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By S. H. Lucas

"This paper describes the development and application of single-stage fluidized bed continuous ion exchange to the following processes: removal of base metals from mine water; cyanide removal from gold mill waste; uranium removal from sulphuric acid leach liquor; and uranium removal from mine water. IntroductionThe recovery of metals by fixed bed ion exchange has been practiced for many years. However, the operation has been semi-continuous and can only treat clarified liquors. Substantial savings could be made if the operation were continuous, thus reducing the resin inventory, and if the costly filtration and clarification steps could be eliminated. Work has been in progress over the past twenty-five years on the recovery of metals from unfiltered ore leach slurries, heap leach liquors and mine drainage waters containing a small amount of suspended solids. In the case of dilute waste solutions, with suspended matter and high flow rates, filtration is usually prohibitively expensive.Continuous ion exchange operation using fluidized beds has made it possible to reduce considerably the resin inventory and to allow the treatment of unclarified liquors and slurries. The subject has been reviewed on several occasions (1. 2. 3, 4). Although fluidized beds in multi-stage columns are an improvement over previous resin-in-pulp treatment, they have complicated internal structure and I or involved automatic control systems to move the resin stages.Slater and Prud'homme performed studies on multi-stage fluid beds, and concluded that the performance of single, deep fluidized beds of resin with continuous or semi-continuous flow of both phases would be worth investigating, particularly in the case of very dilute metal solutionsoi, A pilot plant was designed and built at CANMET in a cooperative project with Dr. Slateroi. The use of deep fluidized beds of resin (with a large height/diameter ratio) for treating dilute solutions has the attraction of engineering simplicity. Extreme simplicity of equipment design is considered of paramount importance when treating, for example, mine waters at high flow rates, possibly of the order of several million gallons per day. This paper describes the development of the concept of single-stage fluidized bed continuous ion exchange arid reviews the pilotplant development of processes for the treatment of base metal mine water, cyanide waste water, uranium leach liquor and uranium mine water."

Jan 1, 1980

By Y. Prevost, D. Verhelst

"A computer-assisted process control system for the Noranda Process Reactor Wa5 installed at the Home smelter in 1991. Operating as an advisory tool for the furnace operators, and using a flexible oxygen balance as the basic building block of the computer program, the new control system has helped to reduce the variability in the grade of the matte produced by 50% . The smelter is currently operating at a 72% copper matte, with a standard deviation of between 1.0 and 1.3, while treating a wide range of concentrates, secondary materials and scrap. Enhanced process control has allowed for longer reactor campaigns, and improved stability of the converting operation.This paper reviews the feed classification and blending practice used at the smelter, outlines the strategy of the new computer control system, and summarizes the benefits realized since the implementation of the system in early 1990.IntroductionUntil 1989 the Noranda Home smelter operated as a multi furnace smelter, at one time having three hot-charge reverberatory furnaces, and finally operating with one wet-charge reverberatory furnace and one Noranda Process Reactor. In 1989 the final reverberatory furnace was closed and the Reactor became the sole smelting vessel, treating around 750 000 tonnes of concentrate, secondary materials, and scrap per year. At the time of the start-up of theNoranda Reactor in 1973, the vessel was capable of treating 750tonnes per day of a heterogeneous mixture of copper concentrates, fluxes and secondary materials to produce copper in one operation. Through process improvements, and the introduction of additional oxygen in 1982(pannell and Mackey, 1988), the daily tonnage increased to 1800 tonnes by 1988. The operation had also been changed to produce a high-grade matte instead of copper. Figure1 illustrates the main features of the reactor, a cylindrical vessel which is charged at one end and tapped at the other. The matteis subsequently processed through Peirce-Smith converters and anode furnaces before being cast into copper anodes. Figure 2 outlines the progressive increase in the furnace smelting capacity. Over this period the percentage of secondary materials, and scrap feeds also increased to approximately 110 000 tonnes per year, or 15% of the total new feed input (Bedard et al., 1991). Now operating wholly as a custom feed operation, the Home smelter can treat over 2200tonnes per day of new metal-bearing material."

