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By J. X. Guo, A. Pindred, J. Moore

"Two modified centrifugal flotation cells (CFC), Q2 and Q3 models, have been developed for froth flotation. The operating mechanism of Q2 uses a rotating bowl to create a centrifugal force on the feeding slurry. Froth concentrate is collected on the central top of the cell and tailing is discharged along the top edge of the rotating bowl. The Q3 utilizes a vertical rotating feed pipe to introduce the feed slurry into a stationary flotation column cell. A centrifugal force is also created by a rotating pipe for the feeding slurry. Froth concentrate floats over the cell and tailing discharges through the bottom of the cell. The centrifugal force causes the flotation recovery of fine particles from CFC to be higher than from the conventional cell. The major operating variables of CFC are the rotating speed, air flow rate into feed slurry, and slurry feeding rate etc. Flotation tests have been carried out on synthetic mixtures of galena-silica and chalcopyrite-silica, as well as some natural base metal ore samples. Parallel tests were also conducted on a conventional flotation cell. The comparative test results demonstrated that both Q2 and Q3 have the advantage of higher recovery of finer than 10 µm mineral particles, which are very difficult for the conventional non-centrifugal flotation cell to recover. The grade-recovery curves produced from CFC are equivalent to or better than the conventional flotation cell. There is a potential in the future for the CFC to be employed in slime flotation.1. IntroductionA centrifugal flotation cell (CFC) was originally developed by Clean Earth Technologies [I] to separate oil from water and was subsequently applied in froth flotation. Figure 1 illustrates the design of the original CFC, in which a vertical feeding pipe is attached to the center part of a rotating drum and a stationary shell encloses it to direct the float froth and tailing discharge. The slurry flows down to the bottom of the spinning drum, gets in touch with a screen of air bubbles and then migrates outwards under the effect of centrifugal force. After the mineralized froth enters the rotating cell, it is pushed back to the inner surface. The overflowing froth is collected in the froth launder while the tailing slurry is discharged through orifices along the outer rim of the drum."

Jan 1, 1998

By Thomas J. Ferree

"There has been a major evolution of mineral processing technology in the past decade. The evolution has been driven by recognition of broad environmental changes which have impacted various areas of the world. Numerous government regulations have required many mining and mineral processing organizations to introduce changes to reduce the undesirable effects on the environment; many mining companies have voluntarily introduced new technologies and practices which surpass the limits imposed by the regulations. Surprisingly, some of the 'new' technologies are modifications of concepts developed decades earlier. Gravity separation technology is in this category.Mineral Deposits Limited, an Australian company, has been a world leader in the development of mining concepts, reclamation practices and equipment technology that is environmentally suited to today's requirements. This paper relates to equipment used for gravity concentration and electrostatic separation systems, including Reichert cones, spirals, and electrostatic separators. The basic devices and technologies were developed more than forty-five years ago, but use of the systems was quite limited until environmental and conservation practices were recognized and introduced. Spiral concentrators, for example, were patented in the mid-l940s by I.B. Humphreys in Denver , Colorado. The two types of spirals marketed (the Model 24A for minerals and the Model 24C for coal) remained essentially unchanged until expiration of the patents. During this early period most spirals were used in iron-ore wash plants on the Mesabi Range to recover fine iron which previously had been discarded. About the same time sand mining plants in Florida incorporated spirals to recover titanium minerals and zircon from lowgrade beach sands. The Labrador Trough opened new opportunities for spirals when hard-rock iron-ore deposits were developed; approximately 10 000 Model 24A Humphreys spirals were installed between 1958 and 1966. About 5000 spirals of this model are still operating in the same capacity, in spite of the fact that extensive improvements have since been made in gravity separation technology."

Jan 1, 1993

By Douglas H. Silvey

"The historical transformations in fire -resistant conveyor belt compositions in the past 20 years have been influenced by factors of material cost, changing flame and burn test requirements, material replacements, adhesion and oil resistance considerations, increased responsibility and legal liability. Specifically, this has resulted in a swing from neoprene-cotton belts to S.B.R. blends and synthetic fibre carcasses, and tighter quality control and documentation of test data.Flame dynamics are discussed and the function and performance of the multiple types of flame-resistant additives in the flame front are presented. Specific flame and burn test requirements are described. The belt manufacturers' requirements and criteria in selecting the types of flame-resistant additives are presented.The author will describe the composition of the types of flame-resistant compounds used in a practical production, the belts that are available on today's mark et and future developments in requirements that may further transform the composition of fire-resistant belts.IntroductionThe historical transformations of conveyor belt compositions in the past two decades have been based on factors common to those leading to changes in man y organic compositional products. These factors are material cost, new materials, the customer' s changing requirements, government regulation s and increased manufacturer' s liability.As a basis for understanding the product, I will provide a brief breakdown of the conveyor belt composite. Figure 1 represents a cross section of a typical belt. A skim rubber coats each fabric ply of the belt carcass and is of a different rubber composition than the outside layer of rubber which forms the cover. This outside cover layer is designed to protect the belt from environmental factors such as heat, oil and abrasion, and other factors involved in the service condition s under which the belt will operate. In the case of fire-retardant belt s, both the skim coat encapsulating the fabric plies and the outer cover must contain sufficient fire retardancy to overcome the contribution to burning of the decomosition products of the fabric."

