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By P. Leo

The Al-Zn-Mg alloys (e.g. 7004, 7020) are employed for land transport and architectural applications whereas the Al-Zn-Mg-Cu alloys (e.g. 7075, 7050) provide the high strength aircraft alloys. The Cu gives rise to suitably sized and distributed precipitation but causes high quench sensitivity. Their hot work processing also has marked differences; the Cu-free alloys have good extrudability whereas the Cu-bearing alloys have low extrudability and are difficult in rolling as well. The behaviors of the two classes are described in terms of flow stresses and ductility over 300 - 500°C, along with constitutive contrasts. The ductility of the alloys with Cu is much reduced above 400°C by precipitation of a phase at the grain boundaries that can be avoided by suitable heating schedules. The high strength alloys are subject to de-lamination corrosion of sheared edges due to severely pancaked grains; this can be alleviated by holding between passes, in mid-rolling schedule. In torsion tests at 10-1 to 10 s-1/ 250 - 500°C, the stresses (ductilities) at 300°C/ 1s-1 and activation energy Q for 7004 were 118 MPa (2) and 262 kJ/mol. The 7075 was tested after several pretreatments: completely over-aged (softened): 107 MPa (5), Q=150 kJ/mol; whereas, solution and quick cooling to testing caused extraordinarily high peaks related to dynamic precipitation resulting in 160 MPa (1.2), Q=300 kJ/mol.

Jan 1, 2011

By Deon van Tonder, Chuck Edwards

"Solvent extraction is used in many acid leach uranium circuits to upgrade and purify the uranium bearing solution prior to final product precipitation. A number of strip options are available to recover the uranium from the loaded organic. This paper compares some of these options on the basis of typical capital and operating costs for both the solvent extraction circuit and downstream uranium processing, and also considers other related issues. The aim is to provide an objective comparative assessment of these options that could be used as a qualitative guide in process selection or improvement.INTRODUCTIONIn the 1950s the uranium industry was the first industry to adopt solvent extraction (SX). Since then, the flow sheets have advanced. Notable was the introduction of tertiary amine extractants by General Mills during the late 1950s and 1960s (Mackenzie, 1997; Mackenzie, 2005), which is still today considered ‘state of the art’ technology. Initially solvent extraction was done in combination with ion-exchange, such as in the Eluex flow sheet, followed later by the adoption of direct solvent extraction from pregnant leach solution (PLS). From a strip-precipitation perspective ammonia strip followed by ammonium diuranate precipitation was widely adopted as the norm in the uranium industry. More recently alternative strip systems have been receiving increasing attention due to the environmental, health and safety issues associated with ammonia and the resulting calcined uranium oxide product. This paper explores some of these alternative options in comparison with the ammonia route.It is important to acknowledge that each operation is unique, and therefore the information and philosophy outlined in this paper serve to provide a general guide that can be potentially applied and adapted to any site or project."

Jan 1, 2013

By Claude Bazin, François Robichaud

"The control of the operation of semi autogeneous grinding (SAG) mills is mainly achieved through adjustments of feed rate, water addition and speed of rotation. With the introduction of image based sensors it is possible to have on-line measurements of the size distribution of the ore feeding a SAG mill. This information could be used for the development of a control strategy to regulate a SAG mill feed size distribution to a set-point. The control of the size distribution is carried out by adjustment of the speed of the ore extractors underneath the ore stockpile. The potential and difficulty associated with the implementation of such control are assessed using a multivariable statistical analysis of data collected from the operation of the grinding circuit of the Agnico-Eagle Goldex mine in Val d’Or, QC. A control strategy for the ore feed size distribution is assessed using a simulator of the ore stockpile operation.INTRODUCTIONThe popularity of autogeneous (FAG) and semi autogeneous (SAG) grinding mills has been increasing over the last 20 years (Napier Munn et al., 1996). These mills use the ore or a combination of ore and steel balls as the grinding media. Since the ore is used as a media for the size reduction the mill ore feed size distribution impacts the performance of the mill. Several papers were recently published on this subject (Dance et al.,2011; Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011) showing the advantages of preparing the feed size distribution to maximize the circuit throughput or mill efficiency (Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011). With the availability of particle size sensor based on image analysis of the ore on a conveyor (Partanen et al., 2001; Bobo et al. 2004; Edwards et al., 2002; Wei & Craig, 2009) it is possible to control the size distribution of the ore feeding autogeneous or semi autogeneous mills (Partanen et al., 2001).This paper describes a method to control a size distribution index to a set point via the manipulation of the speed of the extractors that withdraw the ore from the stockpile feeding the grinding circuit. The paper consists of 4 sections. The first section describes the considered grinding circuit. A statistical analysis of the collected data on the grinding circuit is presented in the second section. This analysis demonstrates the necessity of using real-time adaptation mechanisms to develop a control strategy for the size distribution of a grinding circuit. The third section describes the method used to simulate the ore stockpile operation and the last section describes the preliminary development and assessment of a control strategy of size distribution of the ore feeding a SAG mill."

