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By R. Bridge

Mining companies are progressively finding themselves having to develop operations for more complex, deeper ore bodies, which require the use of more complicated mining and mineral processing methods. This coupled with the generally increasing trend of higher capital and operational costs have led to a need in the industry for improving ore body knowledge, mining and mineral processing. Geometallurgy is a relatively new concept that can link the three aspects, and when used effectively it is a method that can produce the necessary information for designing and building operations for reliable and optimum performance. Geometallurgy not only involves the integration of geological, mining, and metallurgical fields, but it can also be expanded to include environmental and economic information to produce a block model that can be used to obtain the best outcome in terms of plant response and plant sustainability. The geometallurgical program should not end once the block model has been populated. A detailed geometallurgical approach should be applied to all stages of the operation encompassing sample selection, through to equipment sizing and production forecasting. It can also become a useful optimisation tool thereby allowing the operation to be proactive rather than reactive against changes in the ore characteristics. With the development of metallurgical performance predictors it is possible to quantify how the ore will behave in the process plant in terms of throughput, recoveries, reagent consumption, etc. Therefore, in terms of responsible mining and sustainable resource development, a geometallurgical approach to operations is the logical step for plant development. In this paper we describe the state of the technology with respect to the application of geometallurgical modeling and its utilisation in the fields of design, forecasting and optimisation of mining operations.

Aug 1, 2013

By R. Saulters, M. Garshasb

The Transportable Moisture Limit (TML) is the copper concentrate moisture content beyond which liquefaction may occur during shipping. This study indicates that TML is a function of mineralogical characteristics of concentrate, which in turn may directly affect moisture retention properties. A comprehensive mineralogical and chemical analysis was conducted to determine the impact of composition and mineral content on concentrate’s TML. The analysis indicates that the high TML value is associated with particular minerals such as gypsum, swelling clay, muscovite and pyrite. Since higher TML values allow higher moisture contents for shipping, this property is desirable for concentrate inventory management and shipping flexibility. Scarcity of minerals that may increase TML contributes to limitations of shippable inventory. INTRODUCTION TML is the maximum moisture content of copper concentrate beyond which liquefaction may occur. Moisture content above the TML is considered a shipping hazard as it may result in loss of lives and cargo. The details of TML testing are given in ISO 12742:2007 (ISO 12742). In principle, TML is 90% of the Flow Moisture Point. The method of testing for Flow Moisture Point and TML is based on procedures that provide pulsations that simulate the cyclic force of ocean waves to determine moisture levels at which the onset of concentrate flow may occur. International Maritime Organization (IMO) provides guidelines for developing and approving TML testing and moisture determination in the context of its Moisture Management Plan (MMP) for shipment of copper concentrate cargoes (IMO 2013). The IMO regulations are specifically provided in MSC.1/Circ. 1454 “Guidelines for developing and approving procedures for sampling, testing and controlling the moisture content for solid bulk cargoes which may liquefy”. Since each port has its own operational characteristics, resource availability, peculiarities of governmental structure, and desired methods of interaction with the authorities, it is necessary to adopt the IMO guidelines to the local requirements. In order to establish a TML and MMP measurement system, it is required to implement and follow proper guidelines for the relevant procedures. For example, sampling procedures are developed in accordance of ISO 12743, “Copper, lead, zinc, and nickel concentrates – Sampling procedures for determination of metal and moisture content”. Moisture content of copper concentrate samples are determined in accordance with ISO 10251, “Copper, lead, zinc, and nickel concentrates – Determination of mass loss of bulk material on drying”. To assure strict adherence to the MMP procedures, the overall system needs to be integrated into the site QA/QC system, which would make them auditable and certified in accordance with standards such as ISO 9001, ISO 17025, and ISO 17020.

Jan 1, 2019

By R. H. Coloquhoun

"INTRODUCTIONDome Mines Limited is located on 3,149 acres in Tisdale, Whitney and Shaw Townships in the City of Timmins in Northern Ontario. The mine and mill facilities are two miles from the town of South Porcupine, now part of greater Timmins.In April of 1909, J.S. Wilson of Massey, Ontario headed a party of prospectors into the Porcupine region. Disembarking from the Temiskaming and Northern Ontario Railway at mileage 222, they established an initial camp on the shore of Frederick House Lake. Following the lake ""break up"" they canoed to their destination, Porcupine Lake, and established a base camp on the west side of the lake by May 23, 1909. At a point approximately two miles from Porcupine Lake the moss on an outcrop was rolled back and gold was discovered.In the spring of 1911, the first mill structure was in the process of being built and consisted of 40 stamp mills, five S'x22' pebble mills and the ancillary equipment for a 300-400 T.P.O. operation. This structure was destroyed along with much of the Porcupine Camp by the famous Porcupine Fire of July 1911. By mid-1912, a steel ""fire proof"" structure was built and utilizing the salvaged mills from the fire, supplemented by the installation of three, 8'x30"" cylindro-conical Hardinge ball mills purchased third hand from Vipond Mines, processed 75,088 tons by year's end.The mill was again levelled by fire in October of 1929, this time caused by an overheated boiler stack. This 11 fire proof"" structure was immediately replaced by a wood structure and again the three Hardinge conical mills and the five 5'x22' pebble mills were restored and utilized. The new mill operated close to its daily capacity of 1500 tons during the last two months of 1930. During the rebuilding phase, the crushing plant was remodelled and a new crusher installed. These plants were protected from fire by a 60,000 gallon water tank and an automatic sprinkler system."