Jan 1, 1993

By Chia-Chi Chiu, Meng-Chia Weng, Tsan-Hwei Huang

"This paper proposes a biconcave shape bond model for the distinct element method (DEM) to simulate cementation between two circular particles in geomaterials. Rather than an original parallel bond model between particles, the proposed model adopts the more realistic shape of cementation and considers the elastic response of cementation under external loading. The stress field and force-displacement relationship of the biconcave shape bond is based on Dvorkin’s theory, and three modifications are adopted: (1) A superposition method is applied to improve the correctness and symmetry of the stress field, and it also provides a method of addressing the status when both connected particles rotate; (2) To simplify the input parameters, the tangential movement is replaced by the normal deformation and rotation between particles; (3) To accelerate the calculations, variables are extracted in analytical procedures, which provide an efficient method for applying the bond model with only one matrix inversion. With these three modifications, the new biconcave shape bond model is created.To assess the validity of the proposed model, this study compared the elastic deformation in a single bond with an elastic analysis performed using the finite element method (FEM). After verification of the proposed model, a series of sensitivity analyses was performed to explore the effect of bond shape and the differences from the parallel bond model. For a single bond, the results showed that a biconcave model can represent the mechanical behavior of a biconcave shape bond and is effective with different bond shapes. The stress fields calculated by the biconcave model were more similar to the FEM results than those obtained using Dvorkin’s theory. The width and thickness of a bond have a considerable effect on its stiffness, and the variation of thickness decreases with the width of the bond. In addition, the required parameters can be acquired directly from real bond material rather than using the back analysis procedure for parameters.The biconcave model was further implemented in DEM software Particle Flow Code 2 Dimension to certify its applicability in geotechnical engineering. Using this model, a series of uniaxial compression test simulations was performed to analyze the macroelastic responses of specimens. The results showed that the macrotangential elastic modulus and Poisson’s ratio had a regular tendency, and they agreed with the stiffness variation in a microbond. This finding can assist in determining correlations between microand macroproperties and provides an innovative method for simulating the behavior of cemented granular material."

Jan 1, 2015

By J. Fixter, P. Prangnell, Z. Pinter, E. Eimer, B. Chehab

"High deposition rate Wire and Arc Additive Manufacture (WAAM) can provide benefits to the aerospace industry, including a reduction in lead time, reduced material waste and, through higher solidification rates, refinement of the as-built microstructure relative to a cast material. However, candidate 2xxx series high strength alloys for this process normally require a solution treatment, which is problematic for large near-net-shape components, and age hardening to achieve maximum mechanical properties. With each layer being subject to multiple thermal cycles, exploiting the thermal conditions in the WAAM process to avoid post-build solution treatment is therefore of interest. A solutionisation model has been developed for the WAAM process, based on diffusion controlled dissolution using DICTRA, which was validated through thermal simulations combined with 2D and 3D image analysis. The model has subsequently been used to simulate the solutionisation behavior during a typical WAAM thermal cycle and to explore what could theoretically be achieved by, for example, refining the starting microstructure.INTRODUCTION There is currently great interest in exploiting additive manufacturing (AM) in the aerospace sector due to several benefits it can potentially provide; including reduced material waste, shorter product lead times and greater design freedom (Herzog, Seyda, Wycisk, & Emmelmann, 2016). However, with aluminium alloys the material properties obtained are generally inferior to those of conventional high-strength wrought products; therefore more research is required to explore the potential for exploiting the unique process conditions in AM to improve the as-deposited microstructure. A range of AM techniques are now available (Debroy et al., 2018), which differ in terms of the feedstock type and power source and it is necessary to match the appropriate technology to a particular application. This generally requires a compromise between the build rate and shape resolution. Wire and plasma-arc additive manufacturing involves using an electric arc to melt wire onto a substrate, and fabricates a part by building up overlapping layers of deposited weld beads (Addison et al., 2015; Williams et al., 2015). The benefits of this process over other AM methods are the much higher deposition rates (e.g. kg’s/hour) achievable and lower capital equipment costs relative to powder bed techniques, which makes it suitable for large scale applications. However, limitations include a lower part complexity, lower solidification rates and a greater need for post-process machining (Frazier, 2014)."

Jan 1, 2018

By A. L. Mackie

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

Aug 1, 2013

By C. Matthews, A. Vo

Australia is one of the world’s leading exporters of bulk commodities and the extensive use of sampling equipment is key to the successful management of each mining operation. At the process plant, without a full understanding of the feed, product and tails, it is impossible to fully optimise the performance of the plant, such that the metallurgical balance is often open to error. At the port, a statistically correct, bias free and compliant sampling system is crucial to ensuring exported minerals are reflective of their true value and to avoid penalties. Over the last decade, as a result of the iron ore boom, there has been high demand for reliable, representative and standard compliant solids sample systems in Australia. This paper reviews several solids sampling applications at Australian mine sites, the challenges and experience gained and how the implementation of reliable representative sampling systems has provided better understanding of plant process performance, improved plant control and overall benefits to the operations. INTRODUCTION Mining is the primary contributor to the Australian economy with Australia ranking in the top 4 in economic resources for 21 primary industrial minerals in 2016 (Britt, et al., 2017). Australia is one of the world leaders in the export of the majority of its commodities including gold, copper and iron ore (Britt, et al., 2017). Therefore, a suitable and standard compliant sampling system is essential for evaluating plant performance and content verification. At the mine and at the port, the larger iron ore miners in Australia have developed stockpile models to predict the grade of their ore down to individual blocks of the stockpile. This allows the miner to create the right blend of material for sale to their customers and ensures they are not underselling their product or at risk of penalties for impurities or low grade. As an example, the average difference between the benchmark 62% Fe and low grade 58% Fe fines at the time of this report averaged US $20/tonne according to the Metal Bulletin Iron Ore Index (Fastmarkets MB, 2019). At a shipload of 180kt this would be a variation of US $3.6M. Whilst this is an extreme example, it highlights the economic importance of producing the correct grade product and a recognised representative sample of said product.