Jan 1, 2013

By Gerald Whittle

What is Enterprise Optimization? Releasing value through better planning decisions: Philosophy, Methodology, Software The key to 5% -35% increase in enterprise value (NPV) The culmination of 25 years development in optimization for the mining industry ?Jeff Whittle Simultaneously optimizing: ?all steps in the value chain ?all assets in the enterprise portfolio ?all periods together

May 1, 2010

By R. R. Klimpel

"SYNOPSISA discussion of the importance of flotation rate in the analysis of mineral flotation is presented. In particular, the highly interactive nature of the flotation ""system"" is emphasized where there are components of chemistry, equipment, and operating conditions simultaneously involved. It is further shown that: 1) flotation rate must be measured in order to economically optimize the performance of an industrial flotation circuit, 2) that consistent patterns exist in both flotation rate and equilibrium recovery when changing various settings of the system components, 3) that knowledge of these patterns is necessary in order to avoid making system changes that are either self-compensating or dramatically re-enforcing in some particular aspect of performance, and 4) the rate related role of the chemistry component (e.g.frothers and collectors) is often much bigger than is sometimes realized (especially in a negative sense involving overdosing of reagents). INTRODUCTIONFroth flotation is widely usedl-6 in mineral beneficiation processes as a method for separating different types of finely divided solids present in a liquid (usually aqueous) suspension. The approach is based on the contacting of solids with upward moving air bubbles in an agitated vessel where the solids are exposed to chemicals (collectors) which create a condition favoring the attachment of a certain portion of the solids to the air bubbles. This selected portion is then removed in a froth stabilized by chemicals (frothers) and subsequently recovered from the flotation vessel. Depending on the settings of the various components of the flotation ""system"" (see Figure 1) it is possible to achieve a wide range of flotation performance."

Jan 1, 1984

By C. J. Van Treeck

PGE concentrations of 18 g/t occur in the pyroxene rich rocks at the Union Bay Alaskantype complex. This PGE enrichment is hosted in magnetite veins. These veins cross magmatic layering and have an irregular morphology in both outcrop and thin section. Fe-Ti oxide geothermometry suggests that this mineralization took place between ~700-400°C. A variation in the amount of ilmenite and Mg- Al spinel exsolution in magnetite correlates with temperature estimates, lower temperature magnetite has less exsolution. Hydrous silicates form an alteration envelope around magnetite veins with a formation temperature of <575oC, vary with temperature, and contain appreciable amounts of Cl. Interaction of a PGE-Fe-Cl rich fluid with clinopyroxene increased the fluids pH and decreased PGE and Fe solubility depositing a PGE enriched magnetite within the Pyroxenites and Olivine Pyroxenites of the Union Bay complex.

May 1, 2003

By R. Barbaroux, B. Dubreuil, D. Lemieux, K. Nasrallah, G. Garcia Curiel

"As conventional, base or precious metal ores are being depleted worldwide, the mining industry is being faced with an increasing exploitation of arsenical sulfidic ores. Enargite is a common example. The downstream processing of these ores poses several problems ranging from high smelter penalties to finding economical and environmentally acceptable means of disposing of the arsenic. Several hydrometallurgical processes have been proposed for extracting arsenic from enargite concentrates, notably using alkaline sulphide leaching (ASL). However, few alternatives exist for removing the arsenic from solution and stabilizing it in an approach that is both economical and environmentally acceptable.Dundee Sustainable Technologies (“DST”) has developed a technology for the sequestration of arsenic by vitrification. Originally optimized to cost-effectively stabilize arsenic trioxide (As2O3), DST has shown that vitrification could be successfully extended to the stabilization of other arsenical products, such as calcium arsenate (Ca3(AsO4)2) and sodium arsenate (Na3AsO4). In this paper, the vitrification of arsenic in the context of an enargite leaching was newly explored and demonstrated successfully.Two processing routes are being proposed by DST for the removal of arsenic from enargite concentrates and their stabilization by vitrification. In one route, arsenic is extracted by ASL and is then precipitated under atmospheric conditions and with minimal use of reagents and time. The resulting precipitate is an unstable hydrated ferric arsenate (FeAsO4•2H2O) that is then vitrified in a non-dilutive method, significantly increasing the stability of arsenic and meeting the TCLP guideline. Unlike is required for scorodite, stabilization was successfully achieved without concern for the crystallinity of the precipitate."

