Search Documents

By Z. Zhong, Y. J. Hu, D. Elsworth

"A standard triaxial apparatus was set up to perform sliding tests. Slide-hold-slide (SHS) tests, with hold times from 10 to 3000s and at invariant effective confining pressures of 1MPa, incremented to 3MPa and then 5MPa, were conducted on both dry and wet split limestone samples to study the evolution of fracture friction under holding and sliding processes. Limestone fracture shows time-dependent behaviors in SHS tests, friction drops (creep) gradually during hold periods, and recovers (healing) after re-slide. Frictional healing and creep increment are proportional with hold times, while are inversely proportional with effective confining pressures. Moreover, the wet sample displays heavier healing than the dry one, whereas creep increment of the wet sample is lighter than the dry one. We conclude that the evolution of limestone fracture friction is significantly controlled by mechanisms of fluid-rock interactions.INTRODUCTIONThe evolution of the mechanical properties of fractures and faults exerts a critical influence on the behavior of fault zones and on the earthquake cycle. In addition to natural processes, this behavior has broad implications in induced seismicity and influences oil and gas production, the sequestration of carbon dioxide and the disposal of nuclear wastes.Fault friction is typically measured by slide-hold-slide experiments, where fault gouge is deformed at low velocities (<0.1 mms-1) and over small shear offsets (<30 mm) (Dieterich, 1979, 1981; Mair et al., 1999; Marone, 1998, 2004) or alternatively with rotary-shear devices, that enable higher slip velocities over very large shear offsets (Hirose & Shimamoto, 2005; Mizoguchi et al., 2006). Typical experiments (Dieterich, 1972; Marone, 1998; Olsen et al., 1998; Bos & Spiers, 2002; Tenthorey et al., 2003) show significant creep during the holding period and frictional healing (strength gain) upon reloading. Fault creep and healing are often time-dependent with longer hold times typically resulting in greater magnitudes of creep and in strength gains (Tenthorey et al., 2003; Yasuhara et al., 2005; Carpenter et al., 2012). Frictional healing may arise from growth of grain contacts (potentially induced by pressure solution) (Rabinowicz, 1951; Dieterich & Kilgore, 1994; Yasuhara et al., 2005), strengthening of the contacts (correlates with cementation) (Hickman & Evans, 1992; Beeler et al., 1994; Karner et al., 1997; Olsen et al., 1998; Muhuri et al., 2003; Tenthorey et al., 2003; Li et al., 2011) or via other changes in state including changes in porosity and rearrangement of force chains. Experimental evidence (Dieterich, 1979, 1981; Johnson, 1981; Marone & Scholz, 1988; Marone, 1998; Saffer & Marone, 2003; Moore & Lockner, 2007; Ikari et al., 2007; Ikari et al., 2011; Carpenter et al., 2012) quantifies both the mode and magnitude of second-order changes in the velocity-dependent (including velocity- strengthening and weakening) evolution of frictional strength."

Jan 1, 2015

By Michael S. Moats, Addin Pranowo, Gerardo R. F. Alvear F., Nigel Aslin

The toxicity of arsenic is well known and documented. However, the presence of arsenic in copper electrorefining anodes and electrolyte is critically necessary to produce high quality cathode. Arsenic as well as antimony and bismuth concentrations vary depending on their relative concentration in the anode copper received from the smelter. This paper will discuss the behaviour and benefits of arsenic in copper electrorefining and the detrimental effects of antimony and bismuth on cathode quality and tankhouse performance. This will include the minimization and mitigation of problems associated with arsenic, antimony and bismuth including cathode contamination, floating slimes, scaling and anode passivation.

Jan 1, 2014

By Philippe Ribagnac, Bruno Courtaud, Jean-Marie Lambert, Valérie Weigel

Mabounié is a polymetallic deposit in Gabon containing niobium (1.2 wt.%), tantalum (0.03 wt.%), rare earth elements (REE) (1.4 wt.%) and uranium (0.03 wt.%) that are mostly carried by pyrochlore minerals. The classical route for niobium recovery is a pyrometallurgical process that consists of pyrochlore flotation, dephosphorization of the niobium concentrate and aluminothermy to obtain ferro-niobium. Pyrochlore concentrates can also be treated by a hydrometallurgical process with a mixture of sulfuric and hydrofluoric acid leaching followed by solvent extraction to separate niobium and tantalum. However, the first process does not allow recovering REE and the second one is not environmentally friendly because of the use of HF. ERAMET is looking for new opportunities on strategic metals market and has recently developed a hydrometallurgical process for the recovery of all these metals based on a sulfuric acid treatment of the ore (Donati, Courtaud, & Weigel, 2014). As for most of the oxidized ore processes, the major impurity is iron (~30 wt.%). Thus, in this hydrometallurgical process, the question arises of the elimination and stabilization of the iron as residue. The iron in solution is ferrous; goethite precipitation seems the most adapted crystalline form. Lab and pilot tests have shown kinetics of precipitation of 6 g.L-1.h-1 at pH 4.5. To reduce the salt discharge, sodium-jarosite form is more adapted, but at pH 2, 95°C only 30% of iron is precipitated in 5 hours, compared to 100% in goethite form at the same time and pH 2, 80°C. Different ways of oxidation were tested, air, pure O2, activated carbon, without drastic effect. The more efficient was ozone use, at room temperature; the kinetics of oxidation is five time faster than the other ways with 30% of sodium removed from the solution.

