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By Y. A. Grinin, M. A. Liamina, V. A. Shumskij, Y. I. Sannikov, M. V. Radashin

"A brief history of the development of the Kivcet flash smelting process is presented. The broad scope of the process with an ability to treat high grade lead sulphide concentrates, as well as oxidized lead-zinc residues, is emphasized. Fine coal in the charge serves two purposes, one as auxiliary fuel and the other as a pre-reductant for ferric iron in residues. Current Kivcet operations incorporate a layer of coke or “coke checker” at the surface of the slag in the reaction shaft that acts as a reducing agent for flame products or slag that flow through it. Details of this invention, are given. IntroductionFor many years, the sinter-blast furnace lead smelting process has been the main internationally accepted method used for the production of primary lead metal. In recent times, this process has been unable to meet legislated environmental restrictions and it became clear in the mid-1960s that a new and cleaner process would have to be developed. The present level of Kivcet performance was developed in three main stages:1. flash smelting natural sulphide lead raw materials with the attached electric furnace for production of lead bullion and crude zinc;2. natural raw material smelting with lead and zinc separation in a “coke filter” or “coke checker” at the surface of the slag in the reaction shaft to produce lead bullion and ZnO-bearing slag that can be fed to conventional slag fuming; and3. processing of oxidized lead-bearing raw materials, such as low grade zinc plant residues, with the production of lead bullion and ZnO-bearing slag periodically directed to slag fuming."

Jan 1, 1998

By R. D. H. Wigmore

"The title of this paper probably sounds quite simple and self-explanatory. The problem, however, was not a simple one, and proved somewhat baffling and expensive over a period of years. After outlining the problem, and the methods which were used to overcome it, we will present our solution, or perhaps it would be more accurate to say the partial solution which we reached.We have in the cyanide circuit at Upper Canada two Dorr Single Tray Thickeners with 30 ft. diameter wood stave tanks 12 ft 4 1/ 2 in. deep. In the years prior to 1948, when the tonnage was 250 to 280 tons per day, two stage thickening was practised. As the tonnage was increased to 450, 500, and finally 560-600 the thickeners were used in parallel.Some settling problems were encountered, and in those days the means to combat these problems were very few. The only remedy was to dump in pounds and pounds of lime. Then along came a few reagents, and the writer remembers one in particular, - 11 Carboxel11 (carboxymethlycellulose). When water or solution was added to this reagent, the result was a gooey mass somewhat resembling tapioca. One drawback was that it stayed in this state, and didn't go anywhere. A method was finally developed of containing the mass (or mess) on a fine screen, and dissolving some of it was a fine spray of solution. The only thing the writer recalls was that it was not a very effective settling agent.Shortly after this time, along came the new flocculents, first a few, and then many, and our settling problems were over.But when the settling problem was solved at Upper Canada, some other problems were creeping up on us."

Jan 1, 1965

By A. J. Rigby

Computer generated, actual-scale furnace models have been imported into a rendering software application to provide for impressively realistic surface features and lighting. When these models are imported into a 3D engine framework, it is possible to create content-rich virtual environments allowing for complete interactivity and could provide for a powerful tool to conceptualize refractory lining designs. This technology has been applied to illustrate the optimized design concept of an anode-refining vessel, which can be expected to be completely relined in six days and to last for 4-5 years with minimal partial relining. This technology could easily be applied to other smelter furnace, converter and refining vessel designs.

Jan 1, 2007

By J. H. McKibbon

This paper outlines the basic steps in the sizing of electrostatic precipitators according to the Deutsch equation. Design parameters that do not appear in the equation are also described.

Jan 1, 1971

By Mark Stephenson, Paul Henharen

"The objective in the mineral processing of primary ores is to liberate and effectively recover each individual economically viable diamond with minimal reduction in its potential value.There are three distinctive phases in diamond processing:Feed PreparationWhere the diamonds are liberated and waste material finer than the economically selected cut-off size is rejected. In this section we will discuss the various techniques utilised worldwide to effect feed preparation.Primary ProcessingWhere the diamonds are recovered into a high grade concentrate suitable for the more expensive and selective final recovery process. In this section we will discuss the development of primary processing from jigs and pans to the more current techniques utilised worldwide.Final RecoveryWhere there is selective rejection of waste particles until only diamonds remain. The diamonds may then be acidised prior to packing and marketing.Within these distinct phases a variety of unit processes have been utilised and this paper will discuss their relative merits, limitations and suitability in respect to the various stages of diamond project development from prospecting through to production.Basic capital and operating costs are compared for the autogenous milling route developed in the CIS and the crushing and scrubbing route utilised in the west."