Jan 1, 2017

By T. Holmes

Investigations reveal that the throughput difference between a perfect production day and the average is often on the order of 40%, and that bulk solids handling system deficiencies are the largest (>50%) class of lost opportunities. Root causes for underperformance of elements in material handling systems can often be traced back to the detailed design and even study-stage oversights. To reduce or eliminate bulk solids handling problems in an existing system requires an analysis of troublesome handling equipment – for example, a concentrate transfer chute that experiences plugging, poor flow to the furnace, unacceptable dust generation, and spillage. These problems often lead to significant maintenance costs related to downtime and personnel time. Understanding the cause and cost of flow issues is the first step in troubleshooting the problem. Next, measuring the material’s flow properties is critical to providing design data required to develop a comprehensive scientific solution. In the case of a new installation, the design data is required to properly design the system right first time and prevent flow problems from occurring so that the new equipment works correctly from start-up. This paper highlights the cost of material handling problems and discusses the best approach to evaluate material handling system deficiencies. A case study is given of how Anglo American's operation in Barro Alto Brazil enjoyed a 5-month payback from a bulk solid transfer chute debottleneck project. INTRODUCTION When operating a plant or designing a greenfield or brownfield operation, the intent is to have perfect production days (PPD) each and every day. Money is made by moving material through the plant in a controlled manner to produce the desired product. Mining operations and copper concentration processes are at the most basic level a practice in bulk solids handling (BSH). Unfortunately, many of the project planning steps are overlooked, particularly related to BSH. Investigations reveal that the throughput difference between a PPD and the actual average is often on the order of 40%, and that BSH system deficiencies are the largest (>50%) class of lost opportunities (Wellwood, 2017). Root causes for underperformance of elements in material handling systems can often be traced back to the detailed design and even study-stage oversights. BSH issues are essentially baked into the fabricated plant and left to operations to accept the deficiencies and deal with them for the life of the mine. These deficiencies then become part of the normal operation of the mine and plant.

Jan 1, 2019

By G. C. Baker

Despite the attention belatedly bestowed on tidal power in recent years, it remains largely a potential rather than an actual source of energy. The resource is, practically speaking, non-depletable. Extraction is technically feasible and environmentally acceptable. The restraints against development are economic, and they could be removed by technical improvements, changes in the cost or availability of other energy sources, or simply by a change in the cost of money. The exponential rise in the demand for energy makes it inevitable that the tides will be harnessed sooner or later.

Jan 1, 1972

By Euler De Souza

This paper presents a study conducted to assess sillmat behaviour during undercut mining and to provide a comprehensive understanding of the effects of arching and sillmat thickness on stability. The design study investigated the effects of such parameters as stope width and height, closure strains, orebody geometry and inclination, wall roughness, fill properties and cure time on the stability of cemented sillmat during undercut mining. Arching effects have been quantified based on centrifuge modeling studies for the different mining conditions. Observations have indicated that the degree of arching is a function of stope geometry, wall roughness, fill cohesion, and wall closure. Arching effects increased with decreasing stope width, increasing backfill cohesion and frictional resistance, increasing wall roughness and with wall closure.

