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By P. Latostenmaa, O. Forsén, M. Lundström, J. Aromaa

The history of electrolytic copper production in Finland is closely connected with the industrialization of Finland and events of the Second World War. Production of cathode copper by electrowinning from the processing of Outokumpu ore started in 1913 based on the Hybinette nickel process, however due to the poor performance of this process, the plant was closed in 1929. Outokumpu commenced copper smelting at Imatra in eastern Finland in 1936 producing blister copper for refining outside the country. Due to the threating military situation, a new electrolytic copper refinery was built in 1941 at Pori in western Finland to refine the anodes. This was the first electrolytic copper refinery in Finland. It was built with 108 cells having a capacity of 12,000 tonnes/y. In addition, the smelter was relocated to nearby Harjavalta in 1944–1945. To fulfill the war reparations included in the peace treaty with the Soviet Union, a doubling of the smelter capacity was realised, with the Pori refinery capacity increasing to 20,000 tonnes/y in 1946 by adding additional cells. The cornerstone of Finnish metallurgical technology — Outokumpu flash smelting — was invented with the new technology commissioned at Harjavalta in 1949 replacing the electric furnace and providing a further increase in copper output with subsequent enlargement of the refinery. By 1963, the plant consisted of 380 cells with a capacity of 48,000 tonnes/y of cathode copper. The Pori electrolytic copper refinery has operated continuously since 1941. Commencing in 2004 the Pori refinery has been part of the Boliden group of Sweden. The most recent plant upgrade was in 2007 with the installation of permanent cathode technology. In 2017, the production of the Pori refinery was 133,000 tonnes of copper cathode, 2,300 tonnes of nickel sulphate, along with 66 tonnes of silver and 4.6 tonnes of gold. This paper describes these outstanding achievements. INTRODUCTION In 1809 the Kingdom of Sweden was defeated by the Russian Empire in the Finnish War, and Finland was annexed by the Russian Empire as the autonomous Grand Duchy of Finland. Finland’s commercial and industrial interests were developed to serve the Empire’s needs. Finland had raw materials, but at the start of the 20th century Finnish industry was based on the use of wood as raw material and agricultural products. The mining industry was only a few percent of the country’s industrial production.

Jan 1, 2019

By H. Matsuoka, M. Kawata, K. Mitsuhashi, C. Tokoro, T. Kato

Copper sulfide minerals such as chalcocite, bornite, and chalcopyrite exhibit high floatability, thus they can be recovered as a froth product by adding only a collector in flotation process. However, when their surface is oxidized, the floatability decreases and the simple addition of collectors is no longer effective. In this case, flotation activators are required to improve the floatability. Sodium hydrosulfide (NaHS) acts as an activator for chalcocite and bornite by re-sulfurizing their oxidized surface portions. Nevertheless, the flotation conditions still require optimization in order to achieve a satisfactory recovery of copper minerals. This work addresses the study of the effect of pH and NaHS dosage on the flotation of a complex Chilean copper ore containing chalcocite, bornite, and chalcopyrite. Flotation experiments were conducted using a Denver-type flotation cell and by investigating the effect of two NaHS dosages (50, 100 g/t) and three pH values (8, 9, 10). Following this process, the surface interaction between the minerals and NaHS were analyzed using zeta potential measurements. INTRODUCTION In recent years in the mining area of metal resources, the complexity of ores and the fineness of grind required to liberate minerals have become serious problems. When ores are complex, several valuable minerals may interact with each other and reactions may become more complicated compared with an ore containing only one valuable mineral. Therefore, more processing steps may be required than for a simple ore. When the valuable minerals exist as fine grains, it is difficult to liberate them and more grinding time is required. Furthermore, the probability of collision of fine particles with bubbles in the flotation process declines and it takes much time to recover them. Flotation is one of the most applied methods to separate valuable minerals from unwanted ones using the difference of hydrophobicity (or hydrophilicity). In the flotation method, the air is fed to a cell filled with the pulp of the mineral of interest and agitation is implemented to introduce air and generate bubbles. The hydrophobic particles attach themselves to the bubbles and float along with the bubbles while the hydrophilic particles remain in the pulp. Copper sulfide minerals such as chalcocite (Cu2S), bornite (Cu5FeS4), chalcopyrite (CuFeS2) are widely known abundant copper resources. Generally, flotation is used to recover and concentrate these minerals. Copper sulfide minerals usually have high floatability. However when their surfaces are oxidized, their floatability becomes lower. When the sulfide surfaces are oxidized, activators are added before flotation to modify the surface property of the Copper sulfide minerals so that they can be floated. Sodium hydrosulfide, NaHS, is known as one of the activators. Zhao (2018) reported that NaHS is often used as a sulfurizer for oxidized sulfide minerals (Becker (2014), Clark (2000), Newell (2006), and Newell (2007)). In earlier work, the authors performed flotation tests on a sample of Chilean copper ore which contained chalcocite, bornite, and chalcopyrite using a Denver-type flotation apparatus. The sample was ground to P80 70 µm with a rod mill and ball mill. In these tests, the following conditions were used: natural pH (about pH 8.5), flotation time of 45 min, 60 g/t AP3477 (sodium diisobutyl dithiophosphate) as collector, and MIBC (methyl isobutyl carbinol) at appropriate dosage as frother. NaHS was applied at the dosage of 0, 67, and 120 g/t, respectively at the conditioning. The concentrates were collected at 6, 18, 30, and 45 min. The weights of the concentrates and tailing were measured and they were assayed for copper by ICP-AES (Optima 5300 DV, PerkinElmer Inc.). The flotation recoveries of the copper minerals were calculated from the results of MLA (mineral liberation analyser) (MLA 650, FEI) measurements.