Jan 1, 1994

By Robert D. Brown

"On April 22, 1980, the Ontario government introduced a totally new incentive program for mineral exploration in the province to be effective September I, 1980. Replacing the former Mineral Exploration Assistance Program (MEAP) and also the assistance formerly provided to the mining sector under the Small Business Development Corporations Act, the new Ontario Mineral Exploration Program (OMEP) promises to offer a valuable benefit to certain eligible applicants.Unfortunately, the program is not without its shortcomings: the eligibility for grants or tax credits under the program is comparatively restrictive, and an October 1980 federal budget provision to amend income tax legislation would blunt some of the originally intended attractiveness of the grants and credits.IntroductionIn his budget of April 22, 1980, Ontario Treasurer Frank Miller introduced a new incentive program to encourage mineral exploration in the province.The Ontario Mineral Exploration Program (OMEP) replaces the Mineral Exploration Assistance Program (MEAP), which will be phased out by March 31, 1981. Exploration programs approved under MEAP will continue to qualify for incentives under that program, but will not qualify for OMEP."

Jan 1, 1981

By K Byrne, W. Hawker

The average grade of copper concentrate is steadily declining, forcing the industry to turn to lower grade and metallurgically challenging orebodies. When faced with a low grade, challenging ore, mining companies are often forced to choose between high recovery at low grade or low recovery at high grade. Hydrometallurgical processing of copper concentrates using metathesis reactions, which replace iron with copper in copper-iron sulphide minerals and increase the copper to sulphur ratio in copper sulphide minerals, provide an opportunity to upgrade copper concentrates and potentially allow high copper recovery for a high grade concentrate. Factors such as concentrate mineralogy, reaction time and temperature, and the concentrations of copper, chloride and acid in solution are reported to affect the reaction extent and mineral phases formed by metathesis. In this paper, a copper concentrate has been studied over a wide range of conditions to investigate the influence of these key parameters on the extent of copper metathesis reactions and product phases formed. Solution and solid phases are assayed with the main mineral phases identified by Quantitative X-Ray Diffraction. INTRODUCTION The minerals present in copper deposits determine the prospective processing options either: pyrometallurgical or hydrometallurgical. For copper oxide ores, copper is typically present as oxides, carbonate, chlorides, silicates and sulphates. These minerals are readily solubilised in acid and are usually treated through hydrometallurgical processes. The typical process route for copper oxides is to leach copper into aqueous solution, purify and upgrade the solution phase through solvent extraction and then recover copper as metal by electrowinning (Davenport, King, Schlesinger, & Biswas, 2002). Copper sulphide deposits may contain a multitude of copper bearing minerals, including chalcopyrite, bornite, enargite, covellite, digenite and chalcocite. The copper ore grade and treatment options are affected by the mineral type and distribution. Sulphide ores are typically treated through comminution and mineral flotation to form concentrates, followed by pyrometallurgical smelting. Hydrometallurgy can also be used to process sulphides depending on the mineralogy and impurity levels (Davenport et al., 2002). The trend of decreasing copper grade with increasing impurity levels are common across the mining industry and pose future challenges to copper processing. This trend has prompted research into hydrometallurgical options for upgrading copper concentrates. One promising hydrometallurgical option is using metathesis reactions. Metathesis reactions, sometimes also called double replacement or displacement reactions, are chemical processes involving the exchange of bonds between two chemical compounds or species. In this case, the metathesis reactions involve copper in solution exchanging with cations in the solid sulphide concentrate. The driving force for copper sulphide metathesis is the relative affinity of the exchanging cation elements for sulphide. Based on sulphide precipitation diagrams, the expected order of affinity for sulphide should be Hg2+ > Ag+ > Cu+ > Cu2+ > Bi3+ > Cd2+ > Pb2+ > Sn2+ > Zn2+ > Co2+ > Ni2+ > Fe2+ > Mn2+ (Monhemius, 1977). Hence, copper metathesis reactions are capable of increasing copper content and removing impurities such as iron, nickel, zinc and lead from copper concentrates.