May 1, 2002

By Peter Bourke, Marc van Olst, Seppo Ronkainen, Natalie Brown

"Since the early days of flotation, process technicians have known that the speed at which the froth is recovered over the lip of the cell has a very direct and consistent influence on the grade and recovery of the circuit. No plant would deny that, even to this day, one of the most important aspects of a process technician’s job is ensuring that the froth is moving at a desired and controlled speed over the lip. Unfortunately, process technicians cannot watch the froth throughout the entire shift, as there are always a lot of other tasks to do.A considerable amount of work has thus been done to develop a “system” that automatically measures the speed of the froth. This measurement would then be used in simple controllers to emulate the steps taken by the process technician in correcting deviations of the froth speed from its desired level. Many of the earlier attempts to measure speed were ineffective, as they were unable to cope with turbulence on the froth surface. The FrothMasterTM instrument makes use of machine vision technologies to measure the speed of the froth over the lip of a cell at a very high sampling rate. This instrument produces an extremely reliable and accurate measurement of the froth speed (as well as other parameters such as bubble size and froth stability). As we will show, the effectiveness of the speed measurement in controlling the performance of the plant is a testimony to its reliability.Three FrothMasterTM units have recently been installed on the Cadia Hills Gold Mine in New South Wales (Australia). The objectives of this installation were:To control the mass-recovery of two Tank Cells in series on the same level where differences in the hydrostatic head between the two cells result in the cells pulling inconsistently with respect to each other.Control of the concentrate grade (%Copper) from the first rougher.Reduction in the frother consumption in the rougher circuit by monitoring the froth speed.A simple controller was designed to react to disturbances in the same way that human operators react. This was done after consultation with the metallurgists and the process technicians at Cadia. The controller had two functions:Stabilisation: The stabilising controller made use of three manipulated variables (level, frother addition rate and aeration rate) to control the froth speed to a setpoint.Optimisation: The optimising controller adjusted the froth speed setpoint based on deviations of the grade (measured in real-time with an OSA) from a setpoint. The plant metallurgist adjusted the grade setpoint based on information about the ore."

Jan 1, 2001

"It is projected that the reactivity of acid generating tailings and waste rock can be minimized by reducing the access of oxygen by placing these materials under water in natural or manmade lakes. A project under the MEND and B.C. AMO Task Forces has been in progress for four years to evaluate active and historic tailing sites in B.C. and Manitoba. The sophisticated geochemistry studies conducted on the interstitial pore water has illustrated encouraging results which are discussed. The additional studies which are needed to complete the scientific database for the study sites, for achieving regulatory and public approval and for the engineering of application of this method of control are outlined.IntroductionWorld wide research effort is in progress to find the solution to the acid mine drainage problem. A great deal is known about acid mine drainage but much work is required to find economic and effective approaches to the problem. Two coordinated programs have been initiated recently in Canada: the National Mine Environment Neutral Drainage (MEND) program and the British Columbia Acid Drainage Task Force.Disposal of tailings materials under a water barrier such as a natural lake, is thought to be one solution to prevent acid generation by preventing oxygen and bacterial action on the sulphide surfaces. The deposition of reactive tailings materials under water in natural lakes has been used historically at various mines in Canada. However, there has never been a thorough review of the long and short term impacts of historic sites which may provide crucial information to understand the physical, chemical and biological processes which are taking place.This paper provides a general overview of a series of studies which examine the general biophysical characteristics and the sediment geochemistry of four lakes where subaqueous disposal has been or is currently being used for acid generating tailings. The subaqueous disposal studies have been conducted from 1989 through 1991 and were initiated and supported by both research task forces. The objectives of this paper are to summarize the history and the background information of the studies conducted, the general results obtained from each lake, an overview of the major collective results and the additional work which is required to achieve the final objective of developing criteria for applying this technology in the industry."