Jan 1, 1984

By A. Probst, N. Moores, J. A. Finch

"In 1994 McGill and INCO (Sudbury operation) initiated transfer of technology originally developed to detect level in flotation columns to thickeners. A conductivity probe was installed in a concentrate thickener to monitor the mud-line and solids load. The probe consisted in a vertical tube with 32 equally-spaced electrode rings connected in pairs to give 31 conductivity cells. Conductivity was measured as a function of depth (so-called conductivity profile) from which the mud line and % solids could be inferred. The probe was installed in October 1994 and has been successfully detecting profiles since then. This communication includes a description of incidents and experiences during this technology transfer process including: probe installation and start-up; software modifications to accommodate operating conditions and operators' needs; calibration procedures to adjust to changes in performance as a consequence of long term use; and preliminary observations on the last stage, still in progress, signal interpretation and use to facilitate operation, particularly by providing early reaction to upstream disturbances.IntroductionThickening is one of the basic unit operations in mineral processing and hydrometallurgy used to separate suspended solid particles from a solid-liquid dispersion by gravity settling. As the dispersion enters the thickener, the solids settle and clarified water overflows the top. A thickener, which may be continuous or batch, is basically a large vertical cylindrical tank provided with devices to feed the dispersion and to remove the settled solids. The tank provides residence time for the particles to settle. The dispersion enters the thickener through the feedwell whose function is to dissipate the momentum of the feed stream and to create relatively quiescent conditions around the entering point. Rotating rake arms move the settled solids towards the discharge point located at the bottom centre of the tank.There are normally three distinct zones in a thickener: a top layer of clear water, a middle layer called a sedimenting zone in which particles are settling, and a bottom layer called a compression zone in which particles are pressed together as a consequence of the weight of accumulated particles. Proper operation of a thickener requires two simultaneous conditions: no solids in the water overflowing the top and the target solids content in the underflow pulp. To achieve these conditions is difficult because sensors for detecting the sediment line (interface between clear water and sedimenting zone) and the solids content in the compression zone are not commercially available."

Jan 1, 1998

By C. R. Edwards

"Equity Silver Mines Limited owns and operates an open pit silver 111ine located 37 km southeast of Houston in the Omineca District of British Columbia, Figure l. The property is approximately 300 km west of Prince George and 417 km east of the Port of Prince Rupert. This is the largest silver mine in Canada and one of the five largest silver producers in the western world.Equity is a subsidiary of Placer Development Limited, which owns 70 of Equity's issued and outstanding Common Shares. Placer was responsible for financing and developing the property and now provides general management and administrative services.All mining is by open pit methods, using three 1600 P&H shovels, eight 45 t Caterpillar and three 77 t Wabco trucks, three 40R rotary drills, two D8 dozers, one 9888 front-end loader and other ancillary equipment. We are now phasing out the Southern Tail pit and bringing the Main Zone into productior; see Figure 2.Comminution in the concentrator is achieved via primary, secondary, and tertiary crushers followed by rod and ball mills. The flotation circuit, with Southern Tail ore, produces a concentrate containing between 3 000 and 6 500 g of silver per tonne, 201 copper, and 25 g of gold per tonne. Running Main Zone ore, concentrate grades will decrease to about 4 500 g of silver per tonne, 17% copper, and 18 g of gold per tonne. Design milling capacity is 5 300 tpd, which is often exceeded.The flotation concentrate contains relatively high levels of antimony and arsenic. Because of these impurities Equity was unable to obtain economically acceptable long-term sales contracts for flotation concentrate. For this reason a leaching process was developed, and a plant designed, built and placed in operation. Due to the complex and unique technology involved, many start-up problems were encountered but by March 1983 the leach plant was considered ""fully operational."" This paper describes some of our progress and problems in bringing the Equity leach plant through its start-up period."

