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A series of fault slip seismic events have occurred at Garson mine. The events have occurred during development advancement exposing faults, after the extraction of single stopes due to mining induced stress change, and due to time effects after delays of many days during which no mining had been conducted. This paper discusses the study completed consisting of a global or holistic approach aimed at mitigating some of the risks associated with the structurally controlled rockbursts at Garson. Background Garson mine is located in the south east area of the Sudbury Basin (Ontario, Canada). Mining occurs in two main ore bodies (the #1 Shear and the #4 Shear) which are copper-nickel-sulphide and occur within shear zones that (as well as the ore bodies) strike approximately east-west and dip steeply south. From the north (footwall), lithologies progress (Figure 1a) from norite (NR), to greenstone (GS), and finally to meta-sediment (MTSD). A generally WNW-ESE striking, bifurcated sub-vertical dyke of olivine diabase (OLDI) cuts the series. The current active mining block at Garson is between 4700L and 5100L (levels denote depth in feet below surface).

May 1, 2009

By Paul R. Clark

Outline of Presentation ? Brief review of gasification and IGCC ? Polygenerationand its potential ? Why gasification today ? Major technology providers and their activities ? US initiatives ? Alberta Energy Resources and application of gasification ? CO2 management ? Advantages and barriers to gasification ? Conclusions

May 1, 2006

By K. G. Thorne, D. R. Lentz, A. F. Park

"The Main zone of the Clarence Stream gold deposit is hosted by a shear zone adjacent to the Saint George batholith in southwestern New Brunswick. The mineralization history relates to progressive deformation in the shear zone and the intrusion of the Magaguadavic granodiorite.Of three phases of deformation that have affected the area (D1-D3), D1 is not observed in the Clarence Stream shear zone. Two suites of igneous rocks intrude the shear zone: post-D2, pre-D3 gabbro; and granitic dikes, which crosscut the gabbros. Granitic dikes show a diverse range of relationships to D3 features, from pre-syn-D3 mylonitic fabrics in transposed and highly deformed enclaves, to nondeformed and entirely discordant dikes (post-D3).A variety of veins occur within this shear zone. Zoisite-bearing and later, crosscutting epidote-bearing veins are preserved as relics in quartz-mylonites, and predate quartz veins that host gold mineralization. The epidote-bearing veins are associated with pseudotachylitic cataclasite dikes, consistent with transient brittle failure during ductile deformation.Deformation relates to right-lateral strike-slip motion across the shear zone, and subhorizontal and near-vertical stretching lineations imply transpression. A late (late- or post-D3) dip-slip phase of motion is also apparent, and cooling followed a metamorphic peak early in D2-D3 history. Deformation of the shear zone, and more localized transient brittle failure, provided the permeability for fluid movement and vein emplacement. This scenario is similar to the findings of other studies, in which geochemical, tracer isotopic, and age-dating (U/Pb, 40Ar/39Ar) results suggest that the adjacent Magaguadavic granodiorite provided heat and metals. La zone principale (Main zone) du gisement d’or de Clarence Stream est encaissée par une zone de cisaillement qui longe le batholite de Saint George dans le sud-ouest du Nouveau Brunswick. L’historique de la mise en place de la minéralisation accorde un rôle important à la déformation progressive dans la zone de cisaillement et à la mise en place de la granodiorite de Magaguadavic.Des trois phases de déformation à avoir affecté la région (D1-D3), les effets de D1 ne sont pas observés dans la zone de cisaillement de Clarence Stream. Deux suites de roches ignées se sont mises en place dans cette zone de cisaillement : un gabbro post D2, pré-D3 et des dykes granitiques qui recoupent le gabbro. Les dykes granitiques montrent une grande diversité de relations avec les structures D3, allant de précoces à synchrones aux fabriques mylonitiques D3 dans des enclaves transportées et fortement déformées à des dykes complètement discordants et non déformés (post-D3).On note diverses veines dans cette zone de cisaillement. Des veines contenant de la zoisite ainsi que des veines d’épidote sécantes plus tardives, lesquelles sont préservées comme reliquats dans les mylonites à quartz et sont plus vieilles que les veines de quartz contenant la minéralisation aurifère. Les veines d’épidote sont associées à des dykes de catacalasite à texture de pseudotachylite, compatibles avec une rupture cassante épisodique durant la déformation ductile.La déformation résulte d’un mouvement décrochant dextre le long de la zone de cisaillement et les linéations d’étirement subhorizontales et quasi-verticales impliquent de la transpression. Une phase tardive (tardi- ou post-D3) de mouvement selon le pendage est aussi apparente, et le refroidissement succède à un pic métamorphique tôt dans l’histoire D2-D3. La déformation de la zone de cisaillement et la rupture cassante épisodique ont fourni la perméabilité nécessaire au mouvement des fluides et à la mise en place des veines. Ce scénario est similaire à celui proposé par d’autres études, dans lesquelles les résultats de géochimie, de traceurs isotopiques et de géochronologie (U/Pb, 40Ar/39Ar) suggèrent que la granodiorite de Magaguadavic avoisinante a fourni la chaleur et les métaux."