Jan 1, 1982

By Jocelyn Koshinsky

The uranium milling process at the Key Lake Operation of Cameco in Northern Saskatchewan, Canada results in the leaching of untargeted metal(loid)s such as arsenic, iron and nickel that are found at high concentrations in the ores. To mitigate the environmental impact of the leached As, Fe and Ni in the mill waste solution (raffinate), the pre-discharge mine tailings (rafffinate and residue) are neutralized with slaked lime, resulting in the formation of secondary As-, Fe- and Ni-bearing mineral phases which serve to control their concentrations to very low levels in the tailings porewater. As part of an ongoing investigation into the long-term evolution (10,000 year) of arsenic mineralogy in the Deilmann Tailings Management Facility (DTMF) at Cameco?s Key Lake Operation, it has become necessary to characterize secondary As-bearing mineral phases during the neutralization process prior to the discharge of the tailings and compare these to in situ As-bearing mineral phases in the tailings facility. This is primarily to determine if changes have occurred over time to As-bearing mineral phases in the tailings facility as well as to investigate whether alternate neutralization techniques may have a beneficial effect on the stability of mineral phases formed. Thus a bench-scale model was developed and preliminary testing conducted to simulate the mill neutralization process (i.e. neutralization of synthetic raffinate solutions with incremental amount of slaked lime to terminal pH values). The resulting ?tailings? (precipitates and supernatant) were then separated and analyzed. XRD analysis of the precipitates shows that the Fe occurred mostly as ferrihydrite. This was complimented with results from Raman spectroscopy which further indicated that the As appeared adsorbed onto the ferrihydrite. Results of geochemical modeling of the bench-scale test compared well with supernatant profiles from ICP-MS analysis.

Aug 1, 2010

By J. M. Cormie

"The orebodies of the East Malarcic mine are interesting from a geological standpoint because of the very small amount of vein quartz, the occurrence of much of the ore in soft greenstone schist, the presence of innumerable bodies of porphyry and diorite, and their obvious relationship to structural features .The property lies in the northeast part of Fourniere township in western Quebec, and within the important fault zone that geological mapping in recent years has shown is traceable wesrward through the Noranda and Kirkland Lake camps and eastward to beyond the Val d'Or region . The claims included in the property were staked by J.P. Norrie and L. K. Coffin in 1932, but gold was not discovered until i934. Production began late in 1938. The total amount of ore milled up to the end of 1945 was 2,837,091 tons from which gold having the value of $16,631,758 was recovered.The region is underlain by an alternating series of Early Precambrian sediments and volcanics cut by innumerable intrusives. These are more fully described in the works of Gunning, Ambrose, and Norman. The East Malanie property is underlain by three formations, which, named from north· to south, are : the Cadillac sediments, the Blake River volcanics, and the Kewagama sediments. In addition, there are countless intrusions of porphyry and diorite, particularly in the greenstone, occurring in all sizes up to those measurable in hundreds and even thousands of feet . These constitute a large proportion of the greenstone in places. In many parts of the ore zone there is more porphyry and diorite than schist. The greenstone itself is not homogeneous but includes peridotite, andesitic lava, and soft talcose schist.Orebodies have been found in both the Blake River volcanics and the Kewagama sediments. They are not known to occur in the Cadillac group. The ore occurs in a zone of pyrite mineralization in which almost no other metallic minerals and very little vein quartz are present. Generally, the grade varies directly with the amount of pyrite. The mineralization is later than all the rocks of the area. Diorite was a particularly favourable rock for mineralization and contains the highest grade of ore, but any rock in the fault zone may be ore-bearing, whereas the same type away from the zone is not."