Jan 1, 1984

By R. E. Whitehead

"The earliest gold-mining operations in Ontario included the Cordova, Deloro, and Sophia mines north of Lake Ontario (see for example, Miller and Knight, 1914). The Cordova deposit, worked intermittently between 1891 and 1940, is reported to have yielded 22 774 oz of gold from 127 670 tons of ore averaging 0.19 oz/ton Au (Gordon et al., 1979). The property remained inactive for many years. In 1988, Gunnar Gold Inc. and Mill City Gold Inc. conducted a drilling program along the main Cordova vein. The present study is based on core from three drill holes through the Cordova gabbro (Fig. 1) and the carbonate alteration zones enveloping the quartz-carbonate veins. The relatively narrow width of the alteration zone and the availability of continuous sections of core provided an opportunity to conduct a compact mineralogical-chemical study of carbonate alteration and its effect on the redistribution of rare earth elements (REEs). Although the Cordova deposit is not of great economic importance, this study has revealed a number of interesting features which are interpreted as resulting from processes that may be applicable to gold-bearing quartz vein deposits elsewhere.The Cordova gabbro is layered in places, varying from fine-to coarse-grained and in color from light gray (feldspathic) to dark gray-green (amphibolitic). However, layering within the section of the gabbro considered here and which hosts the Cordova gold deposit is mainly textural rather than compositional. Samples of unaltered gabbro selected for analyses were medium- to coarse-grained; amphibolitic and anorthositic phases were avoided as being less representative of the general chemical composition of the gabbro.The Cordova intrusion has undergone regional metamorphism and consists dominantly of actinolite and saussuritized plagioclase with smaller amounts of granular clinozoisite and flaky chlorite; minor quartz, ilmenite, titanite, apatite, and calcite are typically present. Some actinolite grains have irregular rims of pleochroic green hornblende that is optically continuous with the actinolite. Both chlorite and clinozoisite occur in two modes. Chlorite occurs as an alteration product of actinolite and as separate grains in the groundmass. Clinozoisite and albite occur together as regional metamorphic alteration products of original plagioclase crystals and as discrete grains in the matrix.The main gold-bearing quartz-carbonate occurrence occupies an easterly trending, steeply southward-dipping carbonatized shear zone up to 15 m wide and 700 m long. The gold occurs principally in irregular lenses and veins of quartz-carbonate up to a foot thick occupying the lower part of the schistose shear zone within the gabbro. This zone consists of chlorite, biotite, and carbonate and is flanked on both sides by chlorite-carbonate schist lacking biotite. The chlorite-carbonate schist in turn grades into regionally metamorphosed actinolite-bearing undeformed gabbro.The dominant sulfide within the quartz-carbonate lenses and in the surrounding schist is pyrite which occurs mainly as small cubes (1 mm to 2 mm) generally making up less than 10% of the rock. Rare blebs of chalcopyrite are present. Magnetite is common throughout the schist."

Jan 1, 2001

By S. Eckert, M. Reuter, O. Keplinger, D. Obiso, N. Shevchenko, M. Akashi

We report on laboratory experiments focusing on bubbling phenomena arising from gas injection through a top submerged lance (TSL) in a liquid metal layer. The objective of this study is to reveal the characteristics of the formation and behaviour of the particular bubbles at different gas flow rates and submergence depths of lance. Visualization was performed in the eutectic alloy GaInSn using X-ray radiography. Argon bubbles were injected through the injection positions at three different submergence depths. Essential parameters such as the bubble size, bubble shape, detachment frequency and the two- dimensional gas distribution in the flat vessel were obtained by image processing. The results show that the deep position of the submerged lance causes an asymmetric large-scale circulation inside the fluid vessel. Bubble detachment frequencies were calculated by Fast Fourier Transformation from fluctuations of the image intensity in the vicinity of the nozzle injection point. This frequency does not show strong variations with respect to changes of the gas flow rate and the submergence depth of the nozzle. An increasing gas flow rate results in an increasing two-dimensional projected bubble area and the occurrence of a significant number of small bubbles being trapped by the strong fluid flow in the liquid metal layer. INTRODUCTION The Top Submerged Lance (TSL) process has been an essential component of the metallurgical industry since the emergence of first concepts about 50 years ago (Floyd, 2005). The process, which is characterized by blowing reactive gas through the submerged lance into a liquid bath, was developed mainly for improving the tin smelting process and has been applied for waste treatment or the production of several metals such as copper, lead, nickel and zinc. The productivity of the reactor is determined mainly by the efficiency of the chemical reaction between the gas or solid particles and the liquid. To achieve a high efficiency requires good exchange between the gas-liquid phase and therefore a large interface area. The efficiency changes greatly with the number and size of bubbles injected from the nozzle and the extent of splashing on the free surface. The degree of turbulence in the bath plays a decisive role here. For the reason, the design of an efficient TSL process requires a profound knowledge of the fluid dynamics especially for complex multiphase flows.