Jan 1, 2019

By I. Orford, V. Rozon

"The Sä Dena Hes Joint Venture Mill was designed to process 450,000 tonnes per year of ore grading 11.3% lead and 13.2% zinc.Run of mine ore is crushed in a jaw crusher and then ground in a closed circuit SAG mill and ball mill. A standard differential float in conventional flotation cells produces quality lead and zinc concentrates. Dewatering of concentrates is achieved using high rate thickeners, lamella clarifiers and pressure filters. Process water is 100% recycle.Plant design was based on metallurgical test work performed on drill core samples. The plant is operating at better than design specifications.INTRODUCTIONSä Dena Hes Joint Venture was formed in 1989 to develop the Sa Dena Hes lead/zinc sulphide mine, located approximately 45 km north of the Town of Watson Lake, Yukon. The mine is owned by a Joint Venture partnership between the operator, Curragh Inc. (87.5%) and Hillsborough Resources Limited (12.5%).The project was completed in a record ten months, working throughout the severe Yukon winter and through an abnormally wet spring and summer. The first truckloads of concentrate were shipped from the mine on August 1, 1991.The mill currently processes 550,000 tonnes of ore per year grading 6.9% lead and 11.5% zinc. Ore is crushed in a jaw crusher and ground to 60% passing 75 microns in a closed circuit operation of one SAG mill and ball mill. Using conventional flotation cells a 150,000 tonnes of concentrate is produced in a year with lead concentrate grading 78% lead and a zinc concentrate grading 60% zinc at overall recoveries of 89% and 90% respectively.Concentrates dewatering is accomplished using high rate thickeners, clarifiers and pressure filters. The lead concentrate and zinc concentrate contain 5.4% water and 7.8% water respectively. Concentrates are trucked to Skagway, Alaska for subsequent shipping to overseas smelters."

Jan 1, 1993

By Li Xibing

The In order to mine safely with reasonable cost, it is necessary to research the fracture characters of backfill and its matches with rock mass. Four kinds backfill with cement-tailing ratio of 0.250, 0.125, 0.100 and 0.083 respectively were tested, and basing on stress-strain curves tested, the deformation characters and fracture laws as well as the damage constitutive equations investigated with damage mechanics whose calculating results are coincided with experiment results well were proposed. The results of damage analysis show that there have different damage characters in different backfill, e. g. the lower the ratio of cement to tailings the lower the damage values when stresses reach to peak value; the lower the content of cement the more rapid in damage and the more abrupt in breakage process. In addition, using catastrophe theory, an energy model was proposed for backfill and rock mass acted each other in mechanics, and the unstable condition of the system was obtained. Research results show that the higher is the stiffness of rock mass the more stable the system, and the lower is the ratio of cement to tailings, the more unstable the system will be. Further more, according to the catastrophe model proposed, the reasonable matches between backfill and rock mass were researched, and a calculation formula that designs reasonable strength of backfill during mining was put forward. It has been proved by practice that the model and theory researched in this paper can guide filling design in mining. Keywords: Cemented tailings backfill, constitutive equation, catastrophe theory, match of backfill and rock mass

May 1, 2007

By Wataru Inagaki, Kozo Kawano, Tetsuya Ando

Cu–Ni–P alloys are the typical precipitation-hardening material. Thus, Cu–Ni–P alloys is expected to apply alternative materials for heat-exchanger tubes. Therefore, it was investigated that the effect of the heat-treatment simulated brazing on mechanical properties of Cu–Ni–P alloy in this study. The hardness of specimen as heat-treatment simulated brazing after solution heat treated (sht) was shown approximately 105 HV. The peak hardness and tensile strength of the specimen as performed aged at 498 K for 43.2 ks following the heat-treatment simulated brazing achieved at approximately 140 HV and 260 MPa, respectively. From these results, it is expected the application to higher-strength tube for heat-exchanger after brazing in a furnace. INTRODUCTION Copper prices have increased dramatically with high in recent years. At the same time, manufacturing industries are required to reduce the environment load by reducing the wall thickness of the copper tubes used in heat exchangers. High-strength copper alloys are developed for reducing the wall thickness, while achieving high strength, of the copper tubes. Moreover, excellent workability and corrosive durability are also required by these copper tubes. Furthermore, the brazing, which are carried out at about 1050 K during the assembly process of heat exchangers, leads to grain growth and re-solving of additive elements. Nomura et al. achieved a tensile strength and Vickers hardness of approximately 260 MPa and 110 HV, respectively, after aging at 723 K for approximately 36 ks. Tamagawa et al. studied that the strength of Cu–Ni–P alloy rapidly increased at about 125 HV during the cooling process after the heat brazing due to the Ni2P precipitates sized 5–10 nm. Furthermore, Inagaki et al. achieved a Vickers hardness of approximately 130 HV by aging at 498 K for 10.8 ks. In the present study, we focused on the lowest aging temperature of 498 K in the previous study, and investigating the effect of heat-treatment simulated brazing on the mechanical properties of the Cu–0.93 mass%Ni–0.24 mass%P alloy.