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

By Paul E. Mather, Matthew J. M. Krane

Anode or fire refining is the final pyrometallurgical step in the production of copper anode. Part of this process purifies liquid copper by reducing its dissolved oxygen with a natural gas-steam mixture. Until now there have been no transient models of this process coupling the injected gas jet mass transport with its chemical reactions. A transient, 1D flow, chemical reaction model was made to estimate the time-dependent behavior of anode copper composition, soot formation, and the transient gas jet composition along its height. Model estimates were made for reduction rate, soot emission rate, and exergy loss and destruction distributions for two melt temperatures and a range of steam addition rates. It was found that increases to melt temperature and steam addition rate increased reduction rate and decreased soot emission, while having little effect on exergetic performance. Estimates for reduction histories were compared to industrial data. INTRODUCTION Anode/fire refining is the final pyrometallurgical step in the production of copper anodes; it is essential for removing dissolved gases which cause porosity in cast anodes. Despite its significance, the open literature appears devoid of any transient models of the entire process; furthermore, there are very little published process data with which to validate any models made. This dual dearth of modelling work and process data inhibits studies of the tradeoffs between process and performance changes, desirable for decreasing reduction time and soot formation as well as quantifying changes in the exergy losses and destructions of the process. This work introduces a reduced-order, transient model of reduction during fire refining, examines performance trends with operating conditions, and compares model estimates with industrial data. Overview of anode refining The purpose of anode refining is to produce liquid copper which can be cast with minimal gas porosity, which disrupts the downstream electrorefining process. Dissolved sulfur is removed first, then oxygen, by reactions with different injected gases. The removal of oxygen (the reduction step) is the subject of this proceeding and is effected by the bubbling of natural gas, typically mixed with steam, through the melt from tuyeres. A typical refining run reduces dissolved oxygen from 3000 ppm to 300 ppm in 2–3 hours (Schlesinger, King, Sole, & Davenport, 2011). The refined copper is then cast as anodes. Figure 1 shows a schematic of the furnace vessel.

Jan 1, 2019

By Xiaoqi Sun

"Rare earth elements (REEs) have a series of specific properties making them invaluable in products including lighting, electronics and magnetism. The elements have been used in many commercial products, for example, turbines, electric vehicles, batteries, lighting, computers, cell phones and lasers. Heavy REEs are attracting the interests of scientists from many countries, and these elements have become critical to develop high-tech materials. Ion-adsorbed rare earth deposits were first discovered in China. The minerals are also called weathered crust elution-deposited rare earths. The particular types of deposits are distributed throughout southern China. The ion-adsorbed deposits are rich in medium and heavy REEs, accounting for more than 80% of the world’s total medium and heavy REEs production. However, the separation of heavy REEs from ion-adsorbed deposit is difficult owing to their similar chemical properties. The conventional separating technologies have caused serious pollution. A novel extraction strategy has been developed in this laboratory and is presented herein.INTRODUCTIONThe unique properties of rare earth elements (REEs) render them important in high-strength magnets, lighting phosphors, polishing compounds, and ceramics. Especially for applications that take advantage of the f-orbital electrons, substitutions outside the REE family are often difficult to impossible (Nassar, Du, & Graedel, 2015). Extracting and separating individual elements from mined ore is not easy, requiring lengthy operations to separate them, and this results in high costs and adverse effects on the environment. The low concentration of REE in the earth’s crust makes economic exploitation difficult (de Boer & Lammertsma, 2013) and the sustainability of the REEs separation industry has aroused considerable attention (McLellan, Corder, Golev, & Ali, 2014). To eliminate protons and break the dimmers in organic phosphonic acids or carboxylic acids, the saponification of acidic extractants by acid-base neutralization reactions are widely used in industrial REEs separation (Sun, Luo, & Dai, 2012a). The saponification procedure contributes to increased extractability and selectivity of acidic extractants to a considerable extent. The common industrial saponification mechanisms and technological processes are shown in Figure 1 and include ammonium saponification (saponified by NH3·H2O), sodium saponification (saponified by NaOH), calcium saponification (saponified by Ca(OH)2), and magnesium saponification (saponified by MgO). These inevitably result in wastewaters containing NH4 +, Na+, Ca2+ or Mg2+ because of the cation-exchange extraction mechanisms (Xiao, Long, Huang, Feng, Cui, & Wang, 2013; Song, Fu, Wang, Liu, Zhou, & Li, 2014; Sun & Waters, 2014)."