Jan 1, 2008

By T. R. Wildeman

"At the Big Five wetland, which was built to treat drainage from a metal mine, encouraging results have attributed to sulfate reduction. Cu and Zn were reduced to below detection limits of 0.50 mg/L, Fe was reduced by at least 90% to below 3 mg/L, and the pH was raised from below 3 to above 6. The study showed that treatment was primarily through microbial processes within the substrate and that plants were not needed to insure success.Constructed wetlands can be taken through design and proofing stages, and into operations scale in a meaningful time frame for on-line acid drainage remediation. In staged design, laboratory experiments are initiated to determine reasonable substrate materials for remediation of a particular drainage. When a reasonable substrate is found, bench scale experiments are initiated as the next stage in the design. In these studies, removal efficiency, permeability, and flow are analyzed to identify candidate substrate which will support sulfate reduction and be hydraulically reasonable. Using this design method at one site, an acid drainage of pH 2.5 was raised to a pH of above 7. Total heavy metal concentrations in this drainage exceeded 900 mg/L. In the pilot constructed wetland Cu, Zn, and Fe were reduced by 99%.Determining how well these anaerobic cells operate in the winter is done at the pilot plant site in Idaho Springs, CO. It's been found that over the last four winters, those cells that emphasize anaerobic removal within the substrate have operated continuously and effectively the year round. Winter operation is aided by the fact that the water temperature of the acid drainage is always between 14 and 16° C, and the cell can be readily insulated from the cold. Activity of the sulfate reducing bacteria may decrease during the winter. However, it appears that organic complexation of metals helps to maintain low concentrations of metals in the effluent."

Jan 1, 1992

By K. G. Thorne, L. J. Cabri, L. R. Fyffe, D. R. Lentz, D. Hoy

"Mineralization at the Main zone of the Clarence Stream gold deposit in southwestern New Brunswick occurs within a parallel series of quartz veins that occupy a NE-trending, steeply north-dipping, brittle-ductile shear zone related to the nearby terrane-bounding Sawyer Brook fault. This deformation zone cuts volcaniclastic sedimentary and volcanic rocks of the Silurian Waweig Formation, and is intruded by mantle-derived East Branch Brook gabbroic dikes and a number of felsic dikes. Intermittent shear-zone reactivation during local magmatic activity induced heterogeneous deformation and alteration of the country rocks and dikes. The terminal phase of shearing is associated with narrow pegmatite-aplite dikes that grade laterally into granophyric granite, and into an auriferous quartz vein. Gold, aurostibite, and gold-antimony intergrowths, predominantly with vein quartz, are associated with a low ƒO2 sulfide mineral assemblage (pyrrhotite, arsenopyrite, and berthierite) that has a sulfur isotopic signature consistent with a magmatic source. Gold saturation and deposition occurred episodically between 300° and 360°C at low activity of sulfur, predominantly in response to pressure reduction during brittle failure combined with the effects of decreasing temperature and sulfidation reactions during fluid-wall-rock interaction. Geochronological studies demonstrate that the timing of the emplacement of the pegmatite-aplite dikes and associated auriferous veins overlap with that of the Early Devonian Magaguadavic granite. Moreover, the composition of the dikes is consistent with their derivation from this granite as late fractionates. The Main zone of the Clarence Stream deposit is therefore interpreted to be intrusion-related, with the Magaguadavic granite being the ultimate source of the gold-bearing fluids. La minéralisation de la zone principale du gisement aurifère de Clarence Stream dans le sud-ouest du Nouveau Brunswick occupe une série de veines de quartz parallèles dans une zone de cisaillement fragile-ductile de direction NE et à pendage abrupt vers le nord apparentée à la faille de Sawyer Brook avoisinante qui délimite deux terranes. Cette zone de déformation recoupe les roches volcaniclastiques et volcaniques de la Formation de Waweig d’âge Silurien et est recoupée par les dykes de gabbro d’East Branch Brook, d’origine mantellique, ainsi que par divers dykes felsiques. La réactivation intermittente de la zone de cisaillement durant l’activité magmatique locale a entraîné la déformation et l’altération hétérogènes des dykes et de l’encaissant. La phase finale du cisaillement est associée à de minces dykes d’aplite-pegmatite qui passent latéralement à des granites granophyriques, puis à des veines de quartz aurifères. L’or, l’aurostibite et les intercroissances d’or et d’antimoine principalement associées aux veines de quartz sont associés à un assemblage minéral de sulfures à faible ƒO2 (pyrrhotite, arsénopyrite et berthierite) présentant une signature isotopique du souffre compatible avec une source magmatique. La saturation en or et la précipitation a eu lieu épisodiquement entre 300° et 350°C dans des conditions de faible activité de souffre, principalement en réponse à la baisse de pression durant la rupture fragile ainsi qu’aux effets de la baisse de température et aux réactions de sulfuration durant l’interaction entre le fluide et l’encaissant. Les études géochronologiques indiquent que l’âge de la mise en place des dykes de pegmatite-aplite et des veines aurifères qui leur sont associées coïncide avec le granite de Magaguadavic, d’âge Dévonien Précoce. De plus, la composition des dykes est compatible avec une origine provenant de ce granite sous la forme de phases fractionnées tardives. Le gisement principale (Main zone) de Clarence Stream est ainsi considéré comme étant associé aux intrusions, le granite de Magaguadavic étant la source ultime des fluides aurifères."