Jan 1, 2019

By J. Keough

During industrial electrowinning of copper, uneven growth can impact current efficiency, cause short-circuits in the cell and negatively impact the final quality of the metal product. Copper electrodeposits can be made smoother by incorporating a host of additives in the electrolyte but uneven copper growth may still occur due to variations in electrolyte composition throughout the cell-house. To quantify roughness, 3D images of the surface may be combined with scaling analysis to gain further insight into the deposit structure. Moreover, scaling analysis may offer a method to predict long-term deposit structure using short-term copper deposits. In this paper, the application of Atomic Force Microscopy (AFM) and scaling analysis are used to monitor copper growth and extract surface roughness characteristics. The cathode was a 304 stainless steel disc that was polished to a mirror finish. Copper was deposited onto this surface over a range of 10, 20 and 30 minutes from an acidic sulfate solution (40 g/L Cu) at 40oC from an electrolyte that contained or was free of guar. AFM images were acquired to obtain 3D topographical information on the samples and scaling analysis was performed to extract the root-mean-squared (rms) roughness, periodicity (grain size), aspect ratio of surface features and growth mechanism. INTRODUCTION During the process of copper electrowinning, the dissolved metal ions are reduced and thereby deposited onto a cathode substrate which is later recovered by stripping the metal from the cathode surface. Although this process is widely used in the industry, the conditions of the deposition procedure may cause the growth of the metal product to be uneven which can not only impact the current efficiency and final copper product quality but may also lead to plant interruptions through short-circuits if the anode and cathode make contact. To create more uniform copper deposits, a host of additives may be included in the electrolyte which may act as brighteners, leveling agents and inhibitors which may include guar gum, glue, thiourea or polysaccharides or polyacrylamides and their industrial use has changed over time (Moats, Luyima, & Cui, 2015). As such, the electrolyte composition is often complex and subtle variations in the solution matrix throughout the cell-house may impact the surface morphology of the final copper product. In an effort to quantify the roughness of copper samples, image analysis is often applied and may include Scanning Electron Microscopy (SEM) or Atomic Force Microscopy characterization. While both are appropriate, AFM offers 3-dimensional information which may be used to both qualitatively and quantitatively characterize the surface characteristics.

Jan 1, 2019

By Ron Bradburn, Derek Perkins

"The Brenda mine is located in south central British Columbia, 22 km north west of Peachland, in the heart of the Okanagan Valley. Situated at an elevation of 1500 metres, the mine area straddles the divide between the headwaters of Trepanier Creek and Peachland Creek. These flow into Okanagan Lake, the chief valley water resource located 1100 metres below. The creeks are of great significance to the water management of the area as they constitute the water supply for the Municipality of Peachland, a thriving retirement community with a population of 5000. A location map is shown in Figure 1.The history of the Brenda Mine operation has been well documented previously. After twenty years of operation, the open pit mine shut down in June 1990 due to the depletion of its ore reserves. During the course of its life, it produced 109 million tonnes of waste rock which is thought to contain approximately 0.10% copper as chalcopyrite and 0.015% molybdenum as molybdenite stored in four rock piles around the pit. Similarly, approximately 200 million tonnes of tailings were produced grading 0.015% copper and 0.005% molybdenum. During operation, tailings were hydrocycloned to produce a coarse sand for dam building with the fine slime stored behind the dam. It is estimated that 40% of the tailings were used for the dam construction. The layout of the mine site is shown in Figure 2.The mine was operated under a zero discharge permit. This was achieved by diverting practically all surface water flowing towards the mine area around the site. Water affected by mine operations, was discharged to the tailings impoundment, and recycled to the mill for reuse. The recycle rate was 90% and seepage from the tailings embankment and saddle dam was pumped back to the tailings pond. Because acid drainage was not a problem with the low grade Brenda rock, runoff from the rockpiles was initially released to the local environment. However, due to the alkaline nature of the waste rock, dissolution of low concentrations of molybdenum in the site runoff waters became evident by the late 1970's. The molybdenum content of Peachland Lake, the water reservoir for the municipality of Peachland, approached the Canadian Drinking Water Standard of 0.25 ppm. By 1979, a series of diversion and collection ditches was completed to direct all runoff water from the rockpiles to the tailings impoundment area. This had a significant detrimental effect on the site water balance causing it to go into a considerable surplus condition. During normal operation, this oversupply was contained by the rate of tailings dam building. An economic shutdown of the mine during 1984 forced a limited discharge from the site to protect the integrity of the tailings dam."