Jan 1, 2019

By S. Kevin McNeil, William Joa, Catherine Johnson, S. Omar Garcia

The geometry of an explosive is known to have a fundamental effect on the resulting shock wave propagation. Typically researchers use a spherical or hemispherical geometry in order to simplify the shock wave so that resultant overpressure and impulse are uniform in all directions. The goal of this research is to understand the overpressure characteristics from a sheet explosive, a geometry often utilized for breeching in the military. In training alone, soldiers can be exposed to several tests per day, which may be contributing to the prevalence of blast induced mild traumatic brain injuries. Using PCB pressure transducers and high speed video, it was found that as the angle departed from normal (perpendicular) to the face of the sheet explosive, the over pressure decreased significantly. When in line with the plane, a decrease of over pressure by as much as 20% was observed. Time of arrival was roughly 20% faster directly in front of the charge with a pressure amplification of 46% relative to a spherical baseline charge of the same mass. This testing may reveal a more complete relationship between the pressure and angle from the explosive that could have applications for military and law enforcement with respect to best practices on blaster location for reduced over pressure exposure.

Jan 1, 2019

By T. Hidayat, D. Shishin, S. Sineva, M. Shevchenko

The effects of the slagging components Al2O3, CaO and MgO on the thermodynamics and phase equilibria of the gas/slag/matte system with major elements “Cu2O”-“FeO”-SiO2-S are characterised using an integrated experimental and thermodynamic modelling approach. The experimental technique included high-temperature equilibration experiments using a spinel substrate in controlled gas atmosphere (CO/CO2/SO2/Ar), quenching of the sample and subsequent measurement of the compositions of equilibrium phases with Electron Probe X-ray Microanalysis (EPMA). Thermodynamic modelling is undertaken using the computer system FactSage. In previous experimental studies on this system, the separate effects of Al2O3, CaO and MgO were determined. In the present investigation, the combined effects of these additional components are investigated at 1200oC and P(SO2) = 0.25 atm and a range of P(O2)’s. Special attention is paid to ensuring chemical equilibrium is achieved in each case, for this reason possible kinetic factors are systematically analysed. The results obtained are presented in the form of graphs showing the compositions of matte and slag phases as functions of copper concentration (matte grade). The new experimental data are used for the improvement of the thermodynamic database for multicomponent copper-containing systems. The effects of the slagging elements on this multicomponent system will be presented in the paper. INTRODUCTION Whilst Fe-O-Si form the principal components of copper smelting and refining slags, additional impurities are present in industrial practice typically at 2–5 wt.% Al2O3, 1–4 wt.% CaO and 1–2 wt.% MgO levels (Davenport, King, Schlesinger, & Biswas, 2002). These elements are introduced into the slag as impurities in concentrates and other feedstocks, in fluxes added to the process and in refractory materials. Understanding, and accurate description, of the changes that take place in the smelting process in the presence of these additional elements is of major importance to the optimization of industrial operations. Although experimental data are available on slag/copper matte equilibria, further work is needed to accurately characterise the influence of Al2O3, CaO and MgO on phase equilibria and elemental distributions of elements in gas/slag/matte systems. The individual effects of the Al2O3, CaO and MgO slagging components on equilibrium phase assemblages under selected conditions have been quantified in a recent series of studies on the Cu–Fe–O–S–Si-(Al, Ca, Mg) gas-slag-matte-metal system (Fallah- Mehrjardi, Hayes, & Jak, 2018; Fallah-Mehrjardi, Hidayat, Hayes, & Jak, 2017a, 2017b; Hidayat, Fallah-Mehrjardi, Hayes, & Jak, 2018a, 2018b; Jak et al., 2016; D. Shishin, Decterov, & Jak, 2018; Denis Shishin, Hayes, & Jak, 2018). Since, thermodynamically, the condensed phases do not behave ideally, it is to be expected that the effects of individual elements may not be additive, and the additional interactions between the Al2O3, CaO and MgO slagging components will also have to be taken into account. These additional effects are expected to be smaller than those of the individual components, so higher accuracy is needed. The present paper describes analysis of the precision and accuracy of the experimental technique on multi-component, multi-phase equilibria in these systems to enable the combined effects of Al2O3 + CaO + MgO on slag/matte/spinel equilibria, i.e. for high Fe/SiO2, to be accurately measured. The effects of Al2O3, CaO and MgO on the gas - slag - matte are analysed using a thermodynamic modelling approach with the computer package FactSage.