Jan 1, 2016

By Masashige Makifuchi, Hiroki Kumakura, Kota Nagai, Takahiro Kikuchi

"A large amount of platinum group metals (PGM) are used in automobile catalyst. It is important to recover the metals from wasted catalyst. Nomura and Kageyama (2013) suggested recovery process that PGM oxides were absorbed to perovskite-type oxide based on LaScO3. In their process, the PGM in wasted catalyst as absorbed to the perovskite-type oxide as PGM oxide. It has potential to each element in the process if the perovskite-type oxide is selected properly. In this study, the PGM oxide absorption properties of various perovskite-type oxides are investigated to develop a new recovery and separation process. The various perovskite-type oxide were synthesized with metallic platinum, palladium, rhodium at high temperature in air atmosphere. And the PGM contents in the perovskite-type oxide after the experiment are determined by chemical analysis. Some of them have the properties to absorb specific PGM oxide selectively. In addition, in order to increase the amount of absorption of PGM oxide and composition of the perovskite-type oxide experimental temperature was investigated. Then it is also found that doping another specific element to the perovskite-type oxide is effective to increase the amount of absorption of PGM oxide. Then a new recovery process of PGM from wasted automobile catalyst, in which PGM can be separate to each element, was designed.INTRODUCTIONA large amount of Platinum (Pt), Palladium (Pd), and Rhodium (Rh) are used in automobile catalysis (Figure 1) (Platinum, 2013). Natural supplies for these metals are limited and PGM mining can detrimental environmental impact. Therefore, the recycling of PGMs from wasted automobile catalysis is very important. The Rose process (Figure 2) was developed as a process to recover PGM from wasted automobile catalysis. In this process, the automobile catalysis is first crushed mechanically, and then extracting agent, copper (Cu), and flux are melted at high temperature. The mixture separates into a metal phase, (Cu alloy containing PGM) and slag. The PGM is enriched by oxidizing. Cu to Cu oxide. The PGM are then recovered and separated to each element in a wet process. However, it is hard to dissolve the PGM with acid because they are chemically stable and strong oxidation agent such as aqua regia is used. Moreover, to separate the PGM to each individual element, a solvent extraction process is required."

Jan 1, 2016

By M. A. Roche

The property of the Hudson Bay Mining & Smelting Company, Limited is located at Flin Flon lake, 85 miles northwest of The Pas, Manitoba, on the Manitoba-Saskatchewan boundary. The area is at latitude North 54°, and is approximately 1,000 feet above sea level. The ore-body occurs in amygdaloidal greenstones, which are to be referred to the earliest basic volcanic flows in the district. It was discovered in an outcrop on a point of land jutting into Flin Flon lake, and subsequent diamond drilling proved that it extends under the lake both to the north and south of the point of discovery. After the taking of soundings from the ice during the winter months, it was estimated that l,000,000 tons of overburden, mainly mud and clay, would have to be removed to expose the underlying ore. Flat diamond-drill holes indicated that the ore-body was separated into more or less independent lenses by horses of unmineralized greenstone, and that in places the overall width, including horses of waste, reached 450 feet. These several] conditions were the main factors influencing the management in their decision to adopt an open-cut method for the mining of the upper portion of the ore-body. Before it was possible to commence stripping operations, it was necessary to dam-off the eastern portion of Flin Flon lake, instal pumps, and de-water the dammed-off area. Dams for this purpose were constructed of mine waste rock and clay filler from nearby hillsides. Material for the bottoms of the dams was taken across Flin Flon lake in hopper-bottom barges. Trestles were then built on the dam bottoms and dams were complete by the time trackage was available. De-watering commenced in April, 1930, and was completed by the latter part of May of the same year. Estimated pump output for the period was roughly four billion gallons. Two Morris 10-inch dredging pumps were used, direct-connected to 250 h.p. electric motors. Individual pump capacity was rated at 3,000 g.p.m. against a 150-foot head. Average head for the operation was 50 feet, which increased the total output considerably. The pumps were mounted on floating barges in cribbings of piling driven at the lowest points in the lake bottom.