Jan 1, 2019

By R. W. Pugh, S. L. Ma

"High carbon rods are typically used as the grinding media in rod mills. Usually an AISI I 090 grade of steel is used which provides the high hardness needed for abrasion resistance and sufficient toughness to prevent rod breakage.Work to improve performance in the past has mainly concentrated on increasing the hardness for improved abrasion resistance. This has been accomplished by adding alloys to the steel which uniformly increase the hardness across the rod section.However uniformly increasing the hardness of the rod produces a more brittle product and at high hardness l1evels can lead to rod breakage problems. Heat treatment processes can be used to selectively increase the surface hardness to high levels while maintaining the core at lower hardness levels to give a product that wears slower without breakage. AltaSteel has developed and mill tested a higher hardness heat treated grinding rodThis paper describes the development of a simple model to predict performance improvement with different hardness profiles and the AltaSteel heat-treated grinding rod process. Mill characteristics of the Endako operation are summarized and predicted performance improvements based on hardmess are compared to the approximate 25% achieved in trials running since July, 1991 at this mill. Early results of other mill trials are also presented.INTRODUCTIONAltaSteel, in Edmonton is a major supplier of grinding media. We have a strong desire to continually increase the quality and performance of our products and thereby improve the efficiency of our customers' processes. In the area of grinding rods, we wanted to produce a product with a significant improvement in wear rate. Second to energy, rod consumption usually represents the next highest cost in processing. This paper presents our development of a heattreated grinding rod that has reduced consumption by about 25% in trials to date."

Jan 1, 1993

By O. Moselhi

This paper describes a method for selection of most suitable project delivery systems for capital projects. It expands upon the method advanced by the Construction Industry Institute (CII) in 2003, and incorporates additional decision criteria and project delivery systems in a multi-tier decision computational platform. The paper integrates the analytical hierarchy process to alleviate the inherent subjectivity associated with the assignments of relative weights to selection criteria used in the CII method. It also expands the range of project delivery options to include Public Private Partnership (PPP) and Integrated Project Delivery (IDP). The range of selection criteria was expanded by 60, beyond the 20 criteria of the CII method. Relative effectiveness values are proposed for the added project delivery systems making use of recent project cases in Canada and the USA. The method was implemented in a spreadsheet application. Multiple scenarios were considered for one of the cases presented in the CII study and a sensitivity analysis performed based on the developments made in this paper. The differences in outputs between the CII method and the proposed method are discussed. This is the first decision framework that incorporates both the presently used PPP and the recently introduced IDP, along with the widely used project delivery systems. The developed method allows users to filter out the factors and alternatives that do not apply to the case at hand, based on key inputs at the upper tier. The method is flexible and can easily be expanded upon and customized by the user.