Jan 1, 2016

By Sean Dessureault

Technological Revolution ?Information age ? brought about by: ?Low-cost processing, storage, information sharing ?Shared & integrated data ? Increasingly more sensors & data sources From data collection to data use ??Data mining??analyzing the huge volumes of data ?Key example used in MBAs: ?Wal-mart, Google Explosion of data, trickle of information ?Paralysis of action Outline Trying to provoke- What drives innovation now? ?Challenges in mining Data Action Approach ?Examples from research & consulting upon request Conclusions ?Unaddressed challenges (Research Opportunities)

May 1, 2006

By Roger M. Nendick

"IntroductionThe stages of development for bringing a mill into production demand that a flowsheet for the process be developed at an early stage. Detailed engineering of the mill may then proceed without duplication of engineering effort on alternative processes. In the case of gold and silver mills, several process alternatives exist which must be evaluated before the flowsheet is finally fixed. The recently developed Carbon-in-Pulp process has added to this choice.Clearly, the major factor in determining the most suitable process is the amenability of the particular ore to a given treatment. Other factors that will also influence the choice are location, climate, power and water supplies, and available capital.Gold and silver milling processes can be divided into three main types:Gravity Concentration;Flotation;Cyanidation.Combinations of the three are frequently used; i.e. flotation followed by cyanidation of the concentrate, or gravity concentration in the grinding circuit with cyanidation of the classifier overflow. The following are the salient advantages and disadvantages of the three main alternatives.Gravity ConcentrationGravity concentration is suitable for free-milling gold ores, i.e. ores in which the gold is relatively coarse and free from sulphides. It should be noted that just because an ore is high grade, it is not necessarily free milling.Gravity concentration methods include jigs, shaking tables, spirals and more recently Reichert cone concentrators. All methods exploit the density difference between gold, which has a specific gravity greater than 15, depending on how it is alloyed, and the gangue, which usually has a specific gravity range of 2.7 to 3.5.A concentrate grade of several hundred ounces per ton might be produced, but recoveries by gravity alone are generally low in comparison to other methods. For this reason, gravity circuits are usually used in combination with flotation or cyanidation circuits."

Jan 1, 1983

By MacKenzie Dickson

"INTRODUCTIONGold is found in many natural states, native nuggets, loose grain in river sands, veins in rock, as well as associated with sulphides, oxides, etc., disseminated more or less minutely.Its high value in recent years has spawned new technologies to win values from sources once considered too lean for commercial recovery. Cyanide leaching from fine milled ore or tailings, from upstream recovery operations, coupled with adsorption onto relatively coarse granules of activated carbon, concentrate the gold enough to be converted to bullion by time-honoured methods.The carbon granules are fragile. Size degradation in handling must be minimized not only to prevent loss of gold on fine fragments of carbon that could pass with the barren tailings, but to maintain adsorption efficiency. Besides, the granulated carbon itself is costly. Circular Vibroscreens are winning acceptance for dewatering carbon granules thanks to their gentle yet effective performance in gold recovery systems around the world. The system at Equity Silver is illustrative.BackgroundEquity Silver Mines Ltd., Houston, British Columbia, first began production in September 1980. In 1984, a feasibility study showed that additional gold and silver could be economically recovered through cyanide leaching of the flotation tailing.Although the carbon-in-pulp (ClP) or carbon-in-leach (CIL) process is commonly used for extracting precious metals from slurries, the use of circular vibrating screens, and Vibroscreens in particular, for the carbon circuit is an innovation that cuts costs.Rectangular vibrating screens have been the conventional separation mechanism of choice across North America. In contrast, the use of Vibroscreens has been proven to be less costly for the initial purchase and installation. Also, these circular vibrating screens require less power and less space. A description of Equity Silver Mines&apos; process shows how these screens have been successfully integrated into the system."