Jan 1, 1992

By Frederick W. Breaks, Andrew G. Tindle, Julie B. Selway

"Abstract - Rare-element pegmatites may host several economic commodities, such as tantalum (Taoxide minerals), tin (cassiterite), lithium (ceramic-grade spodumene and petalite), and cesium (pollucite). Key geological features that are common to pegmatites in the Superior province of Ontario and Manitoba, Canada, and in other large tantalum deposits worldwide, can be used in exploration.An exploration project for rare-element pegmatites should begin with an examination of a regional geology map. Rare-element pegmatites occur along large regional-scale faults in greenschist and amphibolite facies metamorphic terranes. They are typically hosted by mafic metavolcanic or metasedimentary rocks, and are located near peraluminous granite plutons (A/CNK > 1.0). Once a peraluminous granite pluton has been identified, then the next step is to determine if the pluton is barren or fertile. Fertile granites have elevated rare element contents, Mg/Li ratio < 10, and Nb/Ta ratio < 8. They commonly contain blocky K-feldspar and green muscovite. Key fractionation indicators can be plotted on a map of the fertile granite pluton to determine the fractionation direction: presence of tourmaline, beryl, and ferrocolumbite; Mn content in garnet; Rb content in bulk K-feldspar; and Mg/Li and Nb/Ta ratios in bulk granite samples. Pegmatite dikes with the most economic potential for Li-Cs-Ta deposits occur the greatest distance (up to 10 km) from the parent granite.Metasomatized host rocks are an indication of a nearby rare-element pegmatite. Metasomatic aureoles can be identified by their geochemistry: elevated Li, Rb, Cs, B, and F contents; and by their mineralogy: presence of tourmaline, (Rb, Cs)-enriched biotite, holmquistite, muscovite, and rarely garnet.Once a pegmatite dike has been located, the next step is to assess its potential to contain Ta mineralization. Pegmatites with the highest degree of fractionation (and thus the most economic potential for Li-Cs-Ta) contain blocky K-feldspar with >3,000 ppm Rb, K/Rb < 30, and >100 ppm Cs; and coarsegrained green muscovite with >2,000 ppm Li, >10,000 ppm Rb, >500 ppm Cs, and >65 ppm Ta. Pegmatites with Ta mineralization usually contain Li-rich minerals (e.g., spodumene, petalite, lepidolite, elbaite, amblygonite, and lithiophilite) and may contain Cs-rich minerals (e.g., pollucite, Cs-rich beryl). The ore minerals of Ta are commonly manganotantalite, manganocolumbite, wodginite, and microlite; Ta-rich cassiterite is also commonly present. Tantalum mineralization tends to occur in albitic aplite, mica-rich (lepidolite, cleavelandite ± lepidolite), and spodumene/petalite pegmatite zones."

Jan 1, 2005

By P. G. Claridge, G. P. Holmes, M. A. Redfearn

"IntroductionThe purpose of this chapter is to describe current practices in operation and maintenance of flotation circuits in Canadian mineral processing plants. The chapter consists of five sections covering description of equipment, typical flow diagrams, operating practice, instrumentation and maintenance. The approach to this subject matter is intended to be practical in nature, hence no deeply technical treatment of flotation theory will be included.Flotation is an extremely common method of beneficiation of ores. Reasons are the relatively low cost of flotation compared to many other processes, the wide range of minerals to which flotation may be applied and the high efficiency of separation which can usually be achieved. The flotation process as we know it today has been developed within this century and it has had a major impact on the mining industry. While flotation is commonly considered to be important because it provides a relatively simple method of separating minerals in complex ores, the availability of flotation has also made a multitude of mineral deposits in Canada into economically viable mining operations because it allowed production of a concentrate of valuable minerals from the ore. These concentrates could then be transported from geographically remote locations to centraHzed plants for further processing, although the existence of the remote and perhaps small ore body alone would not have justified installation of metallurgical processing plants on site and the grade of ore would not have justified direct shipment of ore.Some of the more common minerals concentrated by flotation are sulphides of copper, nickel, cobalt, iron, zinc, molybdenum, lead and arsenic, as well as non-sulphides such as sodium chloride, potassium chloride, phosphates, silicates, chromite, fluorite, limestone and coal. Precious metals are also recovered by flotation, generally in association with another sulphide mineral."

Jan 1, 1989

By Y. Bai, J. J. Liu, D. Q. Li, S. Z. Liu, X. T. Wang

"This paper presents the results of an industrial test using 8 t/d of Baotou concentrate (50% REO) of a clean metallurgical process, which includes: 1. Roasting with H2SO4 at low temperature (220~260 0C) in a batch roasting furnace followed by water leaching to obtain a leach liquor as feed with a yield of >95% RE and Th, recovering F as fluoride, producing a leach residue amounting to 1/3 of an equivalent high temperature residue and containing <0.03% Th and radioactivity < 3×104 Bq/kg compared to the normal high temperature residue containing 0.2-0.3% Th and 1×105 Bq/kg. 2. Separating Th with N1923 and extracting RE by P507 has been completed in mixer-settlers. Test results indicate that the ThO2 product has a purity of > 95% with a yield of 99%, and the RE has a yield of >95%, in which ThO2/REO <5×10-6, and that P507 loaded RE connect straight with separating RE using P507-HCl system. 3. Treatment and multipurpose use of "three wastes" (water, gas and solids) are discussed in some detail."