Aug 1, 2013

By Carlos Nieble

Ferteco Mineracao, belonging to the CVRD group, is one of the several iron ore mines in the south east of Brazil. Some of these mines have pit walls of over 350-m height and operations under difficult geo-technical conditions are steadily increasing. In order to mine the 15 million tonnes of high grade hematitic ore remaining in the Fabrica pit, Ferteco had to deal with the following restrictions: Warrant power to its concentrator and pellet plant, relocate the load-out rail road loop, stabilise a poor phillyte slope, relocate a creek on a bench, lower the water table and mine at maximum angles at a slope with phillyte and loose itabirites directly under the rail road loop. Considering these aspects, 3 stages of operations were designed, considering a safe operation, but with the lowest possible security factors at the last mining stage, before the pit would be filled in again with tailings. Stage one above the rail road loop and below the high tension power mast, where the poor phillyte slope was secured by a flexible retaining wall, using old rails and anchors up to 27 meters (80ft) with a working pressure of up till 30 tonnes and stage two with different slope angles, and deep wells, high wall blasting perpendicular to the highwall with relief holes and rerouting the creek in steel half culvert and stage three, the filling in with tailings and recuperation of the deep well pumps. Following studies and analysis were made: Geotechnical and geological studies of all rock formations including rock strength test, a hydrological study to define the lowering of the ground water level, slope stability analysis for every bench and all possible combination of benches, definition of blasting parameters close to the pit slope, a monitoring plan, including topographical movements, vibrations due to blasting, piezometers and pressure cells as well as a plan how to fill up the pit again with tailings in combination with the rising of the ground water level. A close co-operation between the geo-technical consultant, pit operations, the blasting contractor and the geological department resulted in an operation without any incidents, lowering the pit with an additional 3 benches and mining an additional 5 million tonnes of high grade ore.

May 1, 2003

By C. D. Kaeding

IN the summer of 1937, the Algoma Steel Corporation, Limited, decided to embark on the full-scale development o their Helen mine and the erection of plants to produce commercial iron sinter from its ores at a rate ? of fifteen hundred tons per day. The Helen deposit has long been known as one of the largest bodies of non-commercial iron ore in Canada, consisting almost entirely of siderite with a band of pyrite along the north wall and some lenticular inclusions of pyrite throughout the mass. It originally contained a nominal amount of hematite (2,800,000 tons), localized at the western end of the ore-body, but this was mined out during the years 1900 to 1918. For details of the:-geology of the Helen range and of the surrounding district, the reader is referred to Memoir 147, Michipicoten Iron Ranges, Geological Survey of Canada, 1926 (1). The siderite mass was explored by diamond drilling in 1917 and the portion so drilled was estimated to contain 67 million tons of siderite averaging 36.0 per cent Fe, 3.2 per cent S, and 8.2 per cent SiO,, without establishing its ultimate depth or extent. The variation in the sulphur content shown by drilling was from 1.5 per cent in 72 per cent of the tonnage, to 7.2 percent in the remaining 28 percent. The pyritic ore aver-aged about 38 percent iron, but otherwise its composition was substantially the same as that of the siderite mass.

Jan 1, 1940

By H. Hejny

Given the current status of political discussion in Europe about raw materials issues, the sustainable development of the extractive industry and the reduction of its environmental and societal footprint are among the key topics in this discussion. In this context, the European Commission approved the EO-MINERS project (Earth Observation for Monitoring and Observing Environmental and Societal Impacts of Mineral Resources Exploration and Exploitation). The overall aim of EO-MINERS is to bring into play EO-based methods and tools to facilitate and improve interaction between the mineral extractive industry, authorities, and the society in view of its sustainable development while improving its societal acceptability. Developing generic, standardised and reliable EO-based products to monitor the environmental and societal footprint of mining projects is the way chosen by EO-MINERS to improve this interaction, along with the organisation of tripartite meetings gathering all relevant stakeholders. The project uses existing EO knowledge and carries out new developments on three demonstration sites (Sokolov lignite mining area, Czech Republic, Witbank coal field, South Africa, Makmal gold mine, Kyrgyzstan) to further evaluate and demonstrate the capabilities of integrated EObased methods and tools in monitoring the environmental and societal footprints of the extractive industry during all phases of a mining project, from the exploration to the exploitation and closure stages. It contributes providing reliable and objective information about affected ecosystems, populations and societies, to serve as a basis for a sound ?trialogue? between corporate stakeholders, governmental organisations and societal stakeholders. The project already developed a set of consolidated indicators based upon previous knowledge about mining-related issues along with interviews of national, regional, and local stakeholders. These indicators reflect known environmental and societal issues at mining sites as well as the interest of various relevant stakeholder groups. The process described in the EO-MINERS project ? policy assessment > indicator identification > choice of suitable EO tools able to serve monitoring task > monitoring by EO tools > making available objective information for fair dialogue among stakeholders ? is generally applicable to all stages of a mining project. The key to success will be the right choice of indicators addressing the problem and subsequently the right choice of suitable EO tools and finally the willingness of all stakeholders to find a common solution. To this end, the EO-MINERS project organises site-specific ?trialogue? workshops at each demonstrations site, which aim to provide relevant information in adequate quality to the stakeholders as common ground for discussions. Local stakeholders, including mining companies, regulatory bodies, and community representatives are invited to participate in these workshops where the relevance of the developed EO-based products to environmental and societal impact assessment will be discussed. The presentation will discuss results obtained during these trialogue workshops, as well as the success and difficulties and reluctances encountered. The project will further indicate ways and means how the EO products and other project results may be used for monitoring purposes in all phases of a mine.