Jan 1, 1989

"Eastern Canadian thermal coal mined at Minto, New Brunswick, and in western Cape Breton at St. Rose, Inverness County, Nova Scotia, contains 6-8 per cent sulphur. Both coals are high volatile bituminous. types. The Minto coal averages 18 per cent ash, 7 per cent sulphur and has a Btu rating of 11,900 per lb. Pyritic sulphur accounts for 75 per cent of the total sulphur and the remainder is largely organic; the amount of sulphate sulphur is negligible. This means that in a 7 per cent sulphur coal, 2 per cent is organic and 5 per cent pyritic sulphur. Pit-head St. Rose coal contains approximately 10 per cent ash, 7 per cent sulphur and 11 ,800 Btu/lb. The distribution of sulphur is different to that of Minto coal because about 12 per cent sulphate sulphur is present. Thus, a 7 per cent sulphur coal would contain 4 per cent pyritic, 2 per cent organic and 1 per cent sulphate sulphur.The iron sulphides, pyrite and, to a lesser extent, marcassite occur over a wide size range from large sulphur balls to crack fillings, lenses, and individual crystals and framboids. Coarse-grained pyrite and sulphur balls are easily separated during washing. It is the fine grain sizes, intimately mixed with the coal constituents, that are difficult to separate. Quantimet analyses of the fine grains in coal from both localities showed that about 50 per cent were smaller than 30 micron(1).No efforts have been made to wash the Inverness coal of Nova Scotia and in 1980 washing ceased in New Brunswick. Previously, the Minto coal was processed in an antiquated jig circuit which reduced ash from the strip operation run-of-mine value of more than 20 per cent to less than 18. More effective removal of ash and sulphur could clearly be obtained. Laboratory tests on both Minto and St. Rose coal have shown that ash rejection rates of 40 to 50 per cent at yields of 70-80 per cent were possible(2), Little information on sulphur removal was available."

Jan 1, 1981

By A. Khan, C. Ciriello, W. Oliveira

"The purpose of this paper is to demonstrate the value of Operational Readiness (OR) in the successful delivery of capital projects. OR can be captured using several technical and practical definitions:• To ensure that the capital project being delivered is capable of an effective start up, ramp-up as planned and a sustainable productive operation.• To function as intended when started, with the appropriate people able to run it.• To increase the value in delivering the product/service through established systems, processes and procedures, while simultaneously minimize risk to both the project and the business.Historically, mining and metallurgical capital projects concentrate efforts on cost and scheduling during the implementation and design phases. The completion and subsequent handover of the operational unit is the main objective.When reviewing the mining and metallurgical industry’s capital projects, it seems many poor results and errors in terms of project design and ramp-up are self-inflicted, with the project neglecting to account for key aspects of the future operational plant. Interestingly, if the problems for the future plant occur during project implementation, they can be prevented during the project phase. It is required to connect the operational knowledge with the project in order to ensure an effective start up and operation.The benefits of OR will be shown to be applicable not only to mining capital projects but also to sustaining and business transformation projects. The paper also includes industrial examples showing both proper and improper use of OR techniques.The paper will cover 3 main areas of Operational Readiness: the OR model, OR related risk elements (i.e., risks associated specifically with the improper execution of OR) and conclude with OR specific case studies."

Jan 1, 2016

By Waghmarem Umesh V., Vinay Jain, Pradip, Beena Rai

"The hematite surface plays a vital role in several processes of industrial and technological importance, such as, in beneficiation of iron ores, catalysis, and geochemical processes. Therefore, it is imperative to have a clear description of the structure and composition of this surface, which forms the underlying basis for such a wide variety of complex phenomena. Particularly, from the perspective of beneficiation of Indian alumina-rich iron ore slimes and fines, a clear understanding of the surface structure and composition of hematite surface and its difference from the alumina surface will facilitate identification and/or design of selective reagents which can be used to separate the two minerals. In this work, the hydroxylated hematite (0001) surface was modeled using density functional theory (DFT). The computed planar relaxations are consistent with X-Ray crystal truncation rod diffraction results reported in the literature. It is demonstrated that only one type of surface, that is, the (OH)3-Fe-H3O3-R is stable under aqueous conditions, as the relaxations for the second type of termination: (OH)3-Fe-Fe-R do not match quantitatively with experiments suggesting that its coexistence with the (OH)3-Fe-H3O3-R surface is unlikely. A detailed analysis of this surface shows that two of the three (OH) groups are not connected to the Fe surface atom. Instead, they form H2O molecules by combining with the H-atoms on the O-layer beneath the Fe-terminated surface leaving the Fe surface-atom bonded to only one OH-group. This undercoordination of Fe-atoms on the surface explains its higher reactivity as compared to hydroxylated alumina.INTRODUCTIONThe hematite surface plays a vital role in several processes of industrial and technological importance. A few examples are: flocculation/froth flotation of iron ores (Ravishankar, Pradip & Khosla, 1995), catalysis (Muhler, Schlögl & Ertl, 1992; Samsonov, 1973), and geochemical processes (Brown, Henrich, Casey, Clark, Eggleston & Felmy et al., 1999; Sposito, 1989). Therefore, it is imperative to have a clear description of the structure and composition of the hematite surface, which form the underlying basis for such a wide variety of complex phenomena. Particularly, from the perspective of beneficiation of Indian alumina-rich iron ore slimes and fines, this is an important problem, as a clear understanding of the surface structure and composition of hematite surface and its difference from the alumina surface will facilitate identification and/or design of selective reagents which can be used to separate the two minerals. Our current research focus is at employing molecular modeling techniques to come up with such selective reagents and then test their efficacy in the practical scenario through selective dispersion – flocculation labscale experiments and subsequently plant trials. Obviously, the correct structure and composition of the hydroxylated hematite surface is a fundamental prerequisite for such investigations."

