is called the ""environmental economy "" that has evolved has become entangled with regulatory and financial issues in a way that is dangerous for all of us. It will also look at some of the promising possibilities that a world-wide and continuing concern for the environment holds out for us.For many decades, British Columbia has provided some of the key examples of the intertwining of environmental and economic issues, and, very importantly, providing the people prepared to do something about them.Canada's first environment minister, the late Jack Davis, came from British Columbia. He led this country's delegation to the Stockholm Environment Conference in 1972, a conference that laid the groundwork of principles necessary to establish some international certainty as to the legal liability for causing trans-national environmental damage.Another British Columbian who was also a national environment minister, John Fraser - now the Speaker of the House of Commons - was one of the first to dramatize the importance of acid rain and force it into the international public arena.Some 20 years ago, British Columbians made Vancouver and their province one of the major centres of opposition to the Alaska Pipeline and Tanker System - with some justification, it now seems.That a British Columbian , then MP David Anderson, was given legal recognition to sue in a United States' court against a system that was either in U.S. territory or international waters was an important precedent in, and of, itself in terms of environmental law. That old controversy is now a very current case study - in terms of dealing with the question of liability.-"

Jan 1, 1991

By Noel Billette, Jean-Luc Chouinard

"An analysis of the number of industrial injuries shows that the mining sector is particularly affected, since mining is one of the most dangerous industrial activities in Canada. This has led to numerous inquiries, and supports the need for an in-depth study aiming at the identification of factors responsible for these accidents. One possible factor often pointed at is the small crews incentives and this paper presents a scientific approach applied to check a potential link between bonus and accidents at two mining companies located in northwestern Quebec.The data required to do so were the accidents, the salaries and the number of hours worked at each mining site by the hourly paid personnel for the period from June 1979 to January 1982. The analysis of the results derived from the initial information shows different behaviours of the frequency rate vs the bonus rate at these mines. The number of days lost per accident is independent of the bonus rate at both mine sites, as shown by Khisquare tests on contingency tables.IntroductionA wide variety of opinions has been pinpointed by a survey of the literature related to individual or small crew incentives and industrial accidents in sectors of activity where it applies. In addition, various formulae of mining contracts take place in the industry on one part, and statistics from this sector indicate a high accident frequency rate on the other part. These fact s have lead to this scientific investigation which does isolate the direct impact of the bonus on workers' accidents. This study was undertaken with the full support of the Quebec Metal Mining Association.The conclusions of this preliminary phase of the researchproject are presented for two mining complexes in northwestern Quebec. The procedure is somewhat similar to the approach of the Belmoral Commission, with the underlying assumption that bonus has a direct impact on accidents."

Jan 1, 1986

By J. W. M. Porritt, M. P. Brameld, R. H. McCorkell

"Soil-gas Rn was measured by two types of alpha-particle-track detectors, by emanometry and by counting alpha-activity on Rn-decay-product collectors. Soil samples were analyzed for V by fluorimetric and fission-track methods, their Rn-emanation by decay-product collectors and their total alpha activities with track detectors. The same uranium anomalies were revealed by each method. Fission-track analysis of leaves for V showed only random variations with sampling site, and the Rn emanation of trees and leaves removed from them was barely detectable . The agreement among the other methods is significant, total alpha and collector methods are simple, rapid and useful, and a combination of methods at a site would reveal disequilibrium in the U-decay series if present. Track-Rn measurements are highly reproducible and insensitive to apparatus design and their sensitivity is not enhanced by covering the detectors with thin foil or placing them in large containers.IntroductionIn the past two years , a number of methods of geochemical exploration for uranium have been tested by the authors over the known copper-uranium occurrence at South March, Ontario. This was done partly to compare methods and partly to develop and test new ones. The South march deposit has been described a number of times by others(1,2,3,4). The methods included in this study were the following.Emanometry, Radon in soil gas is measured by pumping the gas into the cell of a scintillation counter or emanometer(5,6). Alpha particles from radon (222Rn) and its decay products (218Po and 214Po) produce scintillations of light in the silver-activated zinc sunphide (ZnS(Ag) covering the walls of the cell. These are counted by a photomultiplier tube and scaler, the count ra sulphide (ZnS(Ag) covering the walls of the cell. These are counted by a photomultiplier tube and scaler, the count rate being proportional to the concentration of radon in the soil gas."