Jan 1, 1995

By Wayne F. Luinstra, Frédérique Vincent, André R. Laplante

"A test to determine the amount of gravity recoverable gold (GRG) in an ore is described. Typical GRG results are given; possible uses are discussed. A case study of how the technology was used at Hemlo Gold Mines is presented.IntroductionUsing gravity to supplement either flotation or cyanidation is a well established practice in the gold industry. It differs from most recovery circuits in that most of the gold recovered by gravity would be recovered in any case by the circuit downstream, be it flotation or cyanidation. The economic justification of gravity is therefore based on small margins (for example. a net smelter return of gravity gold of 99%, as opposed to 94% by flotation). It was easily demonstrated when either flotation or cyanidation were relatively inefficient processes, and labour costs low (as gravity can be labour intensive). Over the past thirty years, the advent of better flotation machines (flash, column, high capacity), more effective collectors, and better control systems has increased flotation's metallurgical performance, thereby decreasing the incentive for gravity recovery. Cyanidation technology has undergone similar changes, with the advent of activated carbon, oxygen and lead nitrate addition, and improved impeller design.Today, gravity could remain an attractive option only inasmuch as it can be implemented with very low capital and operating costs. This has resulted in a relative shift away from gold gravity recovery (except for alluvial deposits), in the seventies and eighties. For example, as of the early nineties, gold gravity recovery has disappeared from the typical South African flow sheet. The advent of the Knelson Concentrator, at the beginning of the eighties, foreshadowed a resurgence of gravity recovery, as gold's grinding and classification behaviour (Banisi, Laplante and Marois, 1991) makes it possible to achieve adequate gold recoveries with very simple, Knelson based, gravity circuits (Laplante et al., 1994).Consider for example the Hemlo recovery circuit, made of a single 61 cm x 122 cm (2' x 4') screen, an automated 76 cm (30"") Knelson feeding a Gemeni table, to produce a 70-80% Au concentrate which accounts for approximately one fourth of the gold production (about 3 million grams per year). The low capital and operating costs make it possible to justify the gravity circuit on the basis of savings in other operating costs (e.g. fewer carbon stripping and reactivation cycles). Any improvements in recovery, which would at Hemlo be difficult to measure, would be additional benefits. Such a circuit is rapidly becoming a standard in Canadian and Australian operations.The justification for installing a gravity circuit must be first based on an estimate of how much gold it will recover (irrespective of what the potential benefits are). This is a function of the nature of the gravity circuit to be installed; however, Knelson circuits, by their very simplicity and efficiency, limit the options. This leaves the amenability of the ore to gravity recovery as the single most important parameter. Various approaches can be used to characterize this response (Woodcock, 1993). The present work sought a route which would fulfil the following criteria:"

Jan 1, 1996

By O. Benavente, M. C. Hernández

Selenium is obtained mainly from copper anode slimes, which is a by-product of the copper electrorefining process. Selenium is present in copper anode slimes as silver and copper selenides. Traditionally, selenium has been obtained by pyrometallurgical processes; however, hydrometallurgy has now been incorporated either full or partial to the conventional process. Some hydrometallurgical processes are been developed and used at laboratory level an industrial scale. This document describes theoretical background and results of characterization and dissolution of selenium from a decopperized anode slimes in an alkaline oxidizing system. The results have indicated that the oxidizing leaching process reaches a practically complete dissolution of selenium (up to 90 %); the undissolved portion is due to the generation of a AgCl around the particle surfaces that delays the progression of the reaction. The results indicated that 90% of selenium extraction can be obtained by dissolution under moderate conditions of pH, temperature and reagent concentration in oxidizing medium. INTRODUCTION Selenium is distributed in the earth's crust in most rocks and soils. Selenium is essential, as it has wide range of applications in the medical and food industries as well as, used in semiconductor, sensors and photochemical devices. Selenium is rarely found in the native state, 90% of selenium is obtained from copper anode slimes, which is a byproduct of the copper electrorefining process (Lu, Chang, Yang, & Xie, 2015; Xiao et al., 2017; Yang et al., 2018). In anode slimes, selenium is present as an intermetallic compound, as Ag-Cu selenides. Currently, selenium is recovered from the anode slimes via hydrometallurgical and pyrometallurgical processes (Lu et al., 2015). The traditional pyrometallurgical method recovers selenium effectively, but faces difficulty as a result of increased energy costs and strict environmental regulations associated with the process (Li, Yang, Jin, Chen, & Tong, 2017). Therefore, hydrometallurgical processes are proposed in oxidizing medium such as nitric acid, alkaline leaching with sodium nitrate, wet chlorination and pressure leaching, amongst others (Hait, Jana, & Sanyal, 2012; Kim & Wang, 2010; Xiao et al., 2017; Yang et al., 2018).