Jan 1, 2008

By Kyle Selga, Christopher J. Greet

"The Pulp Chemistry Monitor (PCM) has been installed and operated in a number of concentrators in Australasia, and has produced some very interesting results. PCM measures the pH, Eh, dissolved oxygen, temperature and oxygen demand of process streams within a concentrator continuously and in real time.In the example presented in the paper the cleaner feed pulp chemistry is measured, and related back to changes in the mineralogy as well as variations in metallurgical response. For example, a decrease in Eh and dissolved oxygen as well as an increase in the oxygen demand. These changes in the pulp chemistry appear to coincide with increases in the feed grades, probably higher pyrite grades. Further, these variations in pulp chemistry and feed mineralogy have a negative impact on the concentrate grade and recovery.The implication is that if the pulp chemistry is monitored it is possible to predict changes in the mineralogy of the system, and from this the metallurgical response. Further, the changes in pulp chemistry can be used in a control system to counter the changes in mineralogy and maintain the metallurgy.IntroductionIt has long been recognized at a fundamental level that the flotation behaviour of sulphide minerals is dependent on the pH and Eh of the system (for example: Tolun and Kitchener, 1964; Woods, 1976). However, measuring pulp chemical parameters within an operating concentrator has had limited attention. Woodcock and Jones (1970a and b) were among the first to actually collect data from operating plants. Jones and Woodcock (1984) identified some of the chemical complexities of flotation pulps, and noted that most concentrators already operate at high levels of metallurgical performance but there was still considerable scope for improvement. These gains would, in part come from more effective chemical control. In the same volume, Healy (1984) discussed metal ion hydrolysis and precipitation, redox chemistry and how the surface chemistry of sulphide minerals can be influence flotation behaviour. The impact of grinding environment on the pulp chemistry and flotation response of a complex polymetallic sulphide ore is discussed by Forssberg and Subrahmanyam (1993). Their work essentially contrasts the pulp chemical and flotation performance of pulps ground with conventional ferrous grinding media to a fully autogenous system. The data suggested that the reduction or removal of iron hydrolysis contamination from the system could have a positive influence on flotation behaviour, and perhaps the first stage of “controlling” the chemistry of the system should be in the grinding circuit. Grano et al (1995) demonstrated that the collection of pulp, solution and surface chemical data could identify weaknesses within a concentrator, and from these data develop, test and implement solutions that improved the process. Greet et al (2006) showed that the pulp chemical trends observed through primary grinding and rougher flotation circuits of most sulphide flotation concentrators are approximately the same. There are differences in the magnitude of various pulp chemical parameters (pH, Eh, dissolved oxygen, temperature and oxygen demand) as a result of variations in the feed mineralogy and processes employed. Further, it was noted the Eh shifted to more reducing values of the ten days measurements were collected, and this change could be traced to an increase in the iron sulphide content of the feed. Metallurgically, this resulted in an increase in nickel recovery through the recovery of nickeliferous iron sulphides, and a decrease in nickel concentrate grade."