Jan 1, 1981

By Sabine Noelte, Mikko Petteri Rantamäki, Christopher Robben, Anssi Takala, Karl-Erik Rånman

"Sensor-based ore sorting is on its way to establish as a standard option for physical separation. The increasing number of installations in coal, iron ore, industrial minerals and hard-rock metal mining segments underlines the technological and financial potential that is beginning to be untapped. The Ma’aden Phosphate installation with 1.650 t/h sensor-based ore sorter feed sets a new benchmark for the possible capacity of stationary installations. Besides stationary installations, including multiple sorters parallel in different size ranges and multiple stages, containerised semi-mobile plants are applied. These offer two advantages, first being the reduction of transport costs when applied close to the mining face. The second is that semi-mobile systems enable processing of tens of thousands of tonnes of ore in a piloting phase with an approximate duration of 0.5 to six months, verifying the results laboratory mini-bulk tests on 5–10 tonnes of sample. This is especially important for low-grade metal ores because of the inherent Fundamental Sampling Error, as well as the Grouping and Segregation Error. The paper elaborates the results achieved within the I2MINE project, treating polymetallic ore at the Boliden Concentrator in Sweden.INTRODUCTIONSensor-based ore sorting (SBS) is becoming a standard option for physical separation, applied in all segments of mineral production. The Umm Wu’al project sets the bar for stationary SBS installations with 1,650 t/h feed (TOMRA, 2015). TOMRA Sorting Solutions has finalised an agreement in cooperation with Outotec to supply Outotec-branded sorting solutions for the mining and metallurgical industry (TOMRA, 2014). Especially for low-grade hard rock metal ores, an interdisciplinary approach is necessary applying experts for all aspects of mineral production. Due to the impact on the whole process stream, it is important on the end user side to have high level managerial support and involve all disciplines to ensure that all stake holders are involved in the analysis."

Jan 1, 2016

By A. C. Moerman, C. W. Pelley, K. R. Notley

"The software package described in this paper facilitates the handling of diamond drilling information in the creation and maintenance of a geological information database. Once entered, the drilling information is stored in a direct access file for rapid retrieval. The drill data are sorted into vertical cross sections and interpretation is performed using on-screen menus to define areas of common rock type and outline the mineralized zones. Intersections of individual zones can then be examined and outlined in the longitudinal dimension at a graphics terminal.With the aid of surface-fitting routines, the digitized information can be used to generate three-dimensional models of each zone. From these models intermediate cross sections and level plans can be obtained. Several interpretations are possible and plotter drawings can be obtained. An output file for use by ore reserve estimation software can also be created.Future work will tie this software to geostatistics and mine planning programs and will be aimed at converting mainframe software to run on microcomputers. IntroductionComputer applications in underground mine planning and geology are not new. Operations such as Sherritt Gordono and Kidd Creek have been using software developed about a decade ago for handling geological information and designing stoping blocks. However, general use of such systems has been limited mainly due to the cost of hardware which was affordable only to large corporations. Software to perform the work was usually developed on site and was largely unavailable for use by others, being site-specific, or too expensive for other organizations to purchase.The increasing power of microcomputers and mini-computers, and their relatively low cost compared to the mainframe computers which were previously used, has led to the increased use of computers at mining operations of all sizes. Sources of software for use in geological interpretation and mine planning have increased in number with the proliferation of the use of microcomputers."

Jan 1, 1988

By Richard R. Klimpel, Simon Isherwood

"The industrial processing of ores containing gold and/or platinum minerals is generally mining site specific, e.g. McQuiston and Shoemaker (1985). The variety of various mineral processing techniques employed is extensive and includes size classification, gravity concentration, amalgamation, roasting, cyanidation, chemical leaching, ion ex-change/solvent extraction and froth flotation. Because of the relatively small quantities of Au and Pt that exist in the earth’ s crust, the economical recovery of Pt and Au often requires that the Au and Pt be associated with other industrially important minerals. This association of Au and Pt with other minerals can occur in very complicated mineralogical forms. It is probably safe to say that Au and Pt mineral concentration from natural ores represents some of the most complex metallurgical challenges facing the mining industry.Froth flotation is a technique that is receiving increasing industrial attention for the treatment of finely disseminated and/or associated Au and Pt minerals. The high throughput at relatively low capital cost, the potentially high selectivity, and the existence of a well-established technology for treating more abundant associated minerals such as base metal sulfides makes froth flotation an attractive processing alternative. Broadly speaking, the industrial oriented froth flotation experience with Au and Pt minerals exists in two major categories: 1) the flotation of Au and Pt associated with pyrite (or pyrrhotite) and 2) the flotation of Au and Pt associated with primarily Cu, Zn, or Cu/Pb/Zn containing sulfide ores.In the case of the precious metal association with pyrite (or pyrrhotite), much of the published froth flotation experience comes from South Africa. Typically, such processes are done using an acid side flotation with a collector such as mercaptobenzothiazole. The approach obviously is to float pyrite and hence recover associated gold. Generally the higher the total sulfur recovery the higher the gold recovery. Similar experiences of Pt associated with pyrite are a lot less common mineralogically and hence less well documented industrially. For example, in Canada, Pt is often associated with pyrrhotite but is not really processed as a straight pyrrhotite recovery.For the situation of Au and/or Pt associated with other base metal sulfide minerals, the industrial experience base is quite varied and is well documented from both North and South American operations. The nature of most of these flotation processes is to design, operate, and optimize the installed concentrator to get the Cu, Pb, Zn, Ni, etc. that is the bulk of the valuable sulfide ore and then take whatever precious metal recovery comes. Typically, such flotations are on the basic side using standard thiol collectors such as xanthates, dithiophosphates, thionocarbamates, etc. It is very common with such ores to use mixtures of collectors in order to get ""everything"". Often in such plants, there is an inherent balancing act occuring so as to operate in such a manner to reject pyrite (and pyrrhotite) in order to maintain high base metal mineral concentrate grades and yet recover sufficient precious metals such as Au and Pt minerals which often are associated with the iron sulfides being rejected. Obviously in such an environment, the smelter operation often ends up telling the concentrator what the precious metal recovery actually was in a given production period."

