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By R. P. van Hille, N. Mooruth

Remediation of acid rock drainage (ARD) was performed within a linear flow channel reactor (LFCR) via a floating sulphur biofilm (FSB). A full sulphur species material balance was conducted in order to characterize the sulphide oxidation. The reactor was able to achieve an 82% and 93% conversion of sulphide with a hydraulic residence time of 2 and 5 days respectively. 92% and 55.5% of the converted sulphide reported to the biofilm as sulphur for the 2 and 5 day residence times respectively. The mass transport limitations within the biofilm were identified via a dissolved oxygen gradient profiling system.

Jan 1, 2011

By V I. McCallum

In 1954, Falconbridge Nickel Mines Limited began the program of deepening the Falconbridge mine workings for 2,000 feet below the bottom, or 4025, level in order to provide twelve new levels for production. Involved in this program were a production-sized internal shaft, .crushing, pumping and hoisting installations, the integration of the new shaft operations with those of existing shafts, and considerations of ventilation requirements and ground support at depth. This paper describes the major items of equipment, and the depth development of No. 9 shaft as a whole, and also emphasize the planning phases of the program. The factors involved and reasons for the equipment and layouts chosen are given. Introduction and History THIS paper describes the planning and development of that portion of the Falconbridge mine between the 4025 and 6050 mining levels. This mine, the oldest of several being operated by Falcon-bridge Nickel Mines Limited in the Sudbury district, has been a continuous producer for over thirty years.

Jan 1, 1964

By K. Sugawaral, W. Tongampl, K. Hiruta, Takasaki Y., A. Hosoi, T. Katol, A. Shibayamal

Printed circuit boards of discarded personal computers, televisions, mobile phones and so on contain a significant amount of valuable metals. For the recycle of valuable metals contained in the printed circuit board wastes (P CB W), the effect of temperature on chlorination-volatilization of metals under chlorine gas flow was studied. The experimental sample used was powder of two types: crushed PCBW (Sample 1) made from PC and industrial machines, and crushed-combusted PCBW (Sample 2) made from Sample 1. The principal element in Sample 1 and Sample 2 was copper, about 13 wt% and 20 wt% respectively. Sample 1 was composed of pure metals such as copper, tin and lead. In comparison, Sample 2 consisted of oxide metals such as copper oxide, tin dioxide and silicon dioxide. The two samples were heated at predefined temperatures (ranging from 300 °C to 900 °C) in chlorine gas. The chlorine gas flow into the furnace was set at a rate of 100 mL/min. As a result, copper in Sample 1 began to volatilize at 600 °C and completely vaporized at 700 °C. While in Sample 2, copper started to volatilize at 600 °C and reached 100% vaporization at 800 °C. For Sample 1, zinc finished volatilizing at 500 °C, while for Sample 2, zinc volatilization reached only about 80% (maximum). Also, lead, nickel and titanium showed similar results. Consequently the volatilization rate of Sample 1 is higher than Sample 2 and the temperature dependency was verified for copper, lead, zinc, nickel, antimony, titanium and chromium. In fact, the volatilization rate of these metals increased with the temperature. This result reflects the difference between pure metal and metal oxide, as Sample 2 was combusted before the chlorination-volatilization test. Moreover, the chlorination reaction may be accelerated by a reductive atmosphere because Sample 1 contains carbon. In this study, the volatilization behavior of some valuable metals in PCBW by chlorine gas was observed. Further improvement in the rate of volatilization can be achieved by optimizing the experimental conditions.

Jan 1, 2011

By W. Rostoker

Packing Experiments The porosity and most of the other physical properties of a coherent mass of any material built up by the packing together of discrete particles will obviously vary with, and be governed by, the manner of association of the particles. The relation of packing to powder metallurgy has been discussed in some detail by at least two authors. Packing Theory Theories of particle packing have been put forward by C. C. Furnas and A. E. R. Westman. Although their approach to the problem is not the same, their conclusions are similar. In the process of packing, they visualize the voids of a uniformly sized material becoming filled with particles of a smaller size, and the voids between these being filled in turn with still finer particles. Obviously, if this process is carried ad infinitum, the resultant mass will have the same density as the constituent particles. In practice, of course, such a density of packing is never attained. In addition, the premise is made that successive particle sizes must fit into the voids left by the immediately larger particles- an assumption requiring definite size gradations and also particle shapes. It remained, then, for theory to correlate relative diameter ratios, and proportions and numbers of sizes of particles, for requisite packed densities. Briefly, the following principles were postulated: (a) For an infinite diameter ratio between largest and smallest sizes, an infinite number of particle sizes in proper proportion is required to give natural density. (b) For a given diameter ratio, there is a maximum number of permissible particle sizes for maximum packed density (Furnas). (c) Provided the particle shape is spherical, or approximately so, the proportions o particle sizes can be calculated. (d) The ratios of successive diameters of particles are in geometric progression.