Jan 1, 2016

By A. G. Doll

"Multi-stage crushing plants, including high pressure grinding roll (HPGR) circuits, require more materials handling equipment than SAG and ball mill circuits. A temptation exists to neglect the energy consumption of materials handling when doing desktop comparisons of HPGR versus SAG milling circuits because estimation of materials handling power requirements can demand significant general arrangement drafting that is not available when performing preliminary scoping or ""desktop"" studies.The largest component of materials handling power is consumed by conveyors when lifting material between stages of crushing, meaning that a simple potential energy model can be used to evaluate conveying specific power consumption. This potential energy consumption for conveying can then be factored to provide overall materials handling specific energy consumption suitable for desktop studies of HPGR and other multi-stage crushing circuits. The technique also permits different crushing circuit flowsheets to be evaluated, at a preliminary level, by simply counting the number of times a conveyor must ""lift"" ore from ground level up to a bin or other equipment mounted up high.INTRODUCTIONTrade-off studies of multi-stage crushing plants involving high pressure grinding rolls (HPGRs) versus semi-autogenous grinding (SAG) mill circuits typically focus on the operating cost savings of crushing plants versus the capital cost savings of SAG mill plants. Very high-level (or desktop) studies of this type do not include a great deal of engineering work, meaning that the details (and costs) of ancillary systems are often not captured when doing the comparisons. The more complicated nature of the multistage crushing plants means there are more ancillary systems than a SAG mill plant, risking a skewed analysis of the operating costs, benefiting the HPGR option.One of the ancillary systems that may be overlooked is the more complicated conveying systems inherent in multi-stage crushing plants. Due to the mass of material being handled in large mining operations, these conveyors become a significant portion of the operating costs. A simple way to estimate the specific energy consumption of conveying systems without requiring engineering drawings and a site layout would allow high-level studies to capture this operating cost."

Jan 1, 2015

By R. J. C. MacDonald

"The first recorded discovery of placer gold in Canada was on the Chaudiere River in Quebec's eastern townships in 1832 and the first major production came from placer deposits in British Columbia in 1858 (Verity 1974). Placer deposits were the major source of gold until the late 1890's, although lode deposits were found in Nova Scotia in 1862. In 1896 the Klondike Gold Rush commenced with the discovery of gold by Henderson and Carmack in the Yukon. The peak was reached in 1900 when over one million ounces of gold valued at approximately $22.3 million were recovered. With the exhaustion of the easily accessible deposits, more capital intensive operations were necessary such as those initiated in 1906 by the Yukon Gold Corporation. This giant dredging operation continued to produce until 1966. Today only a few small operations remain.The development of Canadian lode gold mining began in earnest around 1900 when the railroads were opening up the interior and western part of the country. At this time many discoveries occurred, particularly where mineral outcroppings were near surface. Tremendous impetus for exploration and the reopening of former producers followed the elimination of the gold standard by Great Britain around 1931, and more importantly the United States in 1933, with subsequent stabilization of the price of gold at $35 per ounce in 1934. Prior to this the price of gold was basically unchanged since 1717, approximately, $20 per ounce (Verity, 1974; Green, 1973).The depression of the 1930s adversely affected most industries, however for the gold producers it was a very profitable time. Prices of materials and equipment declined and labour was cheap and plentiful. Interest in gold mining was further stimulated in 1936 when a three year tax exemption was declared for all new metal mines. In 1938 a decline in the value of the Canadian dollar relative to the U.S. dollar benefitted the industry. The second world war provided yet again further incentive for gold production and in 1941 Canadian gold production reached a high of approximately 5.4 million ounces. This figure was not surpassed until 1991 when approximately 5.6 million ounces were produced (Natural Resources Canada, 1996 )."

Jan 1, 1998

By C. Gosselin, G. M. Payenzo, C. Bazin, M. Desnoyers, G. Chevalier

"ArcelorMittal Mines Canada operates hydraulic classifiers to reduce the silica content of an iron oxide concentrate from 4.5% to less than 1.2% SiO2. Sampling campaigns were conducted to develop a preliminary model for the hydraulic classifier. Results show that iron oxide and quartz are respectively concentrated in the coarse and fine size fractions of the classifier feed. The mineral partition curves in the classifier and the size distribution of the minerals in the feed are used to simulate the classifier operation in order to identify the source of variability observed in the daily plant production data. Results indicate that most of the variation is probably caused by unsteady operating conditions of the classifiers.INTRODUCTION Hydraulic classifiers are a class of gravity separation devices (Wills, 1992) used to classify particles according to their size and specific gravity. Hydraulic classifiers also known as density separators (Hearn, 2002) combine hindered settling and upward flow of water to selectively entrain light particles to an overflow stream. Denser particles settle to an underflow stream. Hydraulic classifiers are used for various purposes ranging from cleaning of coal (Das, Sarkar, & Mehrotra, 2009), recovery of chromite from plant tailings (Kumar, Tripathy, & Rao, 2009) to upgrading iron oxide concentrate (Hearn, 2002). ArcelorMittal Mines Canada operates several hydraulic classifiers to upgrade an iron oxide concentrate by removing fine quartz and other light minerals from hematite that is the main iron oxide mineral. A project was initiated to analyze the operation of the classifier in an attempt to identify the source of the variability in the production results and to assess the potential of increasing iron recovery in the classifier. One possible disturbance for the classifier operation is the change in the classifier feed composition and size distribution, while the other considered disturbance is the operation of the classifier. The paper is divided into five sections. The first section summarizes the ore processing sequence for the production of iron oxide pellets to identify the position of the hydraulic classifier in the process. The second section describes the hydraulic classifier. The third section discusses the sampling campaigns conducted on the hydraulic classifier. The method used to process the data from sampling, the results of the data processing and simulation of the classifier are discussed in the remaining sections."