Jan 1, 2019

By Avi Nanda, Robert Brandt, Alex Piatkiewicz, David M. Jones

In 2010 the United States Environmental Protection Agency (UESPA) revised the National Ambient Air Quality Standards (NAAQS) by lowering the permissible value to a 1-hour average SO2 concentration of 75 ppb. In 2012, the USEPA designated the Miami area as non-attaining for the new SO2 NAAQS. A number of options were studied to assist in bringing the Miami smelter into attainment with the new standard. The chosen solution included: Installation of a large tertiary hooding system over the converter aisle; Hooding of the anode vessels and treatment of the anode gas in a baghouse; Installation of a large 2,325,212 Nm3/hr caustic scrubber to treat the captured converter fugitives and anode off-gases; Installation of a secondary ESP for the tapping fugitive off-gas treatment system; and the continuous scrubbing of all acid plant tail stack gases. The modifications have reduced smelter SO2 emissions by 85%, achieving a world class SO2 stack emissions rate of 0.45 kg of SO2/tonne Cu. INTRODUCTION The Freeport McMoRan Miami smelter has been previously described in literature (Steinborn, Pasca, Sallee, Binegar, Bhappu). For the sake of brevity a brief description is given here. The Miami smelter is located 129 km east of Phoenix, Arizona near the town of Miami, Arizona. The smelter was established in 1915 and has undergone a number of process upgrades and modernizations over the years. The present smelter consists of an Isasmelt furnace, an electric settling furnace, four Hoboken converters, two anode refining vessels & one casting wheel; and one double contact sulfuric acid plant. Concentrates are railed & trucked to Miami from four Freeport McMoRan mines located in Arizona & New Mexico. The wet concentrates, typically containing 10% H2O with a Cu/S ratio of 0.85, are smelted in an Isasmelt furnace to a matte grade of 58%. The entire output from the Isasmelt is laundered into an Elkem electric furnace (originally installed in 1972 as the primary smelting furnace) for separation into matte and slag. Slag containing 0.85% Cu is skimmed from the electric furnace and discarded. Molten matte, containing 60% Cu, is tapped from the electric furnace and processed in one of four Hoboken converters to produce blister copper. Two anode refining vessels further refine the blister copper before the final product is cast on a single, 30 mold, casting wheel. All of the anodes produced at Miami are shipped to the Freeport McMoRan electrolytic refinery in El Paso Texas for refining into cathode copper. Smelter off-gases from the Isasmelt, electric furnace and Hoboken converters are combined and treated in a double contact acid plant to produce sulfuric acid. Acid produced at Miami is internally consumed at Freeport McMoRan mines in the region.

Jan 1, 2019

By M. Elmadagli

Wear mechanisms of two commercial aluminum alloys, namely a spray formed and extruded Al-25%Si alloy, and a sand cast A390 alloy (18.5 %Si) that are known to be successful in cylinder bore technology were studied to have a better understanding of the requirements for wear resistant aluminum alloy design. From durability point of view these alloys are known to be successful for engine bore applications. It is interesting to compare the wear mechanisms in these alloys because they have significantly different silicon content and microstructures, and were produced using different techniques. The wear tests were done using a block-on-ring configuration, in controlled air atmospheres (5 % relative humidity). In spite of the differences in silicon content and microstructure, the wear mechanisms were similar and consisted of formation and destruction of hard tribolayers that are either formed by either mechanical or chemical interactions on contact surfaces. The differences in the wear rates between the two alloys arose from the differences in the hardness of the worn surfaces, and the alloy's ability to sustain the tribolayers on the worn surfaces.

Jan 1, 2004

By S. D. A. Hannot, H. Knapp

"High pressure grinding rolls (HPGR’s) are an energy efficient solution for comminution of industrial minerals and metal ores. The technology is successfully applied in an expanding field of applications throughout the minerals industry. HPGR’s compress and grind the particle bed in the operating gap between two counter-rotating rolls. The equipment provides a high capacity, low energy alternative for applications with conventional tertiary and quaternary crushing and grinding, with size reduction reaching from 100mm down to 25 microns product. In this process, maintaining a controlled and evenly distributed operating pressure between the rolls is very important. Due to feed segregation, the pressure distribution along the roll’s length can become uneven, resulting in a coarser product. This paper will show that the best approach to compensate and adapt for the uneven pressure is to allow the rolls to skew relative to each other. Coarser products are also caused by an inadequate pressure distribution: the much-discussed edge effect. To minimise the consequence of the edge effect, an optimised ratio for the roll length to diameter will be calculated. Furthermore, the design philosophy behind a cheek plate will be detailed. This cheek plate is used to retain the material between the rolls, reducing the edge effect while still allowing roll skew. Together these design considerations help optimise HPGR’s performance.INTRODUCTION Initially developed for the cement industry during the energy crisis in the 1980’s , high pressure grinding has been an accepted technology for many years. It has been successfully implemented throughout the minerals industry, with applications in base and precious metals, as well as industrial minerals processing. The comminution process relies mainly on size reduction through to compression and inter-particle crushing in a particle bed in the operating gap between two counter-rotating rolls. The equipment typically has a high unit capacity and low energy consumption, and is increasingly used as an alternative to conventional tertiary and quaternary crushing and grinding, in particle size ranges reaching from 100mm down to 25 microns. It can be applied to moist ores and dry material, in open circuit arrangement or in a closed circuit with dry or wet classification (screening, air classification)."