Jan 1, 1947

By H. T. Kennedy, D. L. Flock

"The project described in this pa-per was undertaken to provide a more thorough understanding of the effect of contamination by drilling fluids on the electrical properties of porous systems. The basic work per-formed was filtration tests on both consolidated and unconsolidated sands and the determination of their resistivity profiles with the aid of screen-type electrodes spaced at convenient intervals along the core.Four drilling fluids were selected, ranging in resistivity from 447.0 ohm-cm. to •5.76 ohm-cm. The permeability of three unconsolidated and four consolidated cores, on which resistivity measurements were made, varied from 1,257 md. to 4,560 md. Five additional cores with permeabilities as low as 49 md. were tested to investigate the effect of permeability on the depth of invasion of drilling fluids.The effect of drilling fluid invasion on resistivity profiles was determined on cores containing brine only, and on those containing 80 to 90 per cent oil and 20 to 10 per cent brine. The study included measurements of (a) clay content of different sections of the core, (b) saturation of the invaded fluids and oil in different parts of the core, and ( c) changes in resistivity as a result of gravitational re-distribution of fluids and diffusion of electrolytes.The data showed that the resistivity of an invaded core is, as was expected, controlled by the resistivity of the filtrate, the resistivity of residual core water, the resistivity of the clay particles, and the amount and distribution of the filtrate, clays, residual oil, and brine. In most cases there was a significant increase in resistivity owing to the presence of clay solids which had invaded the core. The change in resistivity, as a result of the diffusion of salts, was substantial after a period of five hours. There is evidence in some cases that the residual oil and core water saturations of the unconsolidated cores, after substantial invasion, are very low, indicating almost complete displacement by invading fluid."

Jan 1, 1959

By E. Latorre, T. S. Golosinski

ABSTRACT; his paper proposes new methodology to define the pit limits that yield the maximum net present value (NPV) for a deposit, under a given economic scenario and for a given discount rate. The method uses an algorithm that combines dynamic programming and heuristics. It was tested for evaluating deposits with different sizes and various grade distributions. For the discount rate of zero, the results confirm that the economic pit limits are the same as that defined using Lerchs Grossman method. However, when discount rates higher than zero are used, the pit limits differ, yielding a higher NPV.

Jan 1, 2011

By P. H. Riordon

ONE OF the major sources of the world's supply of asbestos lies in a belt of ultrabasic rocks stretching through the Eastern Townships of Quebec where chrysotile asbestos has been mined continuously since 1878. The productive portion of the belt extends from the town of Asbestos in Shipton town-ship northeastward some 60 miles to the eastern edge of Broughton town-ship. (See map). Twelve mines, currently treating an aggregate of over 60,000 tons of rock daily, produce more than one million tons of asbestos :annually. An equal number of mines are now dormant, having for the most part exhausted their reserves. It is evident from the past and present activity, in this belt that it can look forward to at least half a century or more of large scale production. The host rock to the asbestos de-posits is essentially a peridotite which originally contained I 0 to 15 per •cent pyroxene. The peridotite, associated with dunite, pyroxenite