Jan 1, 1933

By T. Thriveni, Ji Whan Ahn

"The current advanced technologies for the manufacture of light weight metals processing significantly increasing and attributes to the safety, performance, recyclability and economy. Still lack of technologies developed for light weight metals processing waste utilization and application. Environmentally friendly methods need to develop for better understanding the processing waste utilization. Alumina and aluminium processing creates a wide range of waste products, the most significant being is red mud. Bauxite residue is a high volume hazardous waste and nearly 3 billion tons have been dumped worldwide. Approximately, 120 Million tons/yr of red mud being produced without any environmental sustainable solution. Currently the most significant technology is waste utilization. The bauxite residue is suitable for road base construction, golf bunker sand, back fill and, general fill. Sustainable recycling technologies are necessary for the bauxite residue utilization for the recovery of rare earth elements and cement materials. Carbonation or CO2 sequestration of carbon dioxide is a method pioneered for the bauxite residue is modified to reduce alkalinity. The carbonation of alkaline material is an inexpensive and safe process that leads to the formation of thermodynamically stable products. The use of the carbonation can be an advantageous solution for environmental climate change and sustainability problems in the new century.INTRODUCTIONIn 2014, bauxite consumption was estimated to be 9.6 million tons, nearly all of which was imported, worth an estimated 269 million dollars. More than 95% of the bauxite was converted to alumina, and the remaining 5% went to non-metallurgical products, such as abrasives, chemicals, proppants, and refractories. Global aluminum demand is growing steadily spurred by growth in India and China (RBCAluminum Market Outlook, 2013) (Figure 1). Aluminum consumption is expected to rise 5% annually through the end of the decade, reaching 67 million tons per year by 2020. Global bauxite production decreased by 17% from that of 2013, principally owing to lower production in Indonesia. In response to a Government ban on exporting bauxite and other unprocessed mineral ores that took effect on January 12, 2014, bauxite mines throughout Indonesia shut down and production decreased from 55.7 million tons in 2013 to an estimated 500,000 tons in 2014. The ban was part of a mining law that was instituted in 2009 to increase development of downstream processing facilities in Indonesia."

Jan 1, 2016

By A. M. Bell

"The property of Gaspe Copper Mines, Limited, is in the Gaspe Peninsula some 60 miles west of the town of Gaspe and 20 miles in from the north coast. The Copper Mountain deposit was first discovered in 1921 and first optioned and drilled by Noranda in 1938. The Upper Needle Mountain showings a mile south were found in 1938, but it was not until after a delay of 10 years that the more important deep ore at Needle Mountain was located by drilling. Production started in 1955, and the smelter commenced operation in December. Mining is by open stope methods with first production coming from the Lower Needle Mountain orebody, of which only a small portion has been exposed in underground openings.Stratigraphy and AlterationThe country rock is composed of limy siltstones of Devonian age. Away from alteration these are dark and generally carbonaceous. Throughout the drilled area numerous narrow layers of tuffaceous fragmental material occur, and such beds are found to provide excellent horizon markers by virtue of their uniform thickness and wide lateral distribution. These markers are valuable for stope control in mining the bedded replacement deposits.Intruding the sediments at Copper Mountain are porphyry dykes, and one drill hole encountered granite 1,400 feet below the valley. Solutions responsible for the rock alteration and ore deposits seem to have emanated from this intrusive centre.On the mine claims strong hydrothermal alteration has completely obscured the original character of the sediments. The unaltered formations are all very similar in appearance, but following alteration rocks that are very dissimilar in physical characteristics are produced, and such differences have an important influence on ore deposition. The alteration effects are known to extend over a vertical range in excess of 3,000 feet and have a surface area of two square miles. All known ore bodies are within the zone of alteration. In this, rocks are strikingly bleached to a white colour in contrast to the normal dark-coloured sediments, and are much harder and more abrasive, with bedding often obscured. The first effect of alteration is driving off of carbonaceous matter, followed progressively by silicification in the acid rocks and by the formation of lime silicates in the lime rocks, with an accompanying gain in silica and a corresponding loss in lime. Extensive analysis has shown that there is an actual addition of silica involved in the alteration."