Jan 1, 2012

By Carl Bertrand, Patrice Nadeau, Claude Bazin

"Rio Tinto Fer et Titane processes the ore extracted from the Lac Tio orebody to concentrate the hemo-ilmenite (FeO.TiO2-Fe2O3) mineral that is subsequently smelted to produce pig iron and titanium oxide slag. The ore is crushed, heat treated (roasted) and concentrated using dry magnetic separators. This paper describes a model to predict the magnetic susceptibility of the roasted ore and its behaviour in the magnetic separation plant. The model uses a classification of the ore as a function of the size and magnetic susceptibility of the particles. A subsequent division into mineral species allows the simulation of the magnetic separation process. The model is used to assess various operating options when process disturbances occur.INTRODUCTIONOre processing operations consist of ore size reduction followed by the recovery of valuable species in a concentrate that is subsequently sold for its minerals or metal content. The concentration processes are based on the physical and/or chemical properties of the minerals contained in the ore. Flotation is typically used to recover minerals of base metals such as copper, zinc, lead or nickel. Iron ore and other dense minerals are typically concentrated by gravity separation processes. Magnetic separation uses the magnetic properties of the minerals to concentrate magnetite and other minerals, to clean coal concentrate and to recycle ferrosilicon for dense separation processes (Brophy, 1985).Magnetic separators can be divided into low and high intensity separators and high gradient magnetic separators (Gillet, 2008). Most of the low gradient magnetic separators are built as rotary drums that are operated in wet or dry conditions. Dry magnetic drum separators are used by Rio Tinto Fer et Titane in Sorel-Tracy, Canada to concentrate hemo-ilmenite from the ore extracted at the Lac Tio Lake mine near Havre St-Pierre, Canada. The analysis of production data suggests that process improvement could be achieved by identifying key variables for the control of the magnetic separation circuit and/or by modification to the circuit flowsheet. The use of a process model was considered to be an effective approach to conduct a preliminary analysis of the circuit with a view toward process optimization. However, the literature contains few detailed studies dealing with the modelling of magnetic separation processes."

Jan 1, 2011

By Jacques Ouellet, Ferri Hassani

"Within the mining industry paste backfill technology has evolved over the last two decades. This technology aims at producing safe and cost effective backfill of high strength. When used as cemented backfill underground the usual practice in the mining operations was to consider the backfill as a component under the production engineer responsibility. With its growing use in various site conditions some problems were encountered. Over time the mechanical characteristics of paste backfill could deteriorate. Investigations on these problems brought an increased understanding of paste backfill chemistry. Effectively, paste backfill is produced with tailings slurry having sometimes very complex chemistry and mineralogy. The bonding agents can be sensitive to the chemistry of these slurries. Today, a new understanding of the complex chemical reactions involved in cementing paste backfill is derived from concrete research literature. Cemented paste backfill is a manufactured product and as thus all steps involved in its making must be looked at. Both the processing engineers and mining engineers need to be involved. The proposed paper will go over the various developments and new understanding gained over the years on paste backfill chemistry. The authors had the opportunity to work with various Canadian mines that implemented paste backfill in their operations. Chemical alterations case study will serve to illustrate the importance of a good understanding of cement and paste chemistry. The authors will then conclude on research orientations needed to bring a more efficient use of paste backfill technology. The crucial role of the processing engineers in the paste backfill manufacturing will be discussed.INTRODUCTIONPaste backfill is usually prepared from a mixture of solids and water containing at least 15% particles smaller than 20 microns with the addition of binders for mechanical strength. Historically, tailings had been classified to extract the fines and only the coarse part was placed underground as sand slurry. Paste fill was researched in the last three decades to enable mining operations to use all the tailings, including fines, in backfill. The outcome of the early studies showed that paste fill provides high strength but due to preparation complexities and transportation problems is expensive. A greater understanding of the transportation mechanisms and advances in dewatering technology in recent years has resulted in more practical use of paste backfill."