Jan 1, 1962

By A. Trusiak

"Increasing energy costs in mineral processing is making it necessary for tighter controls over various parts of the concentrating process. Continuous monitoring and control of critical process streams can be easily performed by today's many computer systems. Indeed, the list of successful applications of computers to grinding and flotation control continues to grow. The day is rapidly approaching when the use of computers within concentrators for either process control or simply metallurgical accounting will no longer be a luxury, but a necessity. The benefits from such applications far out weight the initial investment.This paper describes the current application of a process control system at Brunswick Mining in north-eastern New Brunswick. Emphasis is placed on the control of a CuPb flotation circuit, and the benefits derived-from this control.No. 12 Cu-Pb Flotation: The complex nature of the Brunswick flotation process has been described else-where. 1,2,3 Four final concentrates are eventually produced, these being a Zn concentrate assaying 53% Zn, a 30% Pb concentrate, a 20% Cu concentrate, and a bulk Pb Zn concentrate assaying 35% Zn and 17% Pb. The Cu concentate, Pb concentrate, and bulk Pb Zn concentrate are all derived from a primary CuPb concentrate assaying 26% Pb, 11% Zn and 1.8% Cu.The No. 12 CuPb circuit produces approximately two thirds of the total primary CuPb concentrate. No. 12 ore, assaying 3.5% Pb, 9% Zn, 0.25% Cu and 100 g/tne silver, is ground in two parallel grinding circuits. Each line uses a 3.2 x 4.2 meter rod mill in closed circuit with two parallel 3.2 x 3.9 meter ball mills. Ball mill and rod mill discharges are classified in two stages of cycloning. The secondary cyclone overflow, or flotation feed, is controlled at 65% passing 37 microns."

Jan 1, 1981

By J. A. Brothers, M. R. Foran

"AbstractThis paper reviews the progress made in the manufacture of hydraulic cement from iron blast-furnace slag by simply grinding the granulated material with other constituents. The influence of the chemical and mineralogical composition of the slag, the conditions of granulation, the manner of drying, and the fineness of grinding, on the hydraulic properties of the cement, is discussed. The effect of additions of lime, portland cement, and anhydrite, alone and in combination, on the physical and cementing qualities of the mortar and concrete made from Sydney blast furnace slag, is shown in tables of data and is reviewed in some de- tail. The work at the Nova Scotia Technical College has shown that a cement equivalent in every respect to portland cement can be made from basic blast furnace slag- by the us2 of certain quantities of portland cement as an activator.IntroductionBlast furnace slag is a potentially valuable material for the manufacture of hydraulic cement. It is a by-product of the manufacture of iron in the blast furnace, and is formed by the combination of the silica and alumina .of the iron ore with the limestone flux. The metallurgist terms a slag .acid or •basic according as the percentage of lime is less or greater than the percentage of silica plus alumina. The basic slags have the •more pronounced hydraulic properties."

Jan 1, 1950

By M. Aljarrah

The phase equilibria in the Mg-Al-Ca system was investigated experimentally by differential scanning calorimetry (DSC) and X-ray diffraction (XRD) techniques similarly the Mg-Al-Sr system has been studied using DSC and XRD as well as scanning electron microscopy/energy dispersive spectrometer (SEM/EDS) and quantitative electron probe micro-analysis (EPMA). A new ternary solid solution MgxAl4,xSr was confirmed in Mg-Al-Sr system, which is a substitutional solid solution of Mg atoms in Al4Sr. Maximum solubilities of 21.3 at.% Al in Mg17Sr2 and 11.4 at.% Al in Mg were observed. It was also noticed that Mg38Sr9 dissolved 12.5 at.% Al. In the Mg-Al-Ca ternary system, one of the invariant transformations predicted by thermodynamic modeling was verified experimentally and found to occur at 512°C with composition close to 10.8 at.% Ca, 79.5 at.% Mg and 9.7 at.% Al. A ternary solid solution (Mg,Al)2Ca was confirmed where large solid solubility of Al in Mg2Ca was observed.

Jan 1, 2006

By J. D. T. Steyl

Anglo American Corporation and the University of British Columbia have developed a medium-temperature oxygen pressure leaching process for the extraction of copper from chalcopyrite concentrates in an acid-sulphate system (AAC/UBC process). Leaching is performed at 150°C and surfactants are added to disperse molten elemental sulphur from the unreacted sulphide surface to prevent liquid sulphur occlusion. Copper extractions of over 95% and elemental sulphur yields of around 60% were observed over a 2-hour average residence time in a continuous pilot autoclave. Sporadic blockages occurring under continuous conditions were found to be caused by the accumulation of molten elemental sulphur mixed with unreacted sulphide mineral. A batch programme of experimentation was undertaken to establish the reason for this phenomenon. Batch experiments were conducted at low pulp densities on concentrate samples with narrow size distributions and at similar solution compositions to those experienced in the pilot autoclave. The overall rate of chalcopyrite oxidation was shown, under approximately constant solution composition, to be a strong function of the total available mineral surface area. Initial copper extraction rates followed the shrinking particle kinetic model under intrinsic surface control. Deviation from shrinking particle kinetics was however observed after a certain oxidation time. Secondary electron images identified small (typically less than 20 µm) sulphur conglomerates at the point of deviation, which 'collected' unreacted chalcopyrite mineral particles and lowered the total mineral surface area available for oxidation. These sulphur clusters are thought to be the precursor for blockages experienced in the continuous pilot autoclave. These results emphasised the importance of fine grinding (P80 5-20 µm) and suggest that shorter residence times may minimise the formation of sulphur conglomerates in a continuous autoclave operating at medium temperature.