Jan 1, 1954

By J. Brent Hiskey

At present, approximately 80 % of the worlds mine production of copper is achieved by smelting and electrorefining. The principle aim of copper electrorefining is to satisfy the large demand for extremely pure copper. It is remarkable that high quality copper cathodes with the required purity (99.99+% Cu, ASTM B115-10) and structural characteristics (i.e. deposit density and morphological texture) are produced from impure copper anodes dissolved in electrolytes containing a variety of impurities. This is accomplished with the proper control of operating parameters such as: current density; temperature; copper and sulfuric acid concentration; agitation; substrate preparation; and the application of electrolyte additives. The basic elements of the electrolysis of copper have not changed significantly over the past 150 years since the early developments of James Elkington. Importantly, our fundamental understanding of the underlying electrochemistry has advanced considerably. This is especially true regarding our knowledge of the mechanistic role of additives during the electrocrystallization of copper. This is a very complex subject and modern electrochemical techniques and the development of surface characterization and imaging instruments has helped immensely in our ultimate understanding of how additives control the deposit structure and morphology. The purpose of this paper is to review the historical development and application of electrolyte additives in copper electrorefining and to examine our current knowledge regarding their use. INTRODUCTION The first transatlantic telegraph cable was completed in August 1858. Regrettably, this cable failed shortly after the initial connections. After several more failed attempts, improvements in materials and copper quality resulted in successful communication between Europe and America on July 28, 1866. In many ways this hallmark event demonstrated the need for high purity copper. The first copper refinery in the world was commissioned in May 1869 in Pembrey, South Wales by partners Josiah Mason and James Balleney Elkington. (Historical Note). The 1865 and 1868 British patents issued to James Balleney Elkington were the foundation for the first commercial production of high purity electrolytic copper (Elkington, 1865, 1869, 1870). This important achievement was made possible primarily by the invention of the dynamo which was based on the fundamental operating principles advanced by Michael Faraday. A symbiotic relationship between copper refining and the dynamo was henceforth established. The dynamo provided electric current for electrolytic refining of copper and the electrorefining tankhouse provided pure copper which was required for the generation and transmission of electricity.

Jan 1, 2019

By Amyn S. Dahya, David J. King

"The object of the paper is to discuss the developments in Carbon-In-Pulp (C.I.P.) technology for gold recovery. The CIP process is comprised of four major operations:(a) adsorption of precious metals onto activated carbon;(b) elution (stripping) of precious metals from the loaded carbon;(c) electrowinning of the precious metals from the concentrated eluate;(d) regeneration of the eluted carbon before recycle to the adsorption process.A detailed review of the technological developments in each of the above operations is presented together with a brief theoretical background of the underlying processes. The advantages and disadvantages of the CIP process are discussed in comparison with conventional zinc precipitation. IntroductionThe carbon-in-pulp (CIP) process involves the use of activated carbon for treatment of pulps containing dissolved precious metals, and provide s an alternative to the conventional gold recover y processes using filtration or counter-current decantation in conjunction with the Merrill-Crowe zinc precipitation process. The commercial application o f the CIP process in gold processing technology is a fairly recent development and it is of growing importance around the world.The fact that activated carbon ha s the property of adsorbing precious metals from solution was first published in 1847. In the following years , a great deal of work was carried out to develop an over-all recover y process using carbon, but all attempts were uneconomical because the loaded carbon had to be smelted to recover the gold.It was not until 1950 that Zadra, at the United States Bureau of Mines Research Centre in Reno, Nevada, developed a technique for stripping the loaded carbon using sodium sulphite-sodium hydroxide mixtures and then recovering the gold from the stripping solution by electrolysis. However, this process was not effective for stripping silver and further testwork was carried out on alternative stripping techniques. In 1952, Zadra published a paper describing the use of a hot sodium hydroxide- sodium cyanide stripping technique, which was effective in stripping both gold and silver. This paper included a description and flowsheet for the CIP process which is essentially the same as we know it today. This technique was used in the first major CIP plant, constructed and commissioned in 1973 by Hornestake Gold Mining Company. Since then there have been numerous CIP plants built in the U.S A., South Africa and Australia."

Jan 1, 1983

By R. M. Giegerich

"This paper begins with a brief history of Madawaska Mines, a brief description of the geology, mining methods and equipment, and the historical development of 7/8-in. tapered drilling equipment. It then outlines the performance characteristics and inherent problems of the traditional 12-degree, tapered connection, blade-type insert bits. Service-free (Unipass) button-type insert bits are presented as a possible solution to these problems. The paper concludes with an analysis of button bit performance to date.Brief History of Madawaska MinesMadawaska Mines Limited, near Bancroft, Ontario, is in Faraday Township, Hastings County, approximately 100 miles northeast of Toronto and 100 miles west of Ottawa.The mine re-opened in 1976committed to 1,500 tons per day to provide approximately 600,000 pounds of yellowcake (U3O8) per year to A.G.I.P. (an Italian government agency) under the joint ownership of Federal Resources Corporation, Salt Lake City (51%), and Consolidated Canadian Faraday Limited (49%). At present, the mine is producing approximately 2 pounds per ton (.1% U3O8) at the rate of 60,000 pounds per month. The mine employs about 400 employees.Brief Description of Geology, Mining Methods and EquipmentGeologyThe Grenville metasediments of the mine area are chiefly paragneiss, amphobolite and some carbonate (marble) rocks, which were intruded by a basic rock, altered to a metagabbrometadiorite. These formations are generally concordant, striking east to northeast and dipping 45 to 60 degrees to the south. Locally, the metagabbro is often discordant. The above group of rocks (mixed hybrid gneisses) forms the hanging wall of the south-dipping Faraday granite sheet and was intruded by nepheline syenite, syenite and granite. The ore occurs in the granite-syenite-pegmatite dykes that are probably genetically related to the Faraday granite. The ore shoots are erratically distributed within the pegmatite bodies, mainly on the hanging or footwalls but also medially. Their size varies from 50 to 350 feet in length, 5 to 40 feet in width and 50 to 200 feet in height. The ore occurs most frequently in a medium-grained granite or syenite, well hematized with 10 to 30% mafic crystals of pyroxene, amphibole and/or relics of the host rock."