Jan 1, 2002

By G. Couture

"The Glencore Raglan mine Concentrator, located in Quebec’s Nunavik region, has been in operation since 1997. The nickel metallurgical plant was originally designed to process 100 tonnes per hour of ore to produce 130 000 tonnes of nickel concentrate per year at 16% Ni. The increased grinding capacity and the discovery of rich ore bodies have increased the annual production to more than 190 000 tonnes of concentrate at 19% Ni. In 2014, the Metallurgical group, with support from BBA, implemented an expert control system on the flotation circuit. This project was initiated to cope with the high variability of the concentrator’s feed. Increasing the plant recovery and final concentrate grade were also important project objectives. The implementation of the expert system ensures better control of concentrate inventory while maintaining maximum plant throughput. This paper describes the successful implementation of the new control system, the control technology selection and the flotation circuit optimization. It will also discuss the performances obtained with this new approach and the improved process stability.IntroductionThe Raglan Mine is located in Northern Quebec along the 62th parallel. It has been in operation since 1997 and was originally designed to produce 130 000 tonnes of concentrate annually at 16% Ni. The increasing of the grinding capacity and the discovery of rich ore bodies have increased the annual production to more than 190 000 tonnes of concentrate between 18 and 20% Ni."

Jan 1, 2015

By Tassos Grammatikopoulos, Sarah Prout, William Mercer, C. Gunning

"The Nechalacho rare metal deposit is located in NWT, Canada. Avalon Rare Metals Inc. is now evaluating processing techniques for recovery of REE to be used in the industry. QEMSCANTM analysis coupled with whole rock, REE, Zr and Nb, and electron microprobe analyses were utilized to identify and quantify the mineralogy of the deposit, and define the REE distribution among minerals. QEMSCANTM mineralogy was performed on fifty intact drill samples using EXPLOMINTM that provided textural information for the mineralization; crushed samples were used to determine the variability of the ore and define REE domains, and composite samples to determine the liberation and association of the REE minerals. The sample matrix predominantly consists of biotite, feldspars, quartz, muscovite/clays, minor chlorite and amphibole, magnetite and hematite and carbonates. REE include in decreasing order of abundance allanite (3.6%), monazite (1.5%), synchysite (0.9%), columbite (0.9%), fergusonite (0.6%), bastnaesite (0.4%), and zircon 11.0%. The majority of the HREE are carried by fergusonite and zircon, and the LREE by monazite, allanite, synchysite and bastnaesite. QEMSCANTM mineralogical data is extremely valuable in assessing the mineralogy of the ore. The data have been implemented in defining the mineralogy of the ore body, support metallurgical test work, predict recoveries and grades, and finally calculate reserves of the deposit.INTRODUCTIONAvalon is in the unique position of owning and developing one of the world’s largest rare earth deposits in the Northwest Territories of Canada.The term Rare Earth Elements (REE) is used to describe the 15 lanthanide elements or metals and, more recently, the metals yttrium and scandium, which display similar physical and chemical properties. REE are not particularly rare and have a similar abundance in the earth’s crust to nickel, tin or tungsten. Unlike these better known metals however, the rare earths have little tendency to become concentrated into commercial ore deposits. When they do become concentrated, deposits of REEs usually display enrichment in one of two groups. The most commonly occurring deposits of REEs contain a sub-group of elements referred to as the “light rare earth elements” (LREEs), which include lanthanum, cerium, praseodymium, neodymium, and samarium. Less commonly, deposits of REEs contain the much more valuable sub-group of elements referred to as the “heavy rare earth elements” (HREEs), which include europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium and lutetium. Yttrium and scandium most commonly occur in association with this second HREE sub-group. Generally, LREEs are produced from the mineral bastnaesite, and the HREEs have been produced from the minerals xenotime and monazite. LREEs continue to be produced from bastnaesite, however in the 1980s, China started production of HREEs from unique laterite clay deposits. China now produces approximately 95% of the world’s rare earths."

Jan 1, 2011

A comprehensive axial test program has been undertaken at the WASM Dynamic Test Facility on the performance of 20mm threadbar. The threadbar tests had a double embedment configuration, with variation to the encapsulation lengths, debonded length and surface hardware and fixtures to examine failure mechanisms. Performance was significantly affected by the collar embedment length and type of fixture nut used. Results ranged from thread shearing occurring between the nut and bar at 1kJ of energy dissipated, to control plastic deformation of a central debonded length dissipating 22kJ.