Jan 1, 2004

By E. Ghali, B. Arsenault

Aluminum alloys play an important role in the manufacturing of air and ground vehicles. However, with such alloys, corrosion damage is often found on structural components subjected to fatigue loading. The alloying provides large benefit in terms of the high specific tensile strength (ultimate strength/density ratio) to the aluminum alloy but causes the aluminum alloy to be more sensitive to localized corrosion due to complex microscopic heterogeneities such as secondary phases. This paper investigates thermal spray coating namely arc spraying to evaluate the protection against environmentally assisted cracking (EAC) and localized corrosion on Aircraft structural 7075 T651 aluminum alloys. In particular, EAC and pitting corrosion at the interface coating-substrate are a challenge for thermal spray protective coatings on aluminum alloys under cyclic load and immersion. To address these issues, different surface preparation techniques prior to arc spraying were considered such as chemical de-oxidation, laser ablation, shot peening and grit blasting. Results indicate that surface preparation and improved coating process parameters play an important role to provide a coating conferring a protection against both localized corrosion and EAC.

Jan 1, 2006

By P. Mundell, R. Torbin

"There are many physical and operational obstacles in pipelines that make the passage of inspection robots impossible including:•,Small radius bends•,Low operating pressure•,Plug valves•,Large changes in pipe size•,No permanent access to the pipeline.The inspection of unpiggable pipelines requires the marriage of a highly agile robotic platform with NDE sensor technology operating as an integrated system. Foster-Miller and GE/PII are developing a robot that is essentially a battery powered, train-like platform. Both front and rear tractors propel the train in either the downstream or upstream direction. Like a train, the platform includes additional ""cars"" to carry the required payloads. The cars are used for various purposes including the NDE sensor module(s), the power supply, and data acquisition/storage components. The robot is designed with a slender aspect ratio and the ability to change shape as required by the physical obstacle presenting itself. The MFL sensor module must also morph itself through the physical obstacles, and thus, will require some level of segmentation. The segmented MFL magnetizers and transformable sensor module represent a major advance in the science of sensor technology and pipeline inspection."

Jan 1, 2005

By G. J. Hawkes

"We all know that every man who works for us should have a thorough knowledge and understanding of his responsibilities. But how about ourselves? Du we have a thorough knowledge and understanding of our own responsibilities as Mill Superintendents?I once had an occasion to say to a Manager that I did not know something was my responsibility and he replied with suitable emphasis ""Everything inside those four walls is your responsibility"". That makes the extent of our responsibilities clear but does not give us a check list to keep an eye on as we do our job from day to day.What are our responsibilities? I suggest they are the following:1. Developing and maintaining efficient operating and maintenance crews.2. Milling the required tonnage.3. Recovering the economical maximum amount of gold.4. Keeping costs at the economical minimum.5. Seeing that the figures are reasonably accurate for tonnages, assays, production, losses, and costs.6. Preventing accidents.7. Preventing fires.8. Maintaining the equipment and building and keeping them clean and tidy.9. Enforcing company rules and regulations."