Jan 1, 2014

By Dave Marsh, John Goode, Niels Verbaan, Kevin Bradley, Henrie Notzl

The mineralogy of the Avalon Nechalacho deposit is complex with significant fractions of HREEs, Zr, and Nb found in refractory minerals such as zircon and fergusonite. Sulphuric acid baking followed by water leaching can be used to extract REEs from flotation concentrates, but acid baking does not completely "crack" the zircon and fergusonite, so the extractions of Zr, Nb, and HREEs are not high. An alternative flowsheet has been developed that is able to obtain significantly improved extractions of HREEs, Zr, and Nb from flotation concentrates or acid bake residues. This is accomplished by cracking the zircon and fergusonite minerals with caustic soda at high temperatures, followed by water washing to recover excess caustic soda and soluble sodium silicates and phosphates. REEs, Zr, and Nb in the solid residue can then be solubilized with HCl or H2SO4. This paper highlights some of the findings from an extensive laboratory program and discusses several flowsheet implications.

Jan 1, 2014

By Jocelyn Bouchard, Claude Bazin, Sami Makni, Rolando Quispe

"Finely disseminated minerals in ores require intensive size reduction to achieve sufficient liberation for the subsequent concentration process. The use of small grinding media in conventional ball mills as a replacement for the usual 2.5 to 3.8 cm balls (1.0 to 1.5 in) produces a finer size distribution at a lower power drawn by the mill motor. However, the conditions ensuring a successful application of using small media for fine grinding are still not fully understood and some plants faced issues to keep the newly introduced steel charge inside the mill during industrial trials. The retention conditions of small media are investigated in this paper using a continuously operated 40 x 40 cm laboratory grinding mill. The tests assess the effect of the steel charge level, the slurry solids concentration and the feed rate on the rate of ejection of the grinding media. The dominant variable is found to be the filling of the mill with the grinding media, followed by the mill feed rate. Results show that increasing the slurry solids concentration within a grinding mill may be favorable for the retention of the media inside the mill.IntroductionProcessing of complex ores often requires the use of fine grinding to liberate finely disseminated minerals. Grinding requirements for a P80 of less than 0.02 mm are not uncommon, and different grinding mills are available for such application (Leonida, 2014; Pease et al., 2006; Davey, 2006; Orford et al., 2006; Nesset et al., 2006). Even conventional horizontal ball mills charged with of 2.5 cm or larger grinding media can be used for fine grinding. However, simulation (Orford and al., 2006) and plant observations (Alves and L-Bouchet, 2005; Brisette, 2010; Brisette and Roman, 2012) demonstrated the advantages of using smaller grinding media in terms of size reduction and energy consumption."

Jan 1, 2015

"Mineral separation by exploiting the differences in mineral or metal specific gravities (sg) has been used since antiquity. In the modern era, this is most commonly seen in the recovery of gold by the use of centrifugal concentrators and the separation of coal (low sg) from waste rock (high sg) using heavy media cyclones and jigs. Other metals, notably oxides of tin and tungsten, are successfully recovered using some combination of jigs, spirals and shaking tables. However, many conventional gravity separation systems are limited to fairly coarse (>1 mm) to medium (>100 micron) particle separations due to the limitations of the gravitational separation forces in conventional equipment.Many of the high specific gravity minerals of interest are friable (e.g. cassiterite, wolframite, etc.) and excessive fines are generated in the grinding/comminution circuits which cannot be recovered with conventional equipment and may be lost directly to the tailings stream. Recent developments in centrifugal gravity separation equipment specifically designed for fine or ultra-fine particle separation and recovery have resulted in heavy mineral particle recoveries as fine as 10-20 microns. Ongoing research work has also identified the benefit of closed-circuit, multi-stage gravity recovery circuits using some combination of centrifugal concentrators and conventional equipment (tables, spirals, flotation, etc.) which can substantially increase overall plant recovery and final concentrate grade.This paper will provide details of the evolution of non-fluidized bed gravity concentrators for high sg mineral recoveries, correlate laboratory scale development work with plant results to confirm scale-up, and show how recent flowsheet developments have contributed to significantly improved plant performance.IntroductionGravity concentration is one of the oldest and most cost effective metal and/or mineral recovery systems, ranging from the classic hand panning of alluvial gold, through to the large scale production of a wide range of metals and minerals, including gold, silver, PGMs, black sands, coal and iron oxides. Conventional gravity recovery systems, comprised of some combination of sluices, spirals, jigs and shaking tables have limited application in fine particle (i.e. <100 microns) recovery, or where there are limited differences in particle specific gravities (sg), e.g. the separation of high sg, high value metal oxides (e.g. those of tin, tantalum or tungsten), from medium sg, lower value metal sulfides such as pyrite and pyrrhotite. These unit operations tend to have limited capacity, and while double and triple turn spirals and multi-deck tables have been developed, extensive floor space is often required in order to process the tonnages incorporated in the modern mineral processing plant."