May 1, 2009

By Nick D. Rose

The importance of structural geology in open pit slope stability is overlooked in many mining operations due to pressures to cut costs and minimize mine personnel. As a result, mine geology staff are often unable to keep up with geological mapping of the pit slopes due to other day to day production oriented tasks. In some instances, key structural controls that contribute to slope instability may not be identified until after a slope instability has initiated, at which time it may be difficult or impossible to delineate these features in the field. Even for small slopes, understanding the engineering and structural geology of the rock mass, as well as the effects of groundwater and blasting practices, are of key importance to slope stability. Two case histories are presented as examples of mining operations that experienced slope instability with little more than anecdotal information about the structural and groundwater conditions encountered during mining. By conducting field reconnaissance, simple slope stability analyses and ongoing slope monitoring, practical remedial measures were implemented, allowing continued safe mining operation.

May 1, 2002

By Mehmet Canbazoglu, Kaarlo Haavanlammi, Hannu Laitala

Cobalt is currently a critical metal because of the increasing number of electrical vehicles forecast for production. Copper and zinc are base metals that are used in many everyday applications, including vehicle production. Eti Bakir Co. Mazidagi Phosphate Plants in the Mazidagi district of Mardin, Turkey is a multi-product facility that produces cobalt, copper, and zinc, as well as fertilizers, from the roaster calcine. Roaster off gas is treated in a sulfuric acid plant (SAP) and the acid itself is utilized in fertilizer production. The comminution circuit, roaster, SAP, and calcine leaching process were developed by Outotec and integrated with the fertilizer production process implemented by Eti Bakir A.Ş. This paper describes the hydrometallurgical calcine leaching process used at the plant, with a focus on the oxidative pressure leaching of the cobalt, zinc, and copper-containing calcine. The cold and hot commissioning phases of the pressure leaching autoclave start-up are discussed together with experiences from the ramp-up period. The production processes of the copper cathode and cobalt and zinc-containing salts are also presented together with some of the plant’s operating parameters. INTRODUCTION The ETI Bakir A.S. fertilizer plant, Turkey, is located in Mardin Province near the border with Syria, (Figure 1). The leaching plant project was started in 2012. The purpose of the plant is to separate the valuable metals from the calcine produced by the pyrite roasters, which are used to produce a SO2 containing gas for the new sulfuric acid plant. Fertilizer production requires sulfuric acid, and for this reason a sulfuric acid plant was built; customer already had access to pyrite concentrate, which is used as the roaster feed. During the engineering phase of the roaster and acid plants it was noticed that the roasted calcine contained considerable amounts of valuable metals like cobalt, copper, and zinc. To recover these metals, a calcine leaching process was developed at the Outotec Research Center in Pori, Finland. The main process connections of the Mazidagi site are shown in Figure 2.

Jan 1, 2019

By Patrick R. Stephenson

The current version of the JORC Code was issued in 2004 and since that date the Australian Securities Exchange, in collaboration with JORC, has issued a number of companies? updates to provide guidance on various reporting issues. An interim review will be undertaken during 2010 to meld these updates with the current 2004 Code. This will be followed by a full review during 2010/11 covering all aspects of the Code and focusing particular attention on: ? clauses / guidelines relating to Exploration Results1, exploration targets and the reporting of historical estimates; ? issues relating to the Public Report2, including definition, the requirements for Competent Person sign-off, level of disclosure and mandatory requirements; ? guidance on techniques to quantify resource / reserve estimation uncertainty; ? the level of study required to convert Mineral Resources to Ore Reserves, and disclosure of prices and economic factors; ? expansion of the requirements for reporting of marketable material for bulk commodities; ? the use of language and tone that anticipates new and emerging technologies and techniques for the qualitative and quantitative measurements of data; and ? linkages to valuation and the valuation code (The VALMIN Code). Although the Canadian and Australian reporting regimes for public statements on Exploration Results, Mineral Resources and Mineral (Ore) Reserves are similar in intent and coverage, they differ in detail, sometimes substantially. Canadian requirements are generally more wide-ranging and prescriptive than Australian requirements, including in the need in many situations to publicly lodge a Technical Report, a requirement that does not exist in Australia. Australian-listed companies seeking to list in Canada, and Competent Persons seeking to act as Qualified Persons in Canada, need to ensure that they are fully conversant with Canadian regulations and how they differ from those in Australia.