Jan 1, 1949

By Robert MacPhail, Louis Racine, Bret G. Cousins

"Sodium Isopropyl Xanthate is employed in a wide variety of sulphide flotation procedures due to the specificity over other xanthates and the cost savings over other sulphide collectors such as thionocarbamates, xanthogen formates and dithiophosphates. An investigation carried out by Tennants Research Technology Applications (TRTA), has produced a chemically enhanced SIPX that has shown all the advantages of normal SIPX along with improved recoveries.Currently, FLEX 31 is manufactured at the Prospec Chemicals xanthate plant in Fort Saskatchewan, Alberta, Canada. Prospec, founded in 1983, produces approximately 4500 tonnes of xanthates per year and is presently undergoing expansion to produce approximately 6000 tonnes. The proprietary process for FLEX 31 is a new addition to plant operations that include the production of xanthates, thionocarbamates and xanthogen formates.This paper will cover the testwork on the zinc flotation circuit at the Louvicourt Mine in Val d'Or, Quebec. This testwork indicated that FLEX 31 could replace SIPX due to the ability of the FLEX to improve zinc flotation kinetics without adversely affecting final concentrate grade. FLEX has presently replaced SIPX in the zinc circuit at Louvicourt.BackgroundThe Louvicourt deposit is hosted by greenschist facies Archean volcanic rocks located in the southeast portion of the Abitibi Greenstone Belt. In the Val d'Or area, the volcanic rocks form an east-west striking, steeply dipping and south-facing sequence composed of thick mafic and ultramafic flows of the Lower Malartic Group.The joint venture has three partners-Aur Louvicourt Inc. (30%), Novicourt Inc. (45%) and Teck Corp (25%). Aur Resources holds 30% interest through Aur Louvicourt, a wholly owned subsidiary and is the operator of the project. Novicourt Inc. is 60.9% held by Noranda.Approximately 75% to 85% of Louvicourt's revenue comes from the sale of copper; 10% from gold (in the copper concentrate); 7% from zinc and 3% from silver (also in the copper concentrate) The copper at Louvicourt is fairly ""fast floating"". About 60% of the copper in the ore can be concentrated in the first flotation cells with feed grades of approximately 4% copper.The initial testwork and pilot trials had indicated that it would be difficult to achieve zinc grades over 52% and recoveries would have to be kept at a level amenable to salable concentrate grades. The actual mill results, with a fine regrind, achieved zinc grades of 55% to 56% at about 75% recovery."

Jan 1, 1996

By A. P. E. Dirige

In order to maximize the recovery of ore in variably dipping ore zones of moderate width, cemented backfill is normally placed to serve as structural support. With the intention of saving on costs, backfill of low cement content can be used, which are often supported by a sillmat of higher strength. The stability of the low cement backfill face, exposed during adjacent mining, must be carefully studied to provide very effective, safe and economic mining operations. Improper design of these stope support structures may result in fill mass failure resulting in relaxation and failure of the stope walls, with consequent losses of production, ore dilution, and in safety problems. This paper presents a study conducted to assess fill performance during adjacent pillar mining and to provide a comprehensive understanding of backfill behaviour, the failure modes that may occur and the consequences to production, ore dilution and to safety problems, and accurately predict their stability. The design study investigated the effects of stope width and height, orebody geometry and inclination, and wall roughness on the stability of cemented backfill during adjacent pillar mining. Fill properties used were based on paste fill specimens cured for 28 days. Paste fill performances were assessed based on analytical and numerical modeling studies for the different mining conditions. Analytical modeling was carried out using limiting equilibrium analysis and numerical modeling was carried out using FLAC3D. The modeling results suggested that, for stopes that are inclined with smooth wall rock conditions, backfill failure, driven by the fill self-weight, has minimum dependency on the binder content and is reduced by resisting forces developed on the footwall-fill contact. For inclined stopes with rough wall rock conditions, wall roughness contributes significantly to the stability of the backfill during adjacent pillar mining. The analytical modeling approach was demonstrated to be useful in providing some approximate parameters for predicting the behaviour of paste fill exposed faces during adjacent mining, but cannot predict the mode or mechanisms of failure. Numerical modeling not only assess the stability behaviour of paste fill free faces, but also is able to provide a better idea of paste fill failure modes and possible failure mechanisms. The depth of failure and the potential for instability for a simulated stope filled with paste fill can be predicted, and may be useful in estimating the mass of material that could possibly fail and ore dilution levels.