Jan 1, 2017

By P. Radziszewski

"The comminution energy curve is a profoundly simple graphical technique that displays detailed operating efficiency information for industrial comminution processes. The database underlying the methodology has recently more than doubled, to now contain a substantial fraction of the international comminution output (1.7 billion tonnes of rock per year) across more than 8 major commodities. This increase has deepened the analytical capabilities and effectiveness for assessing comminution circuit performance in relation to comparable operating mines. This paper will explore how the energy curve methodology can be used to examine the ability of stirred mill technologies to increase efficiency.Specifically, this paper explores how the energy curves methodology can be used to present energy efficiency gains achieved at three mines employing a stirred mill technology under different duties. The first of these cases studies was conducted by Rosa (2014) who compared the performance of two ball mills operating in parallel with a Vertimill in the regrinding circuit at the Samarco Iron project in South East Brazil. The Vertimill was found to achieve the same product specification while consuming 40% less energy. Brissette (2010) investigated the impact of grinding media on the performance of both ball mills and Vertimills. The energy savings achieved varied from 10% to 40% for the ball mills and 30% to 60% for the Vertimills. Lastly, Bergerman (2012) investigated the operating conditions of the Sossego regrind circuit from 2006 to 2011 against design specifications and laboratory testing. The operating work index of this circuit varied from 5.2 kWh/t to 44 kWh/t over this 5 year period. The full suite of energy curves"

Jan 1, 2016

By John S. Stevenson

ZEBALLOS is a relatively new mining camp. It is on the west coast of Vancouver island, approximately 195 nautical miles northwestward from Victoria. Previous to 1934, a small amount of placer gold had been recovered from the Zeballos river and tributary creeks. However, it was not until the winter of 1934-35, when shipments of high-grade gold ore were made from the property of the White Star Gold Mines, Limited, that lode gold mining really began. Since then, several new mines and many promising prospects have been found. Milling in standard sized mills only commenced late in 1938, when the 75-ton mill of Privateer Mines, Limited, and the 50-ton mill of Spud Valley Gold Mines, limited, were first turned over. The most striking feature of the camp is the high-grade quality of the ore as usually mined. At one property, the ore along a drift 650 feet long has ranged from 5.8 ounces per ton to 17.5 ounces, over average vein-widths ranging from 8.3 inches to 14.5 inches in selected blocks along the drift. One . high-grade shipment of 196.6 pounds of ore from the same property contained 142.877 ounces of gold. The present paper deals only with that part of the area in which mining operations are most active, namely the main Zeballos River valley and its tributaries, Spud and Goldvalley creeks. During the summer of 1938, prospecting was carried on beyond this circumscribed area. The geology and early developments in the area have been described in the following reports: (1) CLOTHIER, G. A., Annual Report of the British Columbia Minister of Mines, 1929, p. 376; 1930, p. 441; 1932, p. 205; and 1933, pp. 252-254. (2) GuNNING, H. C., Geological Survey of Canada, Summary Report, 1932, Part A-II, pp. 29-50, Zebaltos River Area, Vancouver Island, B.C. (3) STEVENSON, John S., Annual Report of the British Columbia Min-ister of Mines, 1935, pp. F-38-40. ? ( 4) BANCROFT, M. F., Geological Survey of Canada, Memoir 204, 1937, Gold-Bearing Deposits on the West Coast of Vancouver Island Between Esperanza Inlet and Alberni Canal. (5) STEVENSON, John S., British Columbia Department of Mines, 1938, Lode-Gold Deposits of the Zeballos Area. (6) MAcDoNNELL, G. F., The Privateer Mine, Zeballos; The Miner, Van-couver, B.C., December, 1938.

Jan 1, 1939