Jan 1, 1965

By S. Nakamura, K. Fujiwara, S. Hirosawa, S. Tsukamoto, Y. Koshino, Y. Aruga

"In this study, hardness test, differential scanning calorimetry(DSC)and transmission electron micros copy were carried out to investigate the age-hardening behavior of Al-Mg-Si alloy with a Mg+Si concentration of 1.5mass%, and a Mg/Si ratio of 1.0. Pre-aging treatment was conducted for the solution-treated and then water-quenched specimen at 100, 80, 60, 40, 20 (natural aging), 0 and -20°C for 20min or 5h, followed by artificial aging at 170°C for 20min. It was found that when pre-aging time is prolonged from 20min to 5h, the bake-hardening response is deteriorated irrespective of the pre-aging temperature. In particular, the specimens subjected to pre-aging at 20-40°C for 5 h hardly increased the hardness with the subsequent artificial aging. From the results of DSC measurement, it was suggested that as pre-aging temperature is lowered, the amount of solute clusters formed during pre-aging, i.e. Mg-Si clusters and Si-rich clusters, is decreased, but the ratio of Si-rich clusters is increased, resulting in the deteriorated bake-hardening response.INTRODUCTIONIn recent years, Al-Mg-Si alloys have been widely used forautomotive body panels due to their excellent characteristics such as high strength-to-weight ratio and good formability. Their strength is increased not only by artificial aging during paint-bake treatment; i.e. “bake-hardening”, but also when kept at room temperature(RT);i.e. “natural aging”, in association with the suppressed bake-hardening response. Such a “negative effect of two-step aging” has been a serious problem because Al-Mg-Si alloy coils are unavoidably maintained at RT in the manufacturing process of body panels; e.g. shipping and storage. However, it was reported that pre-aging above 70°C before the two-step aging can restore the bake-hardening response, and such an improvement is related to the formation of solute clusters(Sato, 2006; Abidet al., 2010; Takaki et al., 2014; Zandbergen et al., 2015). From the comprehensive analysis by three-dimensional atom probe (Aruga et al., 2014 and 2015), it is believed that Mg-Si clusters formed during pre-aging transform into the strengtheningß"" phase, whereas Si-rich clusters formed during pre-aging act as a detrimental precursor of ß""phase. In this study, it was aimed to investigate the effect of pre-aging conditions, especially pre-aging temperature and time, on the formation of solute clusters during pre-aging, in order to elucidate the mechanism of the age-hardening behavior of an Al-Mg-Si alloy."

Jan 1, 2018

By Robert White

?Introduction of BMS?Paste fill generalities?Development through paste?Development through rock fill?Conclusion ?Brunswick mine is operated by Xstrata. The mine has been in operation since 1964 and produces around 10 000 tons per day of ore containing (in order) zinc, lead and silver? The ore body consists of close to ten sub-parallel massive sulfide lenses. The length of the ore body is 1200 m with a width of up to 200 m. From the surface, the ore body extends to a depth to 1200m ?There is 100 km of active drifts for a total 400 km ?The mine is divided by 7 levels and 23 actives sub levels

Feb 1, 2008

By Pierre Sauvé, Guy Perrault, Armand Savoie, Pierre Trudel

"La Mine Doyon est une entreprise détenue à parts égales par Cambior Inc. et par Minerais Lac Ltée qui en est l'opérateur. Les statistiques de production au 1"" janvier 1990 indiquent que, depuis le début de son exploitation (1980), 10.3 Mt de minerai d'une teneur moyenne de 5,9 g Au/t ont été usinées. Les réserves sont estimées a 10,4 Mt à une teneur moyenne de 6,6 g Au/t jusqu' à une profondeur de-450m.Trois zones minéralisées sont présentement en exploitation. Ces zones sont caractérisées par la présence de veines de quartz-pyrite accompagnées localement de chalcopyrite. Les roches encaissantes sont les schistes a séricite (zone# 1), les roches volcaniclastiques de composition felsique (zone #2) et intermédiaire (zone #2, zone ouest) ainsi que les roches intrusives appartenant au pluton de Mooshla (zone ouest). Toutes ces roches ont subi un métamorphisme régional dont l'intensité correspond à la limite supérieure du facies schiste vert ( -425°C}.Les zones minéralisées sont situées à I' intérieur des zones de déformation et d'altération les plus intenses. L'étude des structures mineures nous indique que la région a subi plusieurs épisodes de déformation ductile a fragile qui sont antérieurs, contemporains et postérieurs al a minéralisation. L'analyse structurale indique également que la déformation majeure est associée à une importante phase de compression orientée N-S a NNE-SSW.L’ensemble des observations suggère que la minéralisation et I' altération associée sont de type épigénétique. La mise en place du minerai se serait effectuée à la fin de l'épisode principal de métamorphisme et de la déformation."