Jan 1, 1994

By Leslie S. Weldon

THE Lake Shore mine has been developed by two vertical shafts, with level intervals at 200 feet to the 2200 level and at 125 feet below this horizon. A main cross-cut joins the two shafts on all levels to the bottom of No. 3 shaft, and opening into these cross-cuts are the ore passes and sand passes, which are completed from surface To the 4450 level. In the west part of the mine, additional connections have been made between No. 1 and No. 2 veins, through which the major part of the ore from No. 2 vein west is routed to drifts in No. 1 vein. With the lesser amount of stoping on No. 1 vein, this arrangement permits better haulage to be maintained. The ore passes are driven at 50 degrees from the ends of short cross-cuts off the main ones and are broken through to the levels above at the mouths of the short cross-cuts, to allow for dumping on each level. The passes are controlled at each level by ball-and-chain gates and deliver to crushers and thence to loading pockets, shown in Figure 1. That system is quite flexible in that it is possible to by-pass the crushers at any station and load from either shaft at a number of different pockets. The sand passes are driven from the main cross-cuts at the same inclination as, and parallel to, the ore passes. Each is equipped with a chute and dump hole, so that waste or sand fill may be loaded or dumped at each level. A transfer door, hinged on the bottom, is installed opposite the chute for by-passing fill from one level to the next, and in the closed position protects the opening at the top of the pass. The dump hole situated beside the transfer door is protected by a grizzly with 10-inch spaces. Waste rock may be withdrawn from the system through No. 3 shaft loading pocket at the 2950 level to a waste bin on surface. Levels are developed by drifts in No. 1 and No. 2 veins for their entire lengths. Subsidiary breaks are followed at a later date, when full information has been obtained by exploration, and a considerable amount of development by sub-drifts is carried out in the upper part of the mine, particularly in more or less erratic sections of No. 1 vein.

Jan 1, 1936

By D. Beneteau, L. Feldman, D. Milne, A. Hyett, G. Snell

"Behaviour of rebar bolts subjected to combined shear and bending in underground potash mining applications is not well understood. An experimental investigation compared load applied to resin-grouted No. 25 metric rebar to theoretical load, calculated using elastic bending theory. Distributed optical sensing (fibre optic) technology allowed axial strain measurements at 1.25 mm increments along the length of rebar. Experimental procedure, data processing methods, and equipment improvements are needed for field applications of optical sensing technology. Goals were to adopt procedures for instrumenting rebar bolts, conduct tests, and develop data interpretation methods to measure field deformations and loads. ¦ KEYWORDS Fibre optic, Potash mine, Rebar bolts, Transverse load testing RÉSUMÉ On appréhende encore mal le comportement des boulons d’ancrage pour barres d’armature soumis au cisaillement et au cintrage pour un usage dans des mines de potasse souterraines. Une étude expérimentale compare la charge appliquée aux barres d’armature de diamètre n° 25 (25 mm en mesure métrique) enduites de résine époxy à la charge théorique, calculée à l’aide de la théorie de la flexion élastique. La technologie de détection optique distribuée (fibre optique) a permis d’effectuer des mesures de la contrainte axiale à des différentielles de 1,25 mm sur toute la longueur de la barre d’armature. Pour des applications sur le terrain de la technologie de détection optique, il faut envisager une procédure expérimentale, des méthodes de traitement de données et une amélioration de l’équipement. Les objectifs établis consistent à adopter des procédures d’instrumentation des boulons d’ancrage pour barres d’armature, à effectuer des essais et à mettre au point des méthodes d’interprétation des données pour mesurer les déformations et les charges sur le site. ¦ MOTS CLÉS boulons d’ancrage pour barres d’armature, essai par charge transversale, fibre optique, mine de potasseINTRODUCTION Resin grouted rebar bolts are commonly used in potash mines in the province of Saskatchewan, Canada, in areas where shear deformation is observed in the rock, or where potentially unstable ground is present, including locations such as clay seams. Opportunities to observe deformations of rebar, as shown in Figure 1, are rare because it is unusual in underground potash mining to excavate supported ground. In this situation, a block of potash containing a rebar bolt had failed."