Jan 1, 1982

By B. R. Schneiders

"Atikokan is 200 km west of Thunder Bay and earned its mining fame from the discovery of the Steep Rock Iron Range in the late 1930s. Gold, however, was discovered in the Atikokan and Mine Centre areas in the late 1800s and eventually led to one of Ontario's first gold camps, and a total production of approximately 52,000 oz of gold and 174,000 oz of silver from 218,000 tons milled.The general geology of the area consists of Early Precambrian rocks of the Wabigoon and Quetico subprovinces, separated by the east-trending Quetico Fault. The Wabigoon Subprovince, to the north, consists of metavolcanics, minor metasediments and granitic batholiths. The Quetico Subprovince consists of metasediments and granitoids. Small ultramafic plutons occur throughout the metasediments.Within a 50-km radius of Atikokan, more than 50 gold occurrences, prospects and past producers have been discovered since the 18oos. The Atikokan Gold Study performed by the Ontario Geological Sur vey (Wilkinson, 1980, 1982) defined three types of gold mineralization in which gold is concentrated in quartz and quartz-carbonate veins:I) Metavolcanic-Hosted, Stratabound Type-associated with altered felsic fragmental units and chemical sedimentary rocks; possible gold mineralization source rocks. The mineralization is due to metamorphogenic (epigenetic) enrichment. Associated with dynamic metamorphism events and/or felsic intrusive events.2) Marmion Lake Batholith Type-occurs within the gneissic core of the batholith associated with north east-trending lineaments, representing faults or shear zones. Mineralized veins are related to several periods of deformation and shearing, suggesting that metamorphogenic veins were produced during dynamic metamorphism.3) Contact Zone Type -occurs in the contact zone rocks between gneissic batholiths and meta volcanic belts. Hydrothermal solutions produced during several deformational periods acted to leach and concentrate gold mineralization. Localized, discontinuous shear zones host the epigenetic (metamorphogenic) veins.Twenty-one properties in the area were mapped and sampled, and researched petrographically and geochemically. Geological data from the Atikokan Economic Geologist Program of the Ontario Geological Survey are used in discussing the types of mineralization, genetic models, and exploration targets and techiques."

Jan 1, 1983

By D. A. Whitley, R. P. Carone

"Shaft sinking by blind shaft drilling offers the mining industry a different approach to shaft construction. Benefits of blind drilling include less construction time and fewer safety problems while being economically competitive with conventional sinking methods under certain conditions in soft water-bearing formations.Technology and equipment are available today to drill blind shafts up to 20 feet in diameter and over 3,000 feet deep. This is an extension of proven techniques developed by private industry. To drill a 20-foot-diameter hole requires a specially manufactured hydraulic drill unit with a rated hoist of 2 million pounds and 500,000 ft/lb of torque.This brief paper will discuss the following aspects of blind shaft drilling:• benefits;• equipment f or drilling;• circulation methods;• circulation rates;• shaft alignment; and• progress rates to be expected.IntroductionShaft sinking by blind shaft drilling offers the mining industry an improved alternative method for shaft construction. The drilling of 20-foot -diameter shafts to depth s in excess of 3,000 feet can be done by utilizing equipment specifically designed and manufactured for this purpose. This equipment is a distant cousin, at best, to standard oil well drilling rigs. The actual configuration of these drilling machines will be discussed in a following section.To most people, the drilling of mine shafts may seem to be a new or at least recent technology. However, as early as the I920s, German and Dutch miners were drilling 24-foot diameter shafts for coal mines. What led ""miners to resort to drilling were ground conditions that still plague conventional shaft sinkers today: soft , water-bearing formations. The major drawback to these early attempts was a low penetration rate due to poor bit cleaning. This, however, was not the end of early shaft drilling activity. From 1936 to 1950, a number of mine shafts were drilled by the calyx or coring method. This method employs a bit with a large bore down the center. As it drills, it cuts a core. When 5 to 10 feet have been drilled, the core is removed. The disadvantages of this method were the time and work involved in breaking the core free, and it was inadequate in soft water-bearing ground."