May 1, 2009

By W. P. Corking

"The purpose of this paper is to describe the relationship of the goldbearing ore to structure at the Pickle Crow mfoe. Information beyond this scope may be found in the reports and papers of Hurst( l), Bothwell(2), Thomson(3) , and Monette( 4).The property lies on Crow river, in Patricia district, 80 miles northeast of Allanwater, and about 125 miles northeast of Sioux Lookout on the Canadian National railway. From April of 1935, when the mill began operation, to the end of 1944, a total of 564,890 oz. gold and 49,275 oz. silver were recovered from 958,416 tons of ore.Mining on the property is at present carried on in two separate orebodies, the Howell and the No. 2 vein. Stoping has been carried on from every level on the Howell vein to a depth ot 2,450 feet below the surface and from several levels on the No. 2 vein through a mile long cross-cut on the 750-foot level to the Howell vein workings. A steadily increasing volume of ore is being obtained from the No. 2 vein.Each of the two orebodies consists of vein quartz with small amounts of other gangue minerals either in a single vein or a stockwork of veins. Both lie within sheared walls and carry gold in the free state. Fig. 1 shows the position of the veins with respect to the surrounding rocks. It will be noted that the delineation of the beds of iron formation shows that the lava flows and minor, interbedded sediments that make up most of the country rocks are areally folded and contorted.STRUCTURENearly all gold found to date in sufficient quantity and concentration to make ore in the Pickle Crow area is confined to these major quartz veins . Both strike a little north of east and dip north, and in both cases the attitude of the enclosing rocks is somewhat more north of east in strike and more steeply north in dip than that of the veins. As a result of this disposition both veins migrate across their enclosing rocks at a moderate angle in both dip and strike. (In this connection it should be noted that No. 2 vein dips at a considerably flatter angle than its host rock, but has a very pronounced rake to the northeast, effecting the same result as in the Howell vein.)"

Jan 1, 1954

By R. Sharma, I. Miskovic, S. Miskovic

Solvent extraction (SX) is a widely used method for the extraction of copper from pregnant leach solutions or other ionic solutions in hydrometallurgical applications. About 17% of total refined copper production worldwide is produced using SX followed by electrowinning (EW). Conventional SX units, such as mixer-settlers, consume a large amount of extractant, offer poor process control and yield low extraction efficiency. Extraction efficiency is poor due to low specific interface area, crud formation, emulsification, and slow phase separation stage. Micro solvent extraction (micro-SX) is a rapidly developing technology that provides an opportunity for improved process control, higher extraction efficiency, lower residence time and, as such, has a potential to reduce the environmental footprint of SX operation. In this research, extraction of copper ions in an acetate buffer media is investigated using D2EHPA with kerosene as a diluent. Chemical factors such as extractant concentration and ionic concentration of aqueous phase are varied to determine the optimum conditions. Subsequently, the extraction efficiency and reaction time of new micro-SX reactors are compared against previous bulk SX studies. Extraction in the microfluidic reactor was observed to be faster with a residence time more than an order of magnitude lesser than bulk SX. This study forms a prelude for developing new modular micro-SX units which will enable continuous, effective, and safe operation on a large scale. In future studies, numerical and analytical models validated through this work will help eliminate rigorous experimentation and enable faster design and fabrication of micro-SX units. INTRODUCTION Conventional industrial solvent extraction (SX) units such as mixer-settlers, extraction columns present various disadvantages. Low specific interfacial area, crud formation, and emulsification in industrial reactors lead to long extraction times and severely hamper extraction efficiency (Priest et al., 2011; Chen et al., 2017). Microfluidic extraction considerably improves upon the current extraction practices by (1) creating a large surface to volume ratio that enables rapid mass transfer across the interface, (2) better process control, and (3) easier separation of phases (Kashid, Renken, & Kiwi-Minsker, 2011; Ciceri, Perera, & Stevens, 2014). These factors establish micro-SX as a crucial mode for process intensification. Although, a micro SX unit has significantly low processing volume than conventional extraction methods, their throughput can be increased via massive parallelisation (‘numbering up’), which is an area of active research (Darekar, Singh, Mukhopadhyay, & Shenoy, 2016a; Kriel, Woollam, Gordon, Grant, & Priest, 2016).

Jan 1, 2019

By C. Villaseca, F. Penna, R. Villarroel

Acid mist is the undesired result of copper electro-winning (EW) process, generating a contaminated and hazardous environment in the working areas, as well as promoting high corrosion conditions for structures and equipment in the EW tank-house due to the buoyancy of sulfuric acid in aerosol form. Likewise, organic entrainments that occasionally occur in this process contaminate the cathodes, as well as increasing the risk of fire. LateralVex® and LateralFlow® are technologies that effectively and efficiently solve these problems, without affecting the usual activities of the operators, thus presenting a significant distinction with respect to other technologies that seek to solve these problems. Minera Antucoya, an AMSA group company, which produces 80,000 tons of electrolytic Cu per year, decided to install both technologies for its EW tank-house, LateralVex® in commercial cells and LateralFlow® in scavenger cells to keep both the acid mist and the organic entrainments under control. This paper presents the effects of applying these technologies in Minera Antucoya’s electro-winning plant, the challenges addressed, the difficulties encountered, the problems and opportunities that their implementation involved, as well as the surprising results obtained at the time of measuring the acid mist— less than one hundred’s part of the Allowed Weighted Limit—in samples taken with the plant operating at maximum current capacity. Additionally, the paper includes some of the improvements identified during the period, more than 3 years, in which these technologies have been in operation in Minera Antucoya. INTRODUCTION Electrowinning is the most used electro-chemical process in the hydrometallurgical industry for the recovery of highly pure metallic copper as its final product, which is formed in previous stages in an aqueous medium that goes through a selective process using organic solvents to separate impurities and concentrate the metal in a sulfuric acid solution. In the last stage, this solution goes through a process in which high- density current at a low voltage is applied, so that the metal in the solution will deposit in solid state on a plate, usually a stainless-steel plate. This last process involves the generation of other sub-products due to the electro-chemical dissociation of the electrolyte’s components, or to the addition of chemical additives, or due to specific physical conditions or lack of controls during the operation.