Jan 1, 2019

By W. L. Nelson

The orebodies at Copper Mounta1n lie along and within a few hundred feet of the contact of an intrusive stock of gabbro with a series of bedded volcanic rocks of the Nicola formation. These. orebodies are all entirely within the volcanic series but in many cases the centre line of the orebody at depth is directly under the gabbro contact at the surface, due to a slight overhang of the gabbro. The gabbro and volcanic rocks are cut by a number of faults and fault zones which did not materially affect early mining operations, but, as undeground openings increased in size, the faults became more important factors in determining mining methods and procedure. Caving methods were adopted to take advantage of ground weakness in fault zones, and the effect of ground subsidence which resulted from the division of the ground 1nto large, independent blocks by faulting had to be considered in planning the sequence of mining of certain ore blocks. The faults have been divided into two classes: the longitudinal system and the cross faults, or 'breaks' as they are termed locally. In addition to the two classes of faults, the zones of parallel ore fracture also control mining procedure, but to a lesser degree, in regard to caving or ground subsidence. A third structural feature which influences mining procedure is the mine dyke system. The longitudinal or main fault system roughly parallels the gab? bro contact and is made up of a number of branches and auxiliary fractures. The main fault itself is essentially vertical, but there are a number of branches and fractures on each side of it which, though parallel m strike, may dip as flat as 50 degrees toward or away from the main fault. The full width of the main fault system has not been determined, but it is known to extend beyond the zone of mine workings. A marked schistosity follows the main fault and both schistosity and chloritic alteration have been noticed along the secondary faults of this system. Several prominent orebodies lie along the main fault and its branches. The influence of the main fault system on mining is probably related more to subsidence than to actual cavi.ng, although it is effective in both.

Jan 1, 1950

By M. B. Dusseault, A. Pirayehgar

"There is concern that hydraulic fracturing (HF) near an existing critically stressed fault could trigger slip and generate a seismic event of significance. To this end, a mathematical discontinuum model representing a naturally fractured rock (NFR) mass containing a fault is created, and fault slip response to fluid injection is studied. UDECTM is used as a Discrete Element Method (DEM) approach to define the seismic moment, and also to assess the influence of injection and slip on further shear slip propagation. In most of the simulations, slip and dilation take place along a number of oriented fracture surfaces, so it is reasonable to assume that in a real case, microseimic emissions would accompany each slip “event”. The moment magnitude scale (Mw) is used to measure the size of the microseismic event in terms of the energy released. Seismic moment is a measure of the total energy (work) released during a seismic event and is used to measure its size. From another aspect, the injection energy is calculated as the total energy required for achieving the hydraulic fracturing; this is viewed as energy to open fractures against the ambient stress field and much of it is potential stored energy that can be released and create discontinuities (work of fracture), deformations (W=F·d), heat and radiated seismic energy. A small amount of the total released energy is transformed to radiated seismic energy and another small amount results in tensile fractures. Fracture energy is used to calculate the work required to create fracture surfaces. Comparing the energy release by fracturing or seismic events indicates their contribution to the total amount of released energy or total work. The amount of energy associated with the stick-slip event is small compared with the overall work done to generate the distortions (i.e. increases in aperture by forcing them open) in the rock mass, and much of the energy in slip is lost as heat from the plastic deformation, rather than as radiated seismic energy.Microseismic monitoring shows the spatial distribution and magnitude of seismicity associated with slip of bedding planes and natural and incipient fractures. Work calculations have been undertaken to look at the various sources of energy storage (elastic strain energy) or dissipation (slip), and likely the energy losses from viscous dissipation could be addressed as well, helping to clarify HF processes in NFRs."