Jan 1, 2016

By John E. Udd

"Mining research at CANMET is delivered by the Mining Research Laboratories Division (MRL), with facilities in Sydney, Nova Scotia; Bells Comers, Elliot Lake, and Sudbury, Ontario; and Val d&apos;Or, Quebec. This paper gives an overview of the work of the Division as of early 1993. An addendum, at the end of this paper, describes the organizational changes that will take place in April 1994. IntroductionCANMET, the acronym for the Canada Centre for Mineral and Energy Technology, is the name by which the organization founded in 1907 as the federal Mines Branch is presently known. The name was changed in 1974, in order to give recognition to the importance of both minerals and energy technologies in the activities of the Branch. Further reorganizations have taken place since 1987. As a result of these, CANMET has assumed the status of a Sector (equivalent to the Geological Survey of Canada) and has been organized into four branches.The organization as a whole functions in an environment of semi-matrix management. Some of the ""line"" Divisions report to Directors-General for the appropriate branches while other, larger divisions, report directly to the Assistant Deputy Minister.The second dimension of the organizational matrix is composed of a number of ""staff&apos; divisions which report to the DirectorGeneral for Policy and Planning. Each of these Divisions performs a function which is common to all of the ""line"" divisions. For example, the Research Projects Office is responsible for monitoring the range of research activities of CANMET and for recommending allocations of funds for both research contracts and operating budgets. The Library and Documentation Services Division is responsible for CANMET publications and information services. Likewise, the Engineering and Technical Services Division provides support through, among other services, the construction of specialized laboratory equipment."

Jan 1, 1994

By Mariam Melashvili, Chris Fleming, Rene Jackman

"A study was undertaken to develop a flowsheet for treating a high grade silver/gold ore body by cyanidation, CCD and Merrill Crowe precipitation, with tailings detoxification to meet stringent limits for cyanide and thiocyanate concentrations in the final discharge. In one of the flowsheets examined, free cyanide in the final leach liquor was recycled to the mill where it was hoped that it might react beneficially with fresh ore to reduce the cyanide makeup by an equivalent amount. However, laboratory results suggested that most of the cyanide recycled to the mill reacted with iron forming a soluble ferrocyanide complex. This would have had a negative impact on cost, not only because the cyanide recycled to the mill could not be claimed as a credit, but also because the presence of ferrocyanide in the final tailings meant copper sulphate had to be added to the SO2/air cyanide detoxification process to precipitate the iron cyanide. In an alternative flowsheet that was examined, recycling of cyanide to the mill was abandoned, and this allowed the milled slurry to be pre-aerated in a single mixing tank before cyanidation. During alkaline pre-aeration, any ferrous ions that are generated in the mill are oxidized to ferric ions, which, unlike ferrous ions, are inert in cyanide solution. This significantly reduced cyanide consumption and lowered costs in the detoxification of the tailings. In this paper, the cost implication of these two circuits is compared.INTRODUCTIONMost of the gold that is produced by hydrometallurgical processes is recovered by cyanide leaching. The reason for that lies in the superior thermodynamic stability of the gold cyanide complex and the favorable economics this confers on the process (Fleming, 2014). However, the gold in most of the gold-bearing ores processed is a very minor component, and therefore, the leaching response is often dominated by that of the host minerals.Many gold ores treated by the cyanidation process contain iron minerals. Cyanide reacts readily with some iron minerals forming ferrocyanide as one of the compounds. Ferrocyanide is stable when solutions are exposed to air and does not retard the dissolution of gold. In fact, it has been claimed that potassium ferrocyanide in the presence of oxygen is a weak solvent for gold (Hedley & Tabachnick, 1958). The main problem with the formation of this compound is the consumption of cyanide. Ferrocyanide contains six molecules of cyanide for each atom of iron {Fe(CN)64-} so the wasteful consumption of cyanide is greatly increased by its presence. Thiocyanate is also often produced during the reaction of certain iron sulphide minerals with cyanide, and contributes to the wasteful consumption of cyanide. In addition, environmental regulations in certain countries demand that thiocyanate be removed from the tailings, and this adds significantly to the cost of tailings treatment."