Jan 1, 2019

By H. B. Gillis

THE following pages c9ntain a brief description of the processes involved in the manufacture of silica brick at the Sydney plant of the Dominion Steel & Coal Corporation. Silica brick manufacture was commenced at Sydney in 1918, for the purpose of supplying the requirements of the steel plant, where about three million silica bricks were used annually. Previous to that date, all silica refractories were brought to Sydney from American sources and were subject to a freight cost of nine dollars per ton, or about twenty-seven dollars per thousand. As a preliminary step to manufacture, a thorough investigation of all possible sources of silica in the vicinity of Sydney was carried on during 1916 and 1917. ?Eighteen separate deposits were examined. Of this number, only four were found suitable for brick making, and the Leitches Creek silica, being the nearest of these four deposits to Sydney, was selected as the raw material for the proposed plant. CRYSTAL MODIFICATIONS OF SILICA A brief review of the silica minerals may serve to indicate why careful investigation of sources of supply is necessary before actual brick-making is attempted. The crystal modifications of silica, i.e., quartz, tridymite, and cristobalite, all possessing alpha and beta phases, play a most important part in determining the suitability or otherwise of the resulting brick. By heat treatment, any one of these forms of silica may be converted to either of the others. The inversion temperatures (after Fenner) in degrees Centigrade are shown on rage 160.

Jan 1, 1938

By T. Crivits, I. Bellemans, K. Verbeken

In the metallurgical industry, furnace cooling is often employed to enhance furnace lifetime. Hence, the pyrometallurgical phase comes in close contact with a non-air coolant. The proximity of the pyrometallurgical phases to the coolants may lead to dangerous situations. In case of a breach of the cooling circuit, the contact between a hot and a cold liquid may cause a vapour explosion. The outcome of this vapour explosion includes a pressure wave which could result in damage to the furnace, bodily injuries or even casualties. To prevent such dramatic events, the origin of vapour explosions is currently being studied in more detail. For this purpose, OpenFOAM was used to perform Computational Fluid Dynamics (CFD) simulations. In a first step, simulations were run to get a better idea of the influence of the compressibility on the simulations. The obtained results help in getting a better insight on the need of computational power, and possible numerical errors using compressible or incompressible solvers in future simulations. In a second step, systems without vapour explosions were simulated and compared to experimental results with mercury and aluminum. INTRODUCTION Humanity started using metals for tools around the Chalcolitic (~7400 cal BC), when copper was melted in small furnaces (Radivojevic, Rehren, Farid, Pernicka, & Camurcouglu, 2017). Nowadays, metallurgy has grown into a major industry and it has become impossible to imagine society without metals. In the pyrometallurgical industry, coolants (low temperature liquids, for example water) and pyrometallurgical phases (high temperature liquids, such as slags, mattes and alloys) are typically present in close proximity of each other. The interaction between these two phases can have catastrophic consequences (Tveit et al., 2008; Ferguson & Zsamboky, 2017). A vapour explosion may occur as the consequence of the rapid evaporation and expansion of the coolant. Whereas the volume of a gas increases linearly with the temperature (approximately a factor four by heating from 100°C to 1300°C), a simple calculation of the ratio of the two specific volumes shows that evaporation causes the volume of water to increase by a factor of 1700 at 100°C and ambient pressure (Table 1 and equation (1)), a number that was also found in literature (George, Nexhip, George-Kennedy, Foster, & Walton, 2006). The volume change occurs so fast that it resembles an explosion, hence the name vapour explosion. Such an explosion at industrial scale can cause economical losses, bodily injuries or even fatalities.