Jan 1, 2015

By Martin Junker

This Notwithstanding all the debates on climate change and renewable energy, coal remains the world's no.1 primary energy source even in the 21st century. In 2011 alone, total output rose from 420 million t to around 6.6 billion t. Similar growth rates are also expected for the next few years. In order to meet demand, new seams are being tapped and others mined far more extensively. In most cases, this means a move from open-cast to deep mining, and extracting coal from ever greater depths. At the same time, extraction processes such as the room and pillar method are being replaced by longwall mining, which significantly increases seam yield and therefore output. The depth of extraction with new, modified mining methods is increasing rapidly around the world - in most countries, much more rapidly than the corresponding expertise. A trade in mining expertise has therefore developed alongside the coal trade proper. Germany relies on coal imports, but can export mining industry services and advanced mining equipment. Geological conditions in the Ruhr and Saar coalfields forced Germany to drive forward development of deep working for hard coal years ago. For 30 years now, Germany has been mining hard coal at depths of over 1,000 m, and excavation is now at depths up to 1,600 m. Deeper extraction and more extensive excavation posed new challenges for both miners and equipment. Intelligent logistics, modern roadway development methods and longwall mining with ploughs or shearers combined with optimised process coordination offered pioneering solutions - solutions to enable costeffective coal extraction in the long term without compromising safety. Countries such as Russia, the Ukraine, China and the Czech Republic, where extraction is now reaching similar depths, can now benefit from this experience: they do not need to go through the same process of trial and development, but can draw on Germany's findings and ultimately save on costs. Now, Germany is offering this knowledge and experience on the international market on a client- and deposit-specific basis to continue providing mining countries with safe and cost-effective coal mining solutions and secure the energy supply. The presentation gives a number of examples of how international mining companies have used expertise from Germany to better meet the challenges of deep mining.

Aug 1, 2013

By Amauri Garcia, Rodrigo Valenzuela Reyes, Noé Cheung, José Eduardo Spinelli, Leonardo Fernandes Gomes, Rafael Kakitani

"The present research work deals with directional solidification (DS) of both Al-15wt.%Si and Al-15wt.%Si-1.5wt.%Mg alloys and further characterization of the related microstructures. The growth of refined primary Si and refined eutectic is well-known as positive in order to improve mechanical properties of the Al-Si based alloys. Alloying is also an effective method to strengthen the Al-Si based alloys, in which Mg is considered an important alloying element. However, lack of investigation regarding the effects of Mg on microstructure of these alloys is noted. So, in the present research work emphasis is given on both dendritic and eutectic growth being affected by solidification kinetics (V-growth rate and G-thermal gradient) during transient directional solidification. Experimental evolutions of microstructural spacing of both alloys have been determined which becomes possible to understand the effects of Mg coupled with the effects of growth rate on the final microstructural arrangements. A direct comparison of the tensile mechanical properties obtained for the alloys is also performed. It is shown by these correlations that the addition of Mg generally improves the tensile mechanical resistance with a minor decrease in ductility. Furthermore, the Scheil solidification path of the ternary Al-Si-Mg alloy was determined using the computational thermodynamics software Thermo-Calc, resulting in the following sequence of precipitation: Liquid-Si-a (Al)+Si- a(Al)+Si+Mg2Si. This sequence is compared with the experimental microstructure findings.INTRODUCTION Al-Si alloys containing from 4wt.% to 22wt.% Si comprise more than 90% of all Al-based castings. Parts manufactured with these alloys have a wide range of industrial applications due to their excellent castability, crack resistance, good corrosion resistance, good weldability, low density and high strength (Bray, 1976; Zolotorevsky, 2007; Reyes, 2017). Hypereutectic Al-Si alloys are widely used and researched in the automotive and aerospace sectors due to the low coefficient of thermal expansion, resistance at high temperatures, good wear resistance and high strength (Reyes, 2017; Tebib, 2013). Above the equilibrium eutectic composition (~12.5wt.%Si) the amount of primary Si particles becomes significant (Lasa, 2003). In the place of conventional hypoeutectic Al–Si, there is growing interest in hypereutectic Al–Si alloys as a candidate material for the high-quality engines because a higher volume fraction of silicon may bring about superior properties (Matsuura, 2003)"