Jan 1, 2016

By D. W. I. Griffiths, J. C. Bélanger

"La Division Matagami de Minéraux Noranda inc. a profité de la construction de sa nouvelle mine (Isle-Dieu) pour continuer sa modernisation.L&apos;utilisation d&apos;automates programmables (PLC) s&apos;est intensifiée pour contrôler et automatiser de plus en plus d&apos;équipement dans ses minessouterraines, au concentrateur, dans les atelierset bureaux.Tous les PLC comrnuniquent entre eux et les IBM PC sont utilisespour exterioriser l&apos;information et faire Ie lien entre Ie personnel et les automates. Le système Central IBM 36 emmagasine les donnees traitees par lesPLC (a travers les PC). Ces donnees sont transmisesau programme d&apos;entretien de chaque equipement et compile les heures, cycles, tonnes, etc., fait imprimer automatiquement les feuilles d&apos;entretien préventif &apos;, etc.Les principales realisations adate et en progrès pour la mine Isle-Dieu sont:• automation complète du système de remblayage du concentrateur ala mine Isle-Dieu situee a2 km de la mine principale;• automation du fonctionnnement des ventilateurs qui comprendra la position des portes de ventilation, Ie chauffage, Ie débit, etc.;• automation des pompes, surveillance des sous stat ions électriques, concentration des gas (CO, SO2) à chaque niveau en tout temps et du propane à l&apos;entree d&apos;air frais;• tonnage et cycles par Ie treuil; et• contrôle d&apos;autres équipernents à differénts niveaux. D&apos;autres travaux furent egalement effectues par les employes, tels que:• la conversion d&apos;un Scooptram ST5A de diesel à électrique;• I&apos;installation complete de systèmes de depoussierage a I&apos;eau (scrubber) pour le concassage secondaire;• I&apos;installation d&apos;un marteau hydraulique avant les concasseurs coniques;• la rénovation des cellules de flottation ; et• l&apos;installation d&apos;un système de gardiennage par caméra. D&apos;autres projets seront effectués ou étudiés:• automation d&apos;un Scooptram"";• automation d&apos;un marteau hydraulique;• récupération de Scooptrarn; téléguidé:• gestion de l&apos;énergie, contrôle de la demande, surveillance et contrôle de l&apos;équipernent de machines fixes(compresseurs, boulloires, pompes à vide, génératrices, débits et pression d&apos;air, eau pour diagnostiquer les problémes, analyser les coûts;• concasseur portatif pour fabriquer de la pierre concassée pour utiliser dans les chemins sous-terre, et• utilisation de micro-ondes pour transmettre les images captées par camera."

Jan 1, 1990

"Line Creek Resources Ltd. produces three million clean tonnes per year of metallurgical and thermal coals for the international market. The mining operation, which is located in the mountains of southeastern British Columbia, produces coal from four pits by conventional truck and shovel methods.The South Pit, which has been in operation since J982, has a benched highwall 490 m in height. Shallow dipping coal seams at the base of the syncline formation will be exposed in the highwall over the next four years of pit life. The first of these seams, which will measure 500 m in final highwall length, was exposed in J994. An auger, powered by a 1250 hp engine capable of cutting 2.2 m diameter holes, was brought on site in J994 and has proven to be a successful means of recovering coal from the J3 m thick seam.This paper covers the background to the selection of an auger mining method to recover highwall reserves together with the geotechnical and methane considerations for this approach.IntroductionLine Creek Resources Ltd., a wholly owned subsidiary of Manalta Coal Ltd., owns and operates the Line Creek mine. The operation which consists of four active pits, a processing plant, and rail loadout is located in the southeastern coal fields of British Columbia. Figure 1 shows the general area.The coal bearing sequence at the mine is preserved on the east and west limbs of a major syncline. Up to twelve mineable coal seams, ranging in thickness from 2 m to 15 m are contained in the formation. Within the pits the seams are stratigraphically consistent and structurally undisturbed, except for small-scale thrust faults that create multiple repeats of the seams."

Jan 1, 1995

By Walter Curlook

"With Inco recording record profits over the past year, most people have forgotten the painful years of the last recession/depression when Inco was losing up to a million dollars a day for a short period . While high nickel prices have contributed to the exceptional recent profits, the real achievements stem from the tremendous advances in productivity and cost improvements that make Inco one of the lowest cost nickel producers.Some of the major advances in productivity have occurred in the mines. The establishment of Mining Research Departments in both Thompson and Copper Cliff in 1979/80 provided the stimulus and the opportunity to dedicate personnel to specific research projects, on a full-time basis; and more particularly to concentrate on the efficiency of the total mine system. The Copper Cliff North Mine was brought back into production as a research mine dedicated to the development and testing of new mining methods and equipment. A shop facility known as the ""Murray Shop"" was set up as the centre of innovation for the development and construction of prototype equipment required for new bulk mining methods. While other expenditures were being curtailed , investment in mining research was being increased. An early success was an ""in the hole"" drill that could be broken down into small components to access even the remotest sectors of a mine to convert labour intensive undercut-and-fill mining to "" vertical crater retreat "" , VCR, bulk mining. A new mine manufacturing company-Continuous Mining Systems-was incorporated with the objective of building innovative mining equipment, with strong focus on automated continuous sytems. Continuous Mining Systems has now grown to a $25 million per year business. With these synergistic thrusts, the VCR bulk mining method was implemented in all of the Ontario Division mines, raising the percentage of ore mined by bulk mining techniques from 320/0 in 1981 to over 90% today. Coupled with re-designed job responsibilities, sup- ported by appropriate training, mine productivity more than doubled. It was discovered that VCR mining could be adapted to the narrower Thompson mine ores with identical improvements in productivity. Development of the Thompson Open Pit North had an additional favourable impact on production cost s in Thompson. The timing of the open pit development could not have been better. It permitted Inco to rapidly increase production at high productivity and low cost, in years of record demand and prices."