Aug 1, 2013

By A. Ghahreman

Activated carbon is proven to be able to catalyze arsenic oxidation in aqueous solution under acidic conditions. A commercially available granular coconut shell based activated carbon (AC) has been evaluated for catalysis of arsenic oxidation in this study. The rapid small-scale column test (RSSCT) has been developed to simulate and predict the performance of a carbon-based arsenic treatment system in a pilot scale reactor. Effects of residence time, aeration rate, and initial arsenic concentration in the influent solution on the AC catalysis of arsenic oxidation under dynamic and continuous conditions were studied by manipulating these parameters during RSSCTs. The effectiveness of the AC on arsenic oxidation was examined over the course of constant diffusivity based RSSCT operation. Residence time and initial arsenite concentration are proven to be the most effective impacting parameters on the arsenic oxidation process. For the 218 min residence time, the highest steady-state arsenic oxidation efficiency achieved was 98% with 1 g/L As(III) being treated at pH=1 under 0.05 L/min oxygen aeration. In a long-running RSSCT, the oxidation efficiency did not seem to deteriorate over a course of 666 hours operation, corresponding to 13986 mL total influent throughput volume. INTRODUCTION Arsenic is toxic to both animals and plants and inorganic arsenicals are proven carcinogens in humans. The increased extraction of low grade and complex mineral resources has introduced high ratios of arsenic into metallurgical processes. Discharge of arsenic by acid mine drainage resulting from mining activities has added intensive pressure on the need for proper treatment (Lawrence & Higgs, 1999). Since trivalent arsenite is more toxic and has lower affinity to solid surfaces for adsorption, many treatment systems initiate with an oxidation step to convert arsenite to arsenate and reach the ultimate immobilization of arsenic by the following precipitation of arsenate (Nazari, Radzinski, & Ghahreman, 2016). Scorodite precipitation has been studied extensively as a means of arsenic immobilization due to its high stability and negligible potential of releasing arsenic (Filippou & Demopoulos, 1997). As a stable ferric arsenate mineral, crystalline scorodite can contain up to 30 wt% arsenic (Paktunc, Dutrizac, & Gertsman, 2008). Arsenic must be in its pentavalent oxidation state in order for scorodite to be precipitated via oxidizing ferrous ions (Clancy, Hayes, & Raskin, 2013).

Jan 1, 2019

By Ahmad Hemami

A working code is developed to simulate the forces that are involved in the process of penetration of a bucket into bulk media during loading and excavation. The result of this work can be integrated with the simulation software that many manufacturers of heavy earthmoving machines introduce for training purposes. The simulator programs for machines such as backhoes, loaders, LHDs and the like, used for operator training, are usually only based on the kinematics of the machine. As a result, there is no indication of the effect of the medium when the cutting tool of such a machine touches the excavation material. This is like moving the machine in the air. Feeling the resistance force of medium to cutting by a bucket or blade during loading and excavation is more realistic, both mentally (by vision) and physically (feeling resistance in pedals and levers). The results from force simulation can be used by feedback in creating simulated resistance in these training gears, similar to what is done in flight simulators for aircrafts. In addition to serving for the above purpose, the results of this work can be used as a tool for examination and verification of the effect of a control policy for automation of loading and excavation of bulk media. So far no commercial product has been introduced that has such a capability, despite some extensive research on the subject.