Jan 1, 2017

By A. L. Mackie

Mine water from base metal mines is generally characterized as having high concentrations of heavy metals and low pH, resulting in the need for treatment prior to discharge to receiving waters. In this study, two high-rate clarification processes were evaluated at bench-scale: the high density sludge (HDS) process and sand-ballasted flocculation. High-rate clarification processes generate larger, denser flocs which settle more quickly than those formed in conventional flocculation processes. The HDS process uses recirculated sludge in order to seed floc formation, whereas ballasted flocculation uses microsand. Both of these high-rate processes require the addition of quicklime or hydrated lime to increase pH and allow for the formation of insoluble metal hydroxides, which can then be removed in the clarification stage. Previous research has shown that cement kiln dust (CKD) can effectively replace quicklime in conventional sedimentation processes for metals and acidity reduction to meet discharge guidelines (Mackie & Walsh 2012). CKD is an alkaline by-product of cement manufacture that has fine particle sizes and varying free lime concentrations depending on the manufacturing process. The lower solubility (i.e., free lime content) of CKD requires larger doses of material to be added to reach treatment pH targets for metals precipitation. The purpose of this study was to evaluate the treatment efficacy of high-rate clarification processes for mine water treatment, compared with conventional sedimentation, at bench-scale with conventional (i.e., lime) and novel (i.e., CKD) pH adjustment additives. Results of the study showed that all treatment processes (i.e., HDS, ballasted flocculation, and conventional sedimentation) were able to reduce target metal concentrations to well below Canadian guidelines (i.e., total zinc and total nickel < 0.5 mg/L, total copper < 0.3 mg/L) using either CKD or lime. Additionally, mine water samples treated with CKD generated lower sludge volumes than those treated with lime (i.e., 50 versus 100 mL for conventional sedimentation and 100 versus 200 mL for HDS) even with increased concentrations of solids introduced during treatment. This is most likely due to slower neutralization rates when using CKD, due to its lower solubility, as well as finer particle size. The sand-ballasted flocculation process was able to reduce final turbidity in CKD-treated samples while the HDS process did not when compared to conventional sedimentation, making the ballasted flocculation process preferable for reducing final solids concentrations in CKD-treated mine water.

Aug 1, 2013

By E. Shum, V. G. Papangelakis

"Attempts to reduce the energy requirements of water recovery in the process industries have resulted in the development of techniques to freeze and purify aqueous solutions that take advantage of local climatic conditions. Natural freezing is a technique by which aqueous solution is exposed to freezing air temperatures, resulting in the production of a layer of ice on the surface. As the ice layer tends to exclude solute, a concentrated solution is formed. However, due to dendritic ice growth along the advancing growth front, small quantities of solution are trapped within the ice, decreasing the purity of the water recovered. The ice produced can then be allowed to thaw in order to recover the water using low grade plant heat. As the more concentrated parts of the ice melt first, a purer ice remains. In this study, 0.5 molal magnesium sulfate solution was frozen under simulated natural freezing conditions at temperatures between –1 and –6°C. The ice produced was thawed at either room temperature or 50°C. Up to 40% of the water was recovered with a separation efficiency of above 95% when thawing at room temperature.INTRODUCTION Advances in refrigeration technology have made low-temperature separation processes increasingly appealing to an energy-conscious society. The latent heat of freezing of water is about seven times smaller than the latent heat of evaporation providing freezing with a strong thermodynamic advantage over the traditional evaporation-based processes. The result is the development of a wide variety of processes that fall under the term Freeze Concentration (FC), which describes a variety of techniques in which an aqueous solution is concentrated by the creation and subsequent removal of ice (Williams et al., 2015).FC processes exploit the freezing behaviour of aqueous solutions as dictated by their respective water-salt phase diagram. When, for example, a solution of magnesium sulfate is cooled to a temperature below its freezing point but above its eutectic temperature of –4.1?C (Figure 1) then, if the concentration of MgSO4 is less than the eutectic concentration (about 1.7 m), ice will form until the solution concentration reaches its equilibrium concentration at the given temperature on its phase diagram. If the concentration is greater than the eutectic concentration, MgSO4 solid will precipitate until the concentration reaches equilibrium. It is for this reason that ice can be selectively produced in FC processes (Randall & Nathoo, 2015)."

Jan 1, 2017

By F. Bélanger, P. Fournier

"The Detour Lake Mine is located in northeastern, Ontario, approximately 300 kilometres northeast of Timmins and 185 kilometres by road northeast of Cochrane. The mill processes a free milling gold ore using both gravity circuit and leaching with a state-of-the-art carbon-in-pulp circuit. The grinding circuit is a conventional SABC circuit with a tonnage of 60,000 mtpd at 95 microns with the addition of pre-crushing upstream of the SAG mill. This paper will highlight the impact of a pre-detox thickener on the cyanide consumption in a gold leach/CIP circuit. While the pre-detox thickener was built with the intention to reduce the cyanide consumption, the opposite effect was observed when the thickener was put temporarily offline. It will show how the re-circulation of a dilute stream of cyanide into the grinding circuit, increases the overall slurry temperature and by the same effect, increases the dissolved copper concentration in the leaching circuit which subsequently increases the cyanide consumption.INTRODUCTION The processing plant is using a conventional gravity, cyanidation and carbon-in-pulp processing facility currently operating at approximately 60,000 mtpd. The comminution process is based on two parallel SABC lines, each fed from an individual reclaim tunnel. Cyclone overflow from the two ball mill circuits is directed to the pre-leach thickener. The thickener underflow is distributed to two (4) parallel leach trains of 5 tanks each, each leach train providing 29 hours of leaching retention time. After leaching, the gold in solution is recovered in a carrousel CIP circuit consisting of two (2) parallel leach trains. Tails from each CIP circuit are combined and sent to a pre-detox thickener to recover as much cyanide solution as is economically practical before being diluted and fed to a cyanide destruction system. The detoxified tailings are sent to the Tailing Management Area (TMA) for settling and retention. The excess water from the TMA is reclaimed and recycled back to the processing plant. The pre-detox thickener is designed to thicken the leach tails slurry to 64% solids with the overflow being recycled to the grinding circuit. It is estimated that the pre-detox thickener will recover 60% of the cyanide in the CIP discharge solution. It is also designed to reduce the SO2 consumption by diluting the thickener U/F to 50% solid using the tailing pond reclaim water which has very little cyanide content."