Jan 1, 1990

By E. Ghali, A. D. Bas

"Treatment of oxide, e.g. roasted gold ores has received relatively less attention than sulphidic ones and recently, handling of oxide gold ores is becoming essential. In this study, the influence of roasted gold ore (RGO) slurry (at 35% solid ratio) on the electrochemical dissolution of gold has been examined using pure gold and iron oxide minerals (hematite, maghemite, and magnetite) as disc electrodes by zeroresistance ammeter mode either in one or two separate containers. The presence of slurry decreased galvanic currents, i.e. corrosion rates due to the release and increase in the amount of soluble and/or insoluble species that could retard further the anodic behaviour of gold. Higher galvanic currents were observed in two separate containers than that in one container indicating the primary importance of galvanic interactions. SEM-EDS results observed iron hydro/oxide coatings on gold surface in case of magnetite test. However, in case of RGO electrode it is not a perfect passivation, although somewhat decreasing trend was observed on gold dissolution. These findings have revealed that passivation is not only a laboratory phenomenon, but also it could be responsible for the slowdown of gold dissolution in industrial practice.INTRODUCTIONThe cyanidation of gold is an electrochemical corrosion process thus; the dissolution of gold can readily be studied by electrochemical techniques. The outcome/summary of previous electrochemical studies of gold leaching can partially be summarized as follows: Filmer (1982) and Lorenzen and van Deventer (1992) reported the negative effect of pyrite and pyrrhotite on gold leaching due to the galvanic interactions; Aghamirian and Yen (2005) studied gold galvanic corrosion in presence of sulphide mineral electrodes only in one cell using zero-resistance ammeter (ZRA) mode; Dai and Jeffrey (2006) discussed, individually, the anodic and cathodic behaviours between gold and sulphide minerals, and Azizi et al. (2010, 2011) used a bed reactor to examine the influence of sulphide minerals on gold dissolution in one and two separate cells. Accordingly, a slowdown in gold leaching has generally been attributed to the passivation phenomenon, although the indication of this phenomenon is still an undergoing debate. For instance, Mrkusic and Paynter (1970) identified the passivation as the film formation of dissolved species from calcine, Crundwell (2014) explained passivation, in the case of corrosion of metals, as the formation of a layer which is composed of a new chemical species on the mineral surface and this layer inhibits or blocks further reactions. He suggested that passivation is related to the oxidation of the metal, and the formation of a passive layer that retards the rate of dissolution by several orders of magnitude. Consequently, galvanic interactions, passivation phenomena, or combinations of them are considered significant electrochemical factors affecting gold dissolution (van Deventer, Reuter, Lorenzen, & Hoff, 1990; Azizi, Petre, & Larachi, 2012). It is important to note that generalizing the influence of a mineral/metal on gold dissolution may not be correct in all cases, since contrary behaviours/results (for pyrite) have been observed (Aghamirian & Yen, 2005). Therefore, electrochemical study of gold is still a subject of research and need to be examined further in detail."

Jan 1, 2016

By Jean-François Lagueux

Goldex Mine is an underground mine located near the town of Val-d?Or. Multiple innovations were required to work the mine. Due to its low grade, it was important to develop a bulk mining method to economically extract the ore. The method used is a combination between shrinkage stope and long hole stope methods. This modified method is more than adequate; however, all of the ore body must be liberated in a very short period of time. Therefore, to bring down 23 Mt of ore, large blasts up to 1.8 Mt are required. The situation is also complicated by the fact that the mine is located in a residential area. No less that 40 homes are built within a 1 km radius of the mine site. In addition to managing mass blasts, means must also be found to mitigate the vibrations produced by the blast to diminish the disturbance to the neighbourhood. Therefore, the mine used numerous tools to properly plan the blasts and reduce the impacts. A program to control vibrations was one of the first tools put in place. The program?s aim was to draw up an action plan to ensure a follow-up of the homes by inspecting them. The second tool was to develop a methodology to plot a reliable attenuation curve and to adequately determine the maximum charge weight for each blast. Finally, effective communication tools were developed to inform the neighbourhood. There was a huge amount of work to be done in relation to the bench. Reliable and fast loading techniques had to be found to load large blasts in a very short period of time. Adequate logistics had to be developed to send underground more than 20,000 kg of emulsion per day while performing the daily underground operations. Technology had to be developed to adapt the i-kon electronic blasting system to large blasts. Since the control of the charges was the major factor to reduce surface vibrations, it was important to employ loading techniques that would permit decking of the long holes to reduce the charge weight per delays without adversely affecting the loading. All of this work requires an enormous amount of energy but also close collaboration between the different stakeholders (various departments, suppliers and consultants).