Jan 1, 1981

By Howard J. Handewith

"Tunnel boring machines have been tried in underground mining applications as early as 1957. Although tunneling machines hardly threaten to supplant traditional drill-blast-muck operations for drifting, the latest modern machines offer substantial improvements in progress rates over traditional methods under certain conditions.An evaluation guide is offered which identifies conditions that favour machine boring. Different types of costs related to the use of a tunnel machine are discussed. The advantages of the tunnel machine over the traditional method are reviewed. IntroductionThe first tunnel boring machine ever used in a mine was in Quebec in the year 1957. Although less than successful at the time, this experiment served to illustrate the mining industry' s willingness to try new ideas. Since 1957, the manufacture of tunnel boring machines has developed into a full-size industry. Most contractors in Europe and North America will now consider the use of a tunnel boring machine first on any long tunnel that can have a circular section.The modern tunneling machine offers a great potential to the mining industry. Just as the blasthole and raise drill machines have revolutionized mining in the last decade, it is expected that tunnel boring machines will make an equal impact in the future.This is not to suggest that tunnel boring machines will replace explosives in the foreseeable future, but they can supplement drill-and-blast operations. In areas where long drifts, inclines, declines or cross-cuts are needed in a hurry, the tunnel boring machines offer the fastest way to a production return on investment. Figure 1 shows the range of production that two types of tunnel machines have produced as compared with conventional excavation methods in various ground conditions.Rotary hard-rock tunnel boring machines (Fig. 2) commonly produce progress rates in excess of 20 metres per day in uniform ground from 300 to over 2,000 kgl cm2 in unconfined compression strength . These averages are for the completed tunnels. One 7. I6-metre-diameter tunnel boring machine just completed a 1O.7-km tunnel in hard rock of 700 to 2,800 kg/cm>, The last eight months of production averaged 30 metres per day."

Jan 1, 1980

By Laxman M. Amaratunga

"The fine fraction of tailings is usually excluded from the backfill process and generally disposed on the surface in tailings ponds. Because of the ever increasing demand for fine grinding in the extraction process for improved liberation of valuables in the ore, increased production of fines in the tailings has resulted. The negative impact of surface impoundment of tailings on the environment is well known. Amaratunga and Annor (1989) first presented a concept of fine tailings agglomeration for backfill.This paper presents the development of the novel cold-bond tailings agglomeration (CBTA) technology for non-reactive fine gold mill tailings. For cold-bond agglomeration regular portland cement was used as a binder. In order to cut costs, alternate binders such as lime, and blast furnace slag cement were examined. The effects of binder dosage, curing time, and agglomerate size on the fracture load strength of the agglomerate were studied.Fine tailings agglomerates are mixed with coarse tailings or non-agglomerated fine tailings to make high density fill compositions to be used as underground backfill. Alternatively, the agglomerates can be used as aggregates for surface applications. The paper discusses the environmental and operational benefits of this total tailings utilization and disposal technology.IntroductionIn Canadian mines, mill tailings are probably the most commonly used material for backfill. However, only the coarse fraction of tailings is desirable for use as hydraulic fill or high density fills. The fines are disposed on the surface in tailings ponds since they are unsuitable for underground backfilling owing to the following factors:i) slow settling ratesii) longer percolation timeiii) excessive consumption of binder due to drainageiv) poor compactionFor improved value liberation more milling operations have resolved to fine grinding in their comminution circuits. Gold milling operations driven by down stream leaching processes grind their ores excessively fine. In general, gold mines use hardly any tailings for backfilling. This can result in periodic short falls and the need to search for alternate backfill material from other sources. Apparently, the gold producers are the most identifiable sector which suffer from a shortage of backfill materials ( Amaratunga and Annor, 1989)."

Jan 1, 1994

By A. Nielsen

"The Polaris mining operation (Fig. I) is located on Little Cornwallis Island in the Canadian Arctic, an area with a very limited shipping season and a hostile climate. The operation consists of the underground works, barge-mounted concentrator facilities, concentrate storage and shipping facilities, a personnel accommodation complex and an airstrip.The electrical energy for the operation is provided by four 2.3 MW diesel-generators installed on the barge. The power is distributed through a 4.16 KV system connected directly to the generators and to the large motors. All other loads are supplied at 600 V or 120 V.The power plant has been designed as a Total Energy System with maximum energy efficiency and thus designed for maximum heat recovery. The waste heat recovered on a year 'round basis is expected to be sufficient for the process heating and drying as well as the space heating for offices and accommodations. The engine cooling system consists of two glycol-water loops, a "" cold"" loop for charge air and lube oil cooling and a ""warm"" loop for jacket water cooling. Each loop is complete with heat exchangers f or heat recovery and out-door forced air radiators to dispose of excess heat. A water loop in the concentrator area transfers the heat from the jacket water loop and the exhaust gas boiler to the space heating equipment, the process and the heating system for the accommodations. Part of the exhaust gas is diverted to the concentrate dryers.IntroductionThe Polaris Mine is a highly mechanized, trackless mine with access through a decline. Drill equipment is electrically powered to minimize fumes and heat. Dry development and production drilling eliminates the need for an underground water supply system.The concentrator, power generators, work shops and auxiliary facilities are located on a barge which is beached in a prepared berth on the shore of Crozier Strait. Oil storage tanks, concentrate storage and accommodation buildings are built on shore. The concentrate storage building ha s a 12-month capacity and is fed from the concentrator by a conveyor system.The accommodation building houses up to 240 people, includes administration office, dining and recreation facilities. The complex is housed in on e interconnected building with public use areas alternating with accommodation units.At full production annual concentrate shipments will amount to about 220,000 wet tonnes. This will be accommodated by about 8 shipments during the regular shipping period-August 15 to October 15. Re-supply of general cargo and fuel will be carried out annually in about 3 shipments within the same period."