Jan 1, 1990

By Raymond S. Suglo

Increasing depths of mining coupled with the mechanization of most underground mining operations have led to growing complexities in the mine environments with the production of large amounts of gaseous and particulate contaminants. Over 100,000 lives have been lost due methane gas and dust explosions in coal mine workings in both the US and Canada since 1900 when the methane concentrations exceeded stipulated levels in some workings. There is therefore the need to constantly assess and evaluate the performance of the existing mine ventilation systems to maintain safe and acceptable mine environmental conditions in workings to meet the requirements of federal, state or provincial laws. This paper uses the results of continuous monitoring of methane concentrations conducted in selected development headings in coal mines in North America to assess the effectiveness of existing auxiliary ventilation systems in the control of methane gas concentrations in the headings. The results show that the average quantities of fresh air required to dilute, disperse and remove methane concentrations within set levels of one minute varied from 5.43 m3/s in Heading #11 (Condition 2) to 27.97 m3/s in Heading #10 (Condition 1) and this represented about 12 to 26% of the quantities required to dilute methane concentrations to safe levels within one minute under Condition 1 and 32 to 139% under Condition 2. The average dilution times in the headings studied were less than 8 minutes. The forcing efficiencies of the auxiliary systems in the headings studied ranged from poor to excellent implying that the auxiliary ventilation systems were capable of controlling the methane concentrations below statutory levels but may not be able to cope with large and unusual methane concentrations in the headings. This study is important to prevent methane gas and coal dust explosions in coal mines and thus avoid any loss of lives and property.

May 1, 2001

By A. G. Matyas, B. Wasmund, R. C. Francki, K. M. Donaldson

"The construction of large. high-power electric furnaces has been an important factor in the growth of electric smelting technology . New developments in the design and operation of electric furnaces have been spurred by the escalating cost of electric power and increasingly stringent environmental constraints. Key areasofprogress include the application of high voltage to enhance furnace productivity; the installation of novel furnace cooling elements and robust binding systems to extend furnace life; development of process schemes to reduce power consumption by preheating and prereduction of the charge; and environmental control improvements for handling fine dusty feedstocks, and for furnace sealing.This paper reviews these developments and outlines a systematic approach for their application in the custom design of new or retrofit electric furnace facilities.IntroductionThe use of electric smelting furnaces has continued to progress during the last decade. Electric furnaces offer high throughput capacity and productivity. In addition, they produce minimum environmental impact, as the furnaces can be readily sealed, and only a small quantity of off-gas is generated. Electric furnaces present the most favourable choice of smelting technology where high process temperatures in excess of about l400°C are required.Their application is burdened by the escalating cost of electric power, hence it is important to seek new opportunities through further technological development. This paper reviews recent developments in the design and operation of high-power electric smelting furnaces, in the following areas:"

Jan 1, 1993

By P. Hajiani

"It has been shown that the economics of the hydrometallurgy of a rare earth process depend upon the amount of acid needed to extract the metal per unit mass of the ore. In this paper, we show how to save a significant amount of acid treating a low grade ferrocarbonatite and extracting +99% constituent REEs. In order to save the acid, the alkali earth metals, originated from gangue minerals were removed prior to REE leaching. Moreover, the operating conditions of leaching reactor are optimized to leave iron behind and to leach REE selectively and effectively. Consequently, the hydrochloric acid consumption per ton of ore drops significantly compared to other processes.INTRODUCTIONThe viability of an extractive metallurgy process of rare earth elements (REEs) to produce an intermediate mixed concentrate depends upon a low-cost simple flowsheet (Bogner, 2014). It has been known that regardless of the mineral, the acid consumption is intimately associated with the operational cost (OPEX) of any REE project, particularly when the initial REE grade is low. Besides acid, the other reagents consumed in the process impact the OPEX. This makes minimizing acid consumption in an REE method a critical priority. The acid consumption in different projects has been reported to be equivalent of 100 kg to 1000 kg HCl (100%) per tonne of ore (Average: 500 kg/t HCl 100%) (Bogner, 2014). This value is strongly correlated with the grade of REE in the ore and therefore in the pre-concentrate.Previously, it has been stated that the Montviel flotation concentrate requires about 300 kg HCl (100%)/t ore to attain +99% REE recoveries using direct leaching (Yu, Verbaan, Pearse and Britt 2013). Hence, for an ore feed rate of 2500 t/d ROM, a large acid production plant (744 t/d HCl 100%) would be required for acid leaching of such concentrate. The equivalent Chlor-alkali plant is estimated to cost M$ 150 (CAPEX) and may require 100 MW power. Therefore, there is a necessity to improve the extractive hydrometallurgy and to reduce the acid consumption significantly."

Jan 1, 2016