May 1, 2011

"During the development of a mining project, samples are taken for metallurgical tests. The sample can be a composite sample made of material from a selection of various drilling products or a sample from a mining test. It can be also a sample from an existing plant in order to investigate process enhancement. This sample can represent several tonnes of material which are analysed and split for use in various tests.In the current case, a composite sample of gold ore is dedicated to a Gravity Recoverable Gold (GRG) test. The properties of interest are the head grade of gold, the size distribution of the material, and the gold assay per size class. The sample is mixed and a sub-sample is taken to measure the feed properties: the size distribution and the gold assay per size class. The remaining material of the sample is subject to the first stage of concentration using the laboratory Knelson concentrator. The concentrate is analysed in its entirety, while a sample of tailings is taken for the same analysis. The remaining part of tailings is subject to grinding to feed the second and third stages of concentration. Similar sampling is performed at each stage.The concentration performances are analysed in light of these measurements. But what is the accuracy of the measurements of the properties? Pierre Gy’s theory of sampling, associated with a heterogeneity model built using the GRG test results, will answer this question. The current paper presents, through a real example, the calculation of the errors of the measurements of size distributions and gold assays performed during the GRG test. The uncertainties of the concentration performances and liberation size estimate are then calculated and their impact on the results interpretation is analysed.INTRODUCTIONDuring the development of a mining project, samples are taken for metallurgical tests. The sample can be a composite sample made of material from a selection of various drilling products or a sample from a mining test. It can be also a sample from an existing plant in order to investigate process enhancement. This sample can represent several tonnes of material which are analysed and split for use in various tests. In order to take into account the ore variability in terms of gold deportment, it is recommended to perform variability studies (Dominy, Murphy, & Gray, 2011, Dominy, 2013) using samples representing various horizons or years of production. But for a given ore type, it can be necessary to perform many tests with different operating conditions (Brochot, 2012a). In that case, the variability of the samples is not due to the heterogeneity of distribution of the deposit, but only on the heterogeneity of constitution of the primary sample. The level of confidence associated with the estimated performance parameters is directly linked to the overall error (including sampling error) of the measurements performed during the tests."

Jan 1, 2016

By M. Niinae, F. Nakahara, T. Kawamoto, T. Suzuki

"MgO was chosen as an As(V) immobilization agent and a series of immobilization experiments was performed to obtain insights into the behavior of As(V)and MgO during leaching tests. Our experimental and modeling results demonstrated that As(V) immobilization by MgO consists of the following steps: (i) an increase in soil pH, (ii) desorption of As(V) from the soil, and (iii) the re-immobilization of As(V) by MgO/Mg(OH)2particles. Regarding the behavior of MgO, the modeling results showed that when the MgO dosage was 25 mg MgO/4g-drysoilor less, the majority of MgO was used to increase pH, and less than 1% of MgO was used to sorb As(V), which was consistent with the result of leaching tests showing that a high level of As(V) was leached at the MgO dosages. On the other hand, when the MgO dosage was above 25 mg MgO/4g-drysoil, the percentage of MgO used for As(V) sorption increased up to 35%, and correspondingly, the As(V) leaching level decreased to below 0.01 mg/L As at an MgO dosage of 75 mg MgO/4g-drysoil.INTRODUCTIONOne of the major arsenictransport pathways causing groundwater contamination is the migration of arsenicfrom artificially or naturally contaminated surface soils. Therefore, in this study, the immobilization of arsenate (As(V)) in soils was investigated using commercial-grade magnesium oxide (MgO) as a potential cost-effective measurement to prevent the occurrence of groundwater contamination with As(V).With respect to the use of MgO for remediating As(V)-contaminated environmental media, MgO has been used in several studies to remove As(V)from water(Yu et al., 2011; Jia et al., 2013; Tresintsi et al., 2014). Although these studies demonstrated that MgO effectively removes As(V)from water in water-MgO systems, the behavior of As(V)in water-soil-MgO systems has not been evaluated in the literature, and several questions remain to be answered to comprehensively evaluate the use of MgO as an As(V)immobilization agent in soils. For instance, as is well-documented in the literature(Ford, 2002; Lenoble et al., 2002; Zeng et al., 2008), iron oxy-hydroxide (FeOOH) is one of the important adsorbents for As(V)in both natural and engineered systems. Therefore, it is anticipated that the presence of FeOOH in soils significantly influences the behavior of As(V)in water-soil-MgO systems and consequently, the immobilization performance by MgO. Another important question to be answered is the behavior of MgO during the immobilization treatment. In our previous studies (Suzuki et al., 2013a, 2013b), commercial-grade MgO was demonstrated to effectively immobilize lead and fluoride, and the immobilization effectiveness was indicated to last more than 100 years against the exposure to acid rain (pH 4.0). However, these studies focused on the behavior of contaminants. As a result, little is known about the behavior of MgO during the immobilization treatment. The lack of understanding of the behavior of MgO during immobilization treatment is one of the main concerns leading to relatively low public confidence in immobilization treatments."