Jan 1, 1984

By Ngonidzashe Chimwani, David Glasser, Diane Hildebrandt

"Milling circuits play an important role in achieving the desired product; tonnage, size distribution and properties, since they are directly linked to the comminution energy consumption for the whole process. A grinding model with a mathematical simulation was developed for an open milling circuit with a scalped feed from a hydro-cyclone. The circuit was optimized by determining the optimal conditions in terms of energy consumption which would produce a floatable product having a desired size range (i.e. -75 +9 µm) for a platinum-bearing ore. The simulation results showed an operating region comprising the best combination of the ball filling and mill speed to achieve the highest production of a floatable size class, with the minimum energy consumption.INTRODUCTIONBall milling is usually conducted in closed circuits in order to achieve a desired product size without material overgrinding. The technological importance of optimising milling circuits cannot be overemphasised (Toneva and Peukert, 2007). Previous work has brought significant insights into the benefits of modelling milling circuits (Austin et al., 1984; King, 2012; Toneva and Peukert, 2007). However, the ability to model these circuits is dependent on the performance of the classifiers and the fidelity of the grinding mill model used compared to the actual circuit (Toneva and Peukert, 2007). Austin et al. (1984) identified five types of milling circuits as shown in Figure 1 namely; the grinding circuits employing mill in an open circuit, normal closed circuit, reverse closed circuit, open circuit with scalped feed, and, the combined closed circuit. Depending on the presence of the first classifier, the second, or both as well as their corresponding loops (streams), the combined circuit (e) can reduce to; an open (d), a normal closed (b), an open circuit with scalped feed (c) or a reverse closed circuit (a). In the case of an open circuit, the mill is operated without being directly influenced by a classifier. The normal closed circuit does not have the first classifier and therefore the make-up feed is the actual feed stream. Conversely, the reverse closed circuit keeps only the first classifier. The open circuited with scalped feed has only the first classifier but does not have the second one. The overflow stream from the classifier is mixed with the mill product to form the final product for the downstream process."

Jan 1, 2016

By L. Beiza, L. Velásquez-Yévenes

Chalcopyrite is by far the most abundant copper sulphide on Earth, but it has proved to be problematic in its processing by leaching, mainly related to its refractory nature, slow kinetic rates of dissolution and the formation of an inhibiting product layer on its surface during leaching. This study is focussed on the electrochemistry of chalcopyrite dissolution in acidic chloride system, which has recently gained attention as a feasible alternative to conventional sulphate media for chalcopyrite leaching. It is well understood that chalcopyrite dissolution from mineral grains depends strongly on the solution redox potential at the mineral surface as well as the composition of the product layer that may form on its surface. The chloride ions in solution support the Cu(II)/Cu(I) couple as the primary oxidant, as they can form complexes with the cuprous ions formed, increasing the dissolution rate. Additionally, these ions increase the potential range where chalcopyrite can be oxidized. The chemical reaction that occurs on the surface of chalcopyrite in chloride solutions in the presence of cupric ions will be characterized and the coulometry of the reaction will be reported to understand what is happening on the surface of chalcopyrite. Electrochemical experiments have been carried out to investigate the behaviour of mixed potential by open circuit potential (OCP) measurement for 30 minutes using sulphuric acid at pH 1 under nitrogen at 25⁰ C. The effect of chloride was analysed at 0, 20, 50 and 100 g/L and cupric ions from 0 to 50 g/L. The potentials obtained were then used to carry out chronoamperometry tests for a period of 30 minutes in absence of cupric ions. The mixed potential did not show a significant difference with an increase in cupric concentrations in the absence of chloride. Increased mixed potentials and current densities were observed at higher chloride and cupric concentrations. Chronoamperometry tests at those potentials were used to construct a voltammogram trace which has the same trend as a normal cyclic voltammogram but with much lower current densities. INTRODUCTION As a result of the depletion of copper oxides, the leaching of copper sulphide minerals has become the focus of many studies. Chalcopyrite, (CuFeS2) which is the most abundant of copper minerals, has proved to be challenging to process via hydrometallurgy, mainly related to its refractory nature, slow kinetic rates of dissolution and the formation of an inhibiting product layer on its surface, during leaching.