Aug 1, 2013

By Anthony McDonald, Dominic Verhelst

"Hudson Bay Mining and Smelting Co. Ltd. (HBM&S) started up a zinc plant and smelter in late 1930, employing technology of that period. The zinc plant has since converted from a roast/leach operation to a two stage pressure leach operation to eliminate sulphur dioxide emissioms. The smelter, however, is still continuing with much of the original roaster/reverberatory furnace technology. This paper reviews the process flow of the copper smelter, with some comments on past and recent changes.Smelter FlowsheetPrior to July 1993, the HBM&S smelter operation consisted of seven areas, namely feed preparation, roasters, reverberatory furnace, converters, anode casting, fuming plant, and dust collection. A simplified smelter flowsheet is shown in Figure 1. The data for tonnage of materials moving through the circuit are presented in Tables I to 4, and are representative of the time period from July 1992 to June 1993. A change in the zinc plant technology in 1993 resulted in the elimination of a zinc-bearing sulphide leach residue from the smelter feed blend, thereby greatly reducing the zinc content of slag produced in the smelter. Consequently the fuming plant and ancillary equipment, such as the drying plant, were shut down. These areas will, however, be described. With the closure of the zinc fuming furnaces all the slag from the reverberatory furnace is now discarded to stockpile.The over-all smelter metal recoveries are 96% for Cu and rn.5010 for Au and Ag. Figures 2 and 4 indicate historical production numbers of copper, silver and gold, respectively. A review of the minor element distribution in the smelter can be found elsewhere."

Jan 1, 1994

By W. S. Urbanski, A. L. Devernoe, R. W. Stuart, M. S. Walker

"Although many investigators have successfully achieved separations using variations of a concept wherein particles are buoyed upwards against gravity in a magnetically attracted fluid, none have succeeded in separating fine particles in a manner that is practical tor mineral beneficiation. This paper describes laboratory tests and preliminary economic analyses for a new mineral separation process which provides enhanced magnetic buoyancy and separation capabilities. The process provides a means of gravity separation in a regime where conventional gravity separation is inadequate; that is for small differences in mineral densities and for very fine particles. Although early interest centers on gravity separation, the process has also been successfully used for precise magnetic separation.Tests were conducted on a variety of mineral mixtures, ore samples and concentrates including; a sphalerite ore, pyrochlore ores and concentrates, cassiterite pre-concentrates, a chromite pre-concentrate, and heavy mineral placer samples. Separations requiring high discrimination between fine particles narrowly different in density or weak magnetic properties were obtained. Of particular interest is the ability of the process to concentrate or remove oxide minerals or gangues, especially silicates. Potential applications are discussed."

Jan 1, 1985

By Ted Nuorivaara

The use of polymer-surfactant mixtures as foam stabilization agents is a topic that has drawn the attention of several fields but has so far not be thoroughly studied in the context of mineral froth flotation. The present work showcases a study on the impact on flotation performance of cellulose surfactant derivatives and their mixtures with commercial frothers. Bubble size and foam property measurements were run in a two-phase system along with laboratory-scale flotation experiments conducted using a porphyry copper ore from a mine in Finland. A comparative analysis on recovery, grade and flotation kinetics is presented on the products obtained using: i) a partially hydroxylated cellulose derivative (i.e. hydroxypropyl methyl cellulose; HPMC), ii) a commercial polyglycol-ester frother (Nasfroth240; NF240), and iii) mixtures of these two. The results show that using frothers with a total concentration of 30 ppm, the HPMC-NF240 mixtures provide advantages in terms of Cu recovery, (i.e., up to 95%), compared to the use of either NF240 (ca. 76%) or HPMC alone (ca. 1%). Interestingly, only a small proportion of HPMC mixed with NF240 was sufficient to produce a measurable increase in recovery. This improved performance was attributed to a synergistic effect of the polymer-surfactant mixture which further reduced the bubble sizes in a two-phase system compared to that of NF240. The present work offers a further step into understanding the phenomena taking place when a short chained surfactant is either substituted or mixed with a cellulose derivative, offering potential ways of exploiting this behavior for the enrichment of minerals. INTRODUCTION Froth flotation has been the predominant technology used to beneficiate sulfidic copper minerals such as chalcopyrite ever since it was invented in the late 19th century. The reason for the domination of flotation technology lies behind its versatility in selective separation as well as relatively cheap operational costs (Wills & Finch 2015). Even though flotation of sulfidic copper minerals is moderately well understood based on the broad research work done around the subject (e.g., Albijanic et al., 2015, Bournival et al., 2017, Wang et al., 2015, Mikhlin et al., 2015, to cite only a few), the modern era presents new challenges to the primary raw material industry, including questions regarding sustainability of mining operations accompanied by a need to improve the process efficiencies due to declining ore grades (Edraki et al., 2014).

Jan 1, 2019