Jan 1, 2019

By Hökmark. H., M. Lönnqvist

"The Swedish Nuclear Fuel and Waste Management Company are planning to construct a permanent storage facility for high-level nuclear waste at about 500 m depth in fractured crystalline rock. One potential concern for the long-term performance of the repository relates to large fracture shear displacements, since these may damage intersected waste containers. Although details in fracture surface geometries are known to influence slip magnitudes and distributions, fractures are commonly approximated as planar features with lab-scale properties due to lack of relevant field-scale data and to limitations in available modelling tools. The purpose of the work described in this paper is to explore the possibilities to use the distinct element code 3DEC to generate and analyze models with isolated, non-planar fractures and, based on the results of such models, get perspectives on the influence of basic disturbances such as ridges, steps and waves. Given the model set-up, fracture segments with orientations deviating from the critical one will contribute to stabilize the fracture. As a result, all analyzed non-planar fractures slip less than the critically oriented planar fracture. We demonstrate that the angle of deviation is the most influential parameter determining the slip magnitude for comparable surface geometries. Similarly inclined steps and ridges have almost equal slip reductions despite differences in surface area restricting the movement. Although it may be possible to translate the influence of such irregularities to an effective coefficient of friction that could be used as input to models with idealized planar fractures, additional analyses with randomly shaped and distributed asperities, different fracture sizes and with alternative stress loads, need to be carried out to devise a defensible strategy for this type of parameter variation. INTRODUCTIONThe Swedish Nuclear Fuel and Waste Management Company’s (SKB) concept for deep geological storage of high-level nuclear waste is the KBS-3 method according to which the waste will be encapsulated in copper canisters and deposited, surrounded by compacted bentonite clay, in vertical deposition holes at about 500 m depth in fractured crystalline rock (SKB 2011). A site has been selected in the Forsmark area, which is located some 120 km north of the Swedish capital Stockholm. One potential concern for the long-term performance of the repository that has been identified by SKB (SKB 2011) is related to large shear displacements along repository host rock fractures induced by a nearby earthquake, since these may cause mechanical damage to intersected waste containers. Due to lack of relevant data at the appropriate field scale and to limitations in available modelling tools, fractures are commonly approximated as planar features with lab-scale properties. The standard numerical tool, 3DEC (Itasca 2013), used in SKB’s rock mechanics analyses, for example, generates planar fractures by default. Results from dynamic 3DEC earthquake-models (Fälth et al. 2010) currently provide input to layout decisions, i.e., the diameter of the smallest planar fractures that must be avoided when deciding on waste container positions at different distances from potential earthquake faults. It is, however, well known that fracture surfaces are rough or undulating on a wide range of scales (e.g., Candela et al. 2012) and that such deviations from planarity influence the mechanical behaviour of the fractures (e.g., Dieterich and Smith 2009, Marshall and Morris 2012, Ritz and Pollard 2012). It will probably never be a realistic option to attempt to represent the geometrical complexity of a population of fractures of different size, shape and surface structure in any large-scale rock mechanical analysis relevant for the performance assessments of the nuclear waste repository. It may however be possible to find approximate ways to represent different styles and degrees of non-planarity by assigning mod"

Jan 1, 2015

To produce high quality castings, one must monitor the melt quality by evaluating such parameters as the grain refinement and silicon inoculation and modification during the production of the commercial casting. Thermal analysis is a tool one can use for both monitoring and providing input to the process control of the operation. The current technique of performing thermal analysis in commercial foundries is to pour a relatively small quantity of the molten aluminum alloy into a sampling cup and to then acquire the cooling curve. This approach is somewhat simplistic in that one has limited control of the cooling rate and hence the solidification rate of the sample. Moreover, this process requires a fair amount of user intervention and effort. In the past, studies have been conducted in an attempt to develop a system for conducting thermal analysis in-situ. In a study at McGill University, an alkali based heat pipe system was developed for conducting thermal analysis in-situ. While the concept was deemed as very desirable, the complexity of the equipment and procedure was a substantial negative component. In order to improve the technique of accomplishing real-time thermal analysis with an in-situ probe, a new method that is based on McGill heat pipe technology was developed. In the new system, a controllable and flexible water based heat pipe probe is used to solidify a small quantity of metal while residing in the melt. The solidified aluminum is remelted after completing the thermal analysis. This is a key feature of the probe. It resides permanently in the melt yet it is able to both solidify a sample of the melt and then to subsequently remelt it. Various solidification rates encountered in commercial castings are attainable by adjusting the heat extraction rate of the unit.

Jan 1, 2004

By P. M. Pearson

This paper describes the principles ()f systems engineering and their application to the various transport problems confronting the mining industry. The organization of mining operations into sub-system design problems is explained.

Jan 1, 1970

By J. E. Rowland

NIOCALITE is a new mineral species which has been identified in rock from the Oka district, Quebec. It occurs as prismatic crystals up to l 0 mm. in length and I mm. in width, in coarse, white, crystalline calcite with associated apatite, diopside, biotite, pyrochlore, and niobian perovskite. The crystals of niocalite are light yellow in col-our and have a vitreous lustre. The optical properties of niocalite (I) are as follows: Biaxial negative -2V = 56° a = 1.700, [ ] = 1.730 Birefringence y -a = 0.030 Optic orientation: X = b, Z/\c 12°, Y , /\a = 32° The specific gravity is 3.32 and the hardness is approximately 6 (Mohs scale). A detailed analysis performed on 2.4 gm. of hand-picked material reveals that niocalite is essentially a calcium niobium silicate. Table I gives the result of the analysis. Crystals of niocalite, apparently suitable for X-ray diffraction work, were selected, and single crystal photographs were obtained using both Weissenberg and Precession cameras. Initial difficulties were encountered in interpreting these films, and in correlating the crystal data with the optical data. The in-consistencies were explained by the fact that every grain examined was twinned to some extent.

Jan 1, 1957