Jan 1, 2019

By A. Abadías Llamas, N. Bartie, M. Stelter

Evaluating the economic viability as well as the sustainability of the Circular Economy (CE) system requires a deep understanding of the distribution of all elements, compounds, alloys, materials etc. in flows. In this paper, the circularity of the copper value chain, including primary and secondary processing, is rigorously evaluated. The studied system comprises the metallic copper production from primary sources (from mineral to metal), copper-containing commodity production (copper is mixed with other metals) and copper recycling through secondary smelting to close the loop. This is linked to photovoltaic (PV) panels and battery storage. A simulation model of this system is created using HSC Sim, considering more than 30 elements (and its various compounds), 180 unit operations and 800 flows. From the mass and energy balances obtained through the simulation, an exergy analysis is conducted to evaluate the resource consumption from a second law (entropy) perspective. Additionally, these results are complemented through a Life Cycle Assessment (LCA), the recovery of technology elements and by-products is discussed, while quantifying the losses through the value chain. Through the digitalization of the complete system, a better CAPEX and OPEX understanding of the metal recovery and losses can be obtained, as well as the associated resource consumption and environmental impacts. New flowsheets and technologies can be evaluated. Several scenarios show how the resource consumption and the environmental impacts are affected by the recovery of different materials to produce different products. INTRODUCTION A key challenge of the circular economy (CE) is the linkage and optimization of both resource and energy efficiency of the (renewable) energy infrastructure and the resource cycle depicted by Figure 1 simultaneously. In addition, the CE requires the linkage of all stakeholders as shown by Figure 1. This has been discussed in some detail by Reuter and Bartie (2019) and Reuter et al. (2019). Copper and all its associated technology elements is one of the key enabling metals that is an economical copper colored thread through CE society (Reuter & Kojo, 2014). The Metal Wheel (Figure 2) shows the linkage of and recovery of various elements associated with copper and other carrier metals. This reveals that copper acts as an important carrier metal because of its properties i.e. the wealth of technology elements it carries from geological ores, but also its capacity to act as solvent of valuable elements and thence a key closer of various complexly connected element and material cycles when these technology elements are released during copper refining. The same applies to all other carrier metals of modern CE society as shown for example recently in a lead policy brief (Blanpain, Reuter & Malfliet, 2019).

Jan 1, 2019

By Liu Tao, Wang Wei, Liao Wenjiang, Yang Weiyan, Tang Bin

Side-blown Furnace technology is a developing technology and was developed very quickly in China in recent years, dozens of reference smelters can be found for the treatment of copper concentrate, copper slag, Lead slag, zinc leaching slag and different kinds of slurry. This paper introduces the working principle of the technology, displays the innovations and specifications of China Nerin's SBF technology, gives examples of the typical reference smelters, and looks ahead to future applications. INTRODUCTION Side-blowing furnace smelting process The side-blowing furnace smelting process (SBF process) is one of the bath smelting process with independent intellectual property rights of China Nerin. This process was originally applied to treat copper concentrate and high-lead slag. The SBF process includes two categories, i.e. to treat copper concentrate and to treat industrial solid waste. Up to now, the SBF process has been applied in around 5 million tons of copper concentrate processing capacity and 20+ plants with high-lead slag reduction process. In recent years, SBF process has been gradually used in zinc leaching slag, electroplating sludge, industrial solid waste (e.g., various metallurgical slag), and electronic waste disposal. The basic principle of SBF process is that the material to be treated is directly charged to bath through charging hole with crushed coal and flux, then, enriched oxygen air would be sent to slag level (Vanyukov/Jifeng style furnaces)or matte level (Baiyin style furnaces) via primary tuyere. In this case, the material completes matte/slag metallurgical reaction of gas-liquid-solid interaction under vigorous mixing, and separate matte and slag at the bottom of furnace hearth. With the help of secondary tuyere, the low-pressure air or enriched oxygen would be provided for secondary combustion in the freeboard. The core of SBF process is process parameter control and engineering design of SBF. Presented as a fixed rectangular type, the SBF is a metallurgical furnace consisting of copper water jackets and refractory material. The SBF is consisting of hearth, middle of furnace and furnace top with the sequence from bottom to top. The middle of furnace itself can be further divided into bath area and post-combustion area with primary tuyere and secondary tuyere at both sides. The furnace hearth is made of refractory bricks, which are fused and rebonded Mg-Cr brick, Mg-Cr brick and clay brick from inside to outside. The furnace top adopts a steel water jacket with charging hole and uptake shaft. The water-cooled partition wall is placed at uptake shaft and near charging hole.

Jan 1, 2019

By M. R. Swaan, G. J. Stewart

"ABSTRACTThe application of failsafe techniques to ask the ""What if"" questions about any process has been well established in the chemical industry for many years. The unusual number of fatalities that occurred in Ontario during 1986-87, prompted Inco management to look at the Fail Safe techniques as a method of improving safety performance.INTRODUCTIONOne of the major areas at any time for potentially disastrous accidents is the shaft, where all men and materials must travel. In April 1987, Inco's Thompson Mine's first Fail Safe Team was established to look at its shaft operations. The initial group of eight people was made up of hourly equipment operators, mine foremen, engineering, maintenance and the superintendent. This group took the name of Shaft Risk Analysis Team or SRAT.The many tasks or jobs that are performed in a shaft prompted the team to divide the work under six major headings; slinging, moving men, skipping, dewatering, shaft inspections and moving materials. Each of these areas contained many different tasks which are interrelated. The only way to understand this interrelationship was to flow chart the steps that actually occur in getting any job done.The flow charting is done by the foreman and operators using existing standard manuals and their experience of what actually takes place. In many cases, it soon became apparent that they had been ""working around"" many problems created by their standards. Fine tuning of these flow charts takes place at the weekly meetings, where input from all involved leads to a better understanding of the actual job being analyzed and its potential hazards. Two major points need to be emphasized in getting good flow charting at a mine:I. It must never be blame oriented or you will never discover how the jobs are actually done and hence you won't know what to improve.2 . All team members must have equal voting rights in deciding how to put the charts together so that they have ownership of the finished product. It is far too easy to end up with amanagement view of the way it should be. Many different techniques or methods are being employed to improve the tasks and reduce the risks associated with any job once the basic flow chart is established. The ones found to be the most useful are highlighted."