Jan 1, 2019

By Scott D. Long

"Abstract-Grass roots exploration has an emphasis on good precision at low (sub-economic) concentrations of an element, in order to identify "anomalies" which may lead to drilling targets. Although it is wise to have well-defined quality controls in place during early stage sampling, the control procedures at this stage tend to be rudimentary. When a deposit warrants detailed delineation, the need for accurate geological characterization and assaying of ore-grade rock must be supported by appropriate quality control procedures.As a project returns more promising results, and the likelihood of the project warranting a prefeasibility study increases, it is time to tighten quality assurance requirements and quality control protocols. Delays in instituting adequate quality control and verification procedures may lead to costly remedial actions that can also delay completion of the prefeasibility or feasibility study. Sometimes, as a project matures, the quality control and quality assurance aspects fail to keep pace with the changing nature of the work: an increasing emphasis on precise and accurate assaying of oregrade rock.A resource model is like a house of cards. The foundation of the house is the geological observations, the sampling methods and the geological interpretations performed upon those samples. Upon the sampling, which includes all the sub-sampling steps of sample preparation, rest the assay results. The quality of the assays can be no better than the quality of the sampling. Good assaying technique will preserve all biases and imprecisions introduced during sampling and sample preparation; it cannot remove them. If the sub-samples and/or assays are bad and must be redone, all the succeeding levels of the house of cards must also be rebuilt, or at least patched up.From geological observations, a geological interpretation is built, the quality of which is critical to the construction of an acceptable resource model. Geological elements are difficult to assess, and are not discussed here; they are dependent upon the expertise of the geologists involved. Geological knowledge contributes essential quality control elements to resource/reserve determinations (Sinclair and Postolski, 1999; Postolski and Sinclair, 1999).This paper consists of five sections: (1) overview of quality control protocols, (2) maintaining quality control of databases, (3) quality control of drill sampling, (4) optimizing crushing and grinding requirements, and (5) quality control of assay data acquisition."

Jan 1, 1998

By Ross Gilders, Leo Cheung, C. David Desjardins

"For the past twenty years significant progress has been made in bacterial leaching. In fact, a recent article in the Northern Miner has cited that nearly all new copper production is based on bioleaching. The bacteria that play the predominant role in bioleaching is an obligate chemolithoautotroph called thiobacillus ferrooxidans. These bacteria thrive in areas rich in sulphide and prefer low acid media. In the simplest sense, they metabolize sulphide converting it to sulphate either as acid or as soluble metal sulphates. This process is well known because these bacteria generate acid mine water. Over the last five( 5) years, we have been carrying out development work on creating optimum bioleach conditions for the extraction of metals from sulphide rich ores and concentrates. Recently, a study was carried out on a byproduct sulphide concentrate produced in a tin flowsheet developed for the Mt. Pleasant Mine of ADEX Mining Corp. The objective was to determine whether a bioleach process might successfully leach this sulphide material which contains copper, zinc, iron and, what is more important, trace metals such as Indium, Gallium, and Bismuth. The current work describes the successful recovery of these metals achieving extractions of more than 80% by applying bioleaching techniques. An overview of this technical program forms the basis of this paper.IntroductionBio-Oxidation of metal sulphides has emerged to be one of the leading edge hydrometallurgical processes for the economic extraction of metals from ore and concentrates. The fundamental process involves the bacterial assisted oxidation of sulphide to sulphate yielding metal sulphates which can be subsequently leached by conventional means. The bacteria that play the predominate role is an obligate chemolithoautotroph named thiobacillus ferrooxidans (TF). These bacteria thrive in areas rich in sulphide and occur naturally worldwide. In fact, TF is responsible for generating acid in water (acid mine water) from exposed sulphide surfaces primarily through the oxidation of pyrite. TF bacteria thrive under the proper conditions of temperature and oxygen supply in the presence of sulphides and populations in excess of a million per millilitre (ml) of solution are achievable. Typically, at room temperature, the bacteria will bio-oxidize in a pH range of 1.0 to 3.0. Today, bio-oxidation has been applied commercially with great success for the extraction of uranium, copper, and gold."

Jan 1, 1996