Jan 1, 2019

By A. McLean, X. Men

"With high quality sulphide ores diminishing, nickel production from low-grade oxide ores becomes increasingly important and a major concern is the high sulphur level in the resultant ferro-alloy. So, desulphurization of the ferronickel becomes an important consideration. Red mud is a waste product of alumina industry and causes adverse effects on environment. In the present study, experiments were conducted to determine feasibility of using red mud as a major ingredient of refining fluxes for ferronickel desulphurization. In view of their influence on slag-metal separation, contact angle measurements between ferronickel alloy and molten slag made from red mud were also determined. INTRODUCTIONNickel is a very important metal that finds a variety of applications including consumer products, building, automotive and electronics. There are two kinds of ores containing nickel, namely oxide and sulphide ores. Anthony and Flett [1] reported that about 55% of the world’s nickel production originates from sulphide ores while the balance of new nickel is derived from oxide ores. However, oxide ores count for about 70% of the worldwide reserves. Apparently, oxide ores will play a significant role in the production of nickel in the future. Most of nickel produced using oxide ores is from pyrometallurgical processes namely ferronickel smelting or matte smelting. So, we can draw such a conclusion that using nickel oxide ores to produce ferronickel will become more and more important .Before used to make stainless steel or other materials, the ferronickel need pretreatment to remove impurities especially sulphur and phosphorus that will cause negative influence on the properties of steel product. Red mud is a by-product during alumina production. Usually, about 1 to 1.5 tons of red mud will be produced when we get 1 ton of alumina. About 90 million tones of red mud are produced every year in the world [2].Red mud is deemed as a toxic industrial waste which needs treatment before discharge because of alkalinity and chemical and mineralogical properties. Until now, the main means to deal with red mud are to store it in impoundments or dump it in the sea depending on local environment. Nowadays, how to recycle red mud has become a pretty hot topic around the world. The last 40 years have witnessed a lot of patent applications concerning the ways of red mud utilization. However, a little amount of red mud is consumed until now due to the limitations of recycling red mud."

Jan 1, 2012

By C. J. Hodgson

"IntroductionThe Kitsault molybdenum mine is located 600 m above sea level on Lime Creek at the head of the inlet of Alice Arm, 140 km northeast of Prince Rupert, British Columbia (Fig. 1). The area of the mine is characterized by rugged topography and deeply incised creeks which drain into the Portland Canal - Observatory Inlet fjord system. Access is by air service or boat from Prince Rupert, or by secondary road from Terrace.HistoryThe deposit was first staked in 1911, at which time the main feature of interest was a polymetallic vein with values in silver located southeast of the molybdenum deposit (Woodcock and Carter, 1976). Turnbull (1916) was the first to describe the extensive zone of crisscross quartz stringers with Mo~ on the banks of Lime Creek which, although it clearly had the potential to contain a major deposit, was not then believed to be of commercial value.Following sporadic work through the 1920s and early 1930s, the ground lay idle unti11959 when Kennco Explorations (Western) Ltd. commenced a program of systematic diamond drilling on the stockwork-type mineralization adjacent to Lime Creek. In late 1964, Kennco announced the presence of an orebody of 36 million tonnes averaging 0.230Jo Mo~ which was to be put into production at a rate of 5450 tonnes per day under a new subsidiary, British Columbia Molybdenum Ltd., at a forecasted capital cost of $20 million (revised to $30 million).Between January 1968 and April 1972, 10.4 million kg of molybdenum were produced from 9.3 million tonnes of ore, at an average mill recovery of 90%, and with an average mill head grade of 0.206% Mo~. The ore was mined in an open pit with a waste to ore stip ratio of 1.5 to 1. The. operation was terminated in 1972 due to weak molybdenum prices and escalating costs.In May 1973, the deposit was purchased by Climax Molybdenum Corporation of British Columbia, Limited and in 1979 transferred to affiliate Amax of Canada Limited. In early 1979, following completion of detailed feasibility studies, Arnax commenced construction of the 10 900 tonne per day Kitsault mine and mill complex based on a mineable open pit reserve of 95 million tonnes at an average grade of 0.192% Mo~, sufficient for a mine life of 25 years. The original capital cost estimate was $143 million; it subsequently escalated to approximately $200 million."