May 1, 2010

By Serge Ouellet

Identification and quantification of cementitious phases in cemented paste backfill (CPB) is difficult due to the fineness of tailings and the low binder content used in the mixture. Techniques such as X-ray diffraction and thermogravimetry were ineffective (mainly due to tailing mineral interference) to identify and quantify hydration products in most works reported in the literature. On the other hand, scanning electron microscopy (SEM) was used in many cases but mainly for qualitative purposes. This paper presents the development of a new approach based on a computer program (SEM-XMAP) that combines X-ray dot mapping images acquired by SEM and generates a coloured image representing the mineralogy and the microstructure of a given CPB specimen. The technique is applied to one CPB sample mixed with blast furnace slag (BFS) and acidic sulphated mixing water sampled in a mine dewatering system. The technique allowed identifying calcic matrix, unhydrated and partially hydrated BFS particles, and sulphated phases.

May 1, 2007

By S. Kelebek, E. Yalcin

"Two composite samples of a gold-bearing ore from North Western Quebec (Barry Township) were tested to assess their grindability, recovery by flotation and Knelson concentrator as well as extraction using cyanidation.Work index values (BWi) for two composites were found to be 10.4 and 11.0 kWh/tonne, respectively. Flotation tests involved feed samples for which the particle size ranged from a P80 value of 205 ?m to 53 ?m, which resulted in gold recoveries of 91.8% to 95.8% (at gold grades of 130 to 140 g/tonne). Cyanidation tests indicated gold extraction levels ranging from 94.2 to 97.5%. Higher, recoveries were obtained at finer grind sizes and higher cyanide strengths. However, the grind size had a bigger impact on gold leaching. Gold grade and recoveries obtained with the Knelson concentrator device were lower than those obtained by flotation. Mass recoveries to concentrate products strongly depended on the fluidization pressure used. High gold recoveries can be obtained generally at low water pressures of 3 to 4 psi. Higher water pressures produced more selectivity at the expense of higher gold losses.It can be concluded that the gold ore samples tested have good characteristics due to moderate levels of energy requirement, high gold recoveries and extraction rates attainable.INTRODUCTIONBench scale testing of ore samples is an important step in the long series of metallurgical and economic studies involved in bringing new ore bodies into production (Coleman, 1980). Through a laboratory testing program, the behaviour of an ore in a particular process on a larger scale can be established. In an exploratory manner, this testing can be helpful for classification of ores in terms of “good ore” or “problematic ore” by providing preliminary information on recoverability of target minerals. Potential problems if any can be assessed by further investigation of products (e.g., middlings) obtained as a result of testing. A detailed bench scale experimental program in the metallurgical laboratory can also provide the background work for flowsheet development. Depending on the simplicity or complexity of the ore to be tested and type of details required from a testing program, the amount of samples required in the mineral processing laboratory can vary from several kg to more than one hundred kg depending on how detailed the metallurgical testing program is."

Jan 1, 2008

"7.1. Glossary of the Platinum-Group Minerals7.1.1. IntroductionA glossary of the platinum-group minerals (POM) serves as a useful introduction to a chapter on their mineralogy. Earlier glossaries (Cabri, 1972, 1976) were prepared for different purposes and composed in a different manner. As for most mineralogical textbooks, from Dana's System of Mineralogy, 3rd Edition (1850) , to Fleischer's Glossary of Mineral Species (1980), this glossary is primarily a chemical classification, arranged here in two tables. Table 7.1 groups the mineralforming compounds of the six platinum-group elements (POE) according to their essential anion(s) or according to the nonPOE element in intermetallic compounds. Table 7.2 is a similar listing for the native metals and alloys of the POE. Both tables include unidentified POM considered to be bona fide new minerals requiring additional study for complete characterization and recognition (see Chapter 9*). Some information on the chemistry of the POM is given in the sections on X-ray data (7 .2) and ore microscopy (7 .3). Because the overwhelming majority of POM have been characterized by the electron microprobe, a more detailed account of representative analyses is given in Chapter 8**. The total POM listed in Tables 7.1 and 7.2 is about 100, after deleting some which would not qualify as discrete species if found today, but which are still recognized on historical grounds; e.g. iridosmine, osmiridium, platiniridium. Tables 7.1 and 7.2 also include 28 unnamed POM; included because they are considered to be probably new species (Cabri 1981 b). The referral scheme for unnamed minerals is the same as in Chapter 9. There are considerably more palladium minerals known than other POM; the proportion of total species (listed in Tables 7.1 and 7 .2) for the POE is: Pd 48%, Pt 21 o/o , Ru and Ir 9% each, Rh 8% and Os 5%. These proportions are in marked contrast to the estimated cosmic abundance, as compiled for average type C1 chondrite by Naldrett and Duke (1980): Pt 29%, Ru 20%, Pd, Ir and Os 15% each, and Rh 6%."

Jan 1, 1979