Jan 1, 2016

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

Mechanical cells are the dominant unit operation in base metal rougher flotation applications worldwide. As economic demand for metals increase, and as the feasible ore grades decrease, the installed capacity of mechanical flotation units worldwide has greatly increased, adding significant capital and operating cost. One opportunity to improve this situation is to use fundamental knowledge to make the industrial flotation process more efficient. A key insight is that the flotation process can be optimized by dividing the processing of the pulp into two distinct fluidic environments; a collection cell where air and pulp mix in a highly energetic compartment, and a low energy gravity separation cell. This is the basis of the “2-stage flotation device” that has been patented by Eriez and marketed under the tradename StackCell™. Two industrial case studies will be presented to highlight the potential metallurgical and commercial advantages of the StackCell. In the first, a train consisting of three pilot scale StackCells (each 0.6 metre diameter) was run in parallel with a train of conventional mechanical cells in a major copper concentrator and benchmarked against a batch Denver test. In the second, the performance of a 3 metre diameter StackCell in a nickel sulfide cleaning application was benchmarked against a batch Denver test. The flotation kinetics observed in the StackCells were 2.4 to 2.9 times faster when compared with the lab Denver test. In the first case study, the kinetics of the StackCell was about six times faster than conventional mechanical cells. It is hypothesized that the efficiency improvement in kinetics is because of reduced “drop-back” during the froth recovery phase, which is reduced because of the reduction in the shear that is present in a conventional mechanical tank cell. This result suggests that a 2-stage flotation system can be used to reduce the working volume of flotation units by five to six times. This would allow operators to significantly reduce the size of their flotation lines for a comparable flotation objective. Some discussion will follow about the scale-up criteria, maintenance and operability issues, comparisons of layout, and possible cost savings. INTRODUCTION The solution that most equipment suppliers have offered to reduce the cost of mechanical flotation equipment over the last 30 years has been to increase the size of the flotation unit cell (Mankosa, 2017). For example, three 100 m3 machines with three smaller motors, mechanisms and control valves could be replaced with a geometrically similar 300 m3 with a larger version of each ancillary component. While there are some efficiencies and cost improvements to be gained by this approach, it does not take advantage of fundamental insights into the flotation process, and it does not represent a step-change in terms of process design or economic benefit for the customer. The advantage of the “bigger is better” approach is that there is past experience of incremental improvements by making cells larger and there is a well-established scale-up method, which allows practitioners to estimate equipment capital costs based on very simple laboratory test- work. A different paradigm, which is gaining interest rapidly is to use a unit operation in which it is easier to optimize the flotation process, which will be described in this paper. This approach is often referred to “2- stage flotation”. Because this approach is new, it is important to benchmark it against existing conventional technology and to demonstrate a robust methodology for sizing equipment and flowsheets. This paper aims to address these issues with two case studies.

Jan 1, 2019

By A. B. Szwilski

"The various elements of longwall mine design are reviewed, particularly as they apply to condition s in the coalfields of the Plains region of Alberta. The following recommendations are made: (1) optimum panel lengths of 1000 to 2000 metres; (2) face lengths of 100 to 150 metres; (3) the use of retreat methods; (4) the restriction of floor undulations to less than 15 degrees. IntroductionThis article considers the elements of longwall mine design and is part of a short study that has been made of the feasibility of longwall mining in the Plains region, Alberta, Canada. The Plains region, with about 20 million tonnes, accounts for more than 80% of the coal resources of Alberta. The coalbearing strata in the Plains region cover an area of about 95,000 square miles. The geological formations dip very gently toward the southwest and, from a mining point of view, may be considered horizontal. The depth of the coal seams varies from several metres to over 650 metres. However, it is not anticipated that underground mining will be extended beyond the depth range of 25 to 250 metre s.The coal seams in the Plains region have been subject to many small mining operations in the past. The majority of these small mines were limited to a depth of about 50 metres, working mainly by room-and-pillar methods, although hand-filled longwall mining was often used.The development of new underground mining operations in the Plains region will have to be on a &apos;start-from-scratch&apos; basis, especially regarding the installation of equipment. Longwall mining is more mechanized than any other method of mining and requires careful attention for the selection of new and expensive equipment. There have been many proven longwall advance and retreat systems in Europe and North America; mining conditions in the Plains region are more similar to those in the United States. Therefore, it will be essential for new mining pro jects in the Plains region to commence with proven system s and equipment; in time, perhaps an experimentation process could be attempted as experience is gained.Generally, the coal face is the key factor in mine productivity, which depends a great deal on equipment reliability, training and experience of the operators. The original design of coal faces must be flexible enough to cater for potential improvements so that running time can be increased."

Jan 1, 1980