Jan 1, 1989

By Donald Roberts

"Parameters f or design and development of an emergency hoist system (Fig. I) were submitted to Timberland Equipment Limited of Woodstock, Ontario by the Potash Corporation of Saskatchewan Mining Limited.The unit is the first of its kind in Canada. Not only does it comply with the province of Saskatchewan's mining regulations, but has the added advantage of being completely mobile and self-contained.As well as being useful in the mining industry, the mobile hoist unit could be used in other emergency situations such as rescuing people trapped in wells.JustificationRevisions to the Mines Regulations under Section 13 of the Occupational Health and Safety Act, 1977, of the Province of Saskatchewan stipulate as follows:"" Section 410 .04(4) The Mine Employer shall make adequate provision at every mine where any worker is hoisted in a shaft not equipped with a manway to enable an y such worker trapped in a conveyance in the shaft to be rescued . "" It was not possible to equip the shafts at PCS Mining' s Divisions with "" manways"" and therefore an alternative was required.BackgroundThe Potash Corporation of Saskatchewan Mining Limited, a subsidiary of the Potash Corporation of Saskatchewan, operates four potash mines. Of these, three, Cory, Rocanville and Lanigan Divisions are wholly owned while the fourth, Allan Division, is 40% owned by Kidd Creek Mines Ltd., but managed by PCS Mining. The four underground operations all have a service and production shaft.From 1979 to 1981 the PCS Mining Division, along with Central Canada Potash, were developing an idea to use the existing ground-mounted timberland hoist at Lanigan as an emergency hoist. This hoist was used in the past for rope-up work at both Rocanville and Lanigan and other potash mine s. The Department of Labour at this time had concerns about the time it would take to load, t ran sport and re-install the hoist on a foundation when needed for an emergency. Weight and size made this particular unit very cumbersome for transporting.In addition, the speed at which this hoist operates is too slow for rescue work. When the Department of Labour expressed concern that this unit would be unsuitable for these rea sons, a search for a suitable unit was initiated.The project responsibility for procuring an emergency hoist for the Potash Corporation of Saskatchewan Mining Limited was assigned to the author, a shaft/ hoist specialist a t PCS Mining."

Jan 1, 1986

By R. Stauber, B. Friedrich, S. Hilgendorf, G. Homm, C. Gellermann

Due to recent regulations limiting tolerated lead concentrations in copper alloys, new recycling strategies have to be considered. This study investigates the suitability of technical approaches in order to recycle leaded copper alloys exemplified by CuZn39Pb2 and produce an alloy in compliance with future lead regulations in the European Union. Three pyrometallurgical methods are considered to achieve this goal, namely vacuum distillation, intermetallic precipitation and fractional crystallisation. The impact of essential process parameters i.e. pressure, temperature, time, atmosphere and quantity of additives are identified using thermochemical and kinetic modelling. An experimental implementation reveals a general suitability of all three refining processes. However, the individual efficiency of the different methods strongly differs. Thus, a detailed assessment is carried out in order to provide an evaluation on the efficiencies of these methods from a technological, economical and environmental point of view. Taking a possible industrial realisation into account, the authors recommend the approach of vacuum distillation, although the energy demanded is comparatively high due to the enthalpy of evaporation for zinc, or rather, lead. However, required lead concentrations can be reached quickly without introducing an external element to the material. The prediction of selectivity and evaporation speed, which are immanent variables for a successful implementation, can be validated experimentally. INTRODUCTION In 2016 approximately 1 million tonnes of leaded copper alloy semi-finished goods with an average lead content of 2.1 Ma% have been produced in the European Union (EU). The global production in the same period exceeded 5.5 million tonnes, while the lead concentration has stayed constant at around 2.4 Ma% for the last 50 years. Up to 80 % of the current production volume is covered with end-of-life scraps, which are used for direct re-melting, depending on the alloy composition. Lead has been a popular alloying element in copper alloys in general, and in particular in brass due to its positive impact on various mechanical, physical and chemical properties, especially machinability. Health and environmental concerns were drivers for many governments to implement limitations on lead. In order to further utilise the resource potential of leaded copper scrap, it is inevitable to develop methods to efficiently eliminate lead from circulation as far as possible. (Glöser, Soulier, & Tercero Espinoza, 2013; Soulier, Glöser-Chahoud, Goldmann & Tercero Espinoza, 2018)

Jan 1, 2019