Jan 1, 1995

By L. Rao, P. Carabin

"Advances in thermal plasma torches have resulted in this technology becoming a commercially viable solution for chemical and metallurgical processes. The main advantages of plasma are its ability to control process chemistry and to build small footprint reactors due to its high energy density and reactivity of the free radicals that are produced. This paper focuses on thermal plasmas produced by DC torches and their applications for waste treatment, production of high purity metals, and nanomaterials. Both transferred and non-transferred torches have been used as either a source of heat or as a reagent in various industrial processes. INTRODUCTIONPlasma technology has become an ideal solution for many industrial chemical, metallurgical and mechanical processes (Venkataramani, 2002). Thermal plasma powered systems are typical used for their unique characteristics such as high energy densities, high temperatures, small installation sizes, rapid startup and shutdown features, controlled process chemistry and use of electrical energy as a heating source resulting in decoupling the heat source from the oxygen flow rate (Heberlein et al., 2007). Depending on the nature of the main plasma forming gas used, thermal plasma systems offer a high temperature heating source coupled with a highly reactive plasma plume rich in free radicals and ions which promote otherwise hard to drive chemical reactions, as well as high heat transfer rates.Industrial plasmas can be broadly classified as thermal plasmas and non thermal plasmas. Thermal plasmas are atmospheric pressure plasmas characterized by local thermodynamic equilibrium (Boulos et al., 1994). Thermal plasmas are typically established between any two current conducting electrodes separated by an insulator. A plasma forming gas is blown between the two conducting electrodes resulting in a high temperature plasma plume. A plasma torch generates and maintains an electrically conducting gas column between the two electrodes; a cathode (negative electrode) and an anode (positive electrode) (Camacho, 1988). If both electrodes are housed in a single housing, resulting in an arc enclosed inside the plasma torch, then such plasma generating torches are termed as non-transferred (NT) plasma torches, whereas, if the second electrode is external to the torch, creating an arc transferred between the cathode and the working piece, then such plasma torches are termed as transferred (T) plasma torches. Depending on the source of the electricity which can be either direct current (DC), alternating current (AC) at main network frequency or at radio frequency (RF), the plasma torches are classified as DC, AC or RF plasma torches (Venkataramani, 2002)."

Jan 1, 2012

By L. Nichols, H. Keele, K. Czajewski

"The oil sands industry in northern Alberta disposes of the by-products of bitumen extraction into tailings ponds. The mix of tailings sand and mature fines, with gypsum added as a flocculent, constitute a waste stream called consolidated tailings or composite tails (CT). The two main challenges of the tailings disposal process are a chronic shortage of economical containment and the potential mobility of CT deposits in the event of breaching by, for example, long-term erosion, which makes abandonment of tailings ponds more complicated. Thick deposits of CT do not consolidate in adequate time to provide reassurance for abandonment of the CT containment facilities and that is why the containment must be adequately robust or the properties of CT improved for long-term stability. This paper addresses potential solutions to these problems by applying the gravity drainage of CT and quasi-homogeneous construction for containment facilities.Gravity drainage is suggested as a mechanism to facilitate the consolidation of CT in critical areas of the ponds by inserting internal sand layers within the body of the deposit. Critical CT areas are defined as areas around the perimeter of containment. The advantages of this solution are longterm stability of the disposal area by creation of stiff deposits around its perimeter and facilitation of reclamation efforts on its surface.The present solution for containment of CT deposits is not favourable from a materials balance point of view. In general, sand has become a valuable resource as it both forms the primary component of the CT matrix and becomes the containment building material. If sand is required for CT, less is available for dyke construction. This puts additional pressure on the industry as a whole to maximize the use of overburden and interburden for construction of tailings containment structures. Typically, only 50 percent of dry mine waste (overburden + interburden) is incorporated into dyke construction. The remainder of dry waste is stored either in waste dumps or in tailings ponds, consuming valuable in-pond space originally allocated to CT material. Quasi-homogeneous dykes are proposed to contain CT deposits rather than traditional minimum cross-section dykes built of overburden and/or cell sand. The quasi-homogeneous dyke may change the unfavourable material balance and lessen the construction constraints that are presently imposed on the mine/waste disposal process."

Jan 1, 2008