Jan 1, 2019

By S. K. Kawatra, V. Claremboux, S. Valluri

Carbon dioxide capture from coal-fired power plant flue gas and subsequent sequestration is expected to play a vital role in mitigating global climate change. To meet the growing need for CO2 capture Technology, Michigan Technological University is researching CO2 capture and Utilization. We have successfully converted CO2 into oxalic acid using an electrochemical approach. If oxalic acid can be produced in bulk from CO2, it can be used to leach rare earth minerals at low cost. Electrochemical reduction of CO2 to hydrocarbons and other chemicals is a complex multistep reaction with adsorbed intermediates. The exact reaction mechanisms leading to various products are not clear from the literature to date and will likely change over the range of conditions like pH, electrode potential, electrolyte medium, catalyst, etc. In this paper, we will present an idea of how the factors mentioned above will effect the electrochemical reduction of CO2 to Oxalic acid. INTRODUCTION. CO2 is thermodynamically quite stable, as shown by its highly negative heat of formation. Thus, it is expected that the formation of any useful chemical from CO2 will require the input of at least as much energy as geological sequestration. However, to be more effective the resulting economic value of the chemical need only pass the cost of the additional energy required beyond geological sequestration. This lends itself to two extremes: one where the quantity of energy required is low and one where the economic value of the additional energy is low. It is expected that CO2 can be reduced via electrolysis to several compounds. Of particular interest are formic acid, oxalic acid, methanol, ethanol, formaldehyde, and carbon monoxide (as a component to syn-gas). These chemicals have been reported as electrolytic products previously in literature and have considerable potential for application. EXPERIMENTAL Electrocatalytic Production of Oxalate from CO2 We are using membrane electrolysis cell to produce oxalate from CO2. A membrane electrolysis cell is a 2-chamber electrolysis cell where the chambers are separated by a selectively permeable membrane. To produce oxalate, a cation exchange membrane (which selectively exchanges cations) was used. Previous literature suggests that in the cathode chamber an organic electrolyte, such as tetraethylammonium perchlorate or tetraethylammonium bromide in dimethylformamide, are preferred3. In the anode chamber, a sodium hydroxide solution water is used. Carbon dioxide is then bubbled into the cathode chamber as a current is applied.

Jan 1, 2019

By M. Schmidt, P. Nord, M. Larinkari

Cathode stripping machine technology has been developed significantly both in electrorefining (ER) and electrowinning (EW) applications in the last ten years. The main driver has been robotic cathode handling, which has enabled increased flexibility to meet different layout and material handling requirements for both ER and EW processes. Additionally, cathode stripping stations have been continuously developed to optimize speed and to reduce mechanical stress of permanent cathodes. 'taco' type cathodes have been proven as beneficial approach for high capacity operations (particularly in Cu ER), where over 600 cathodes per hour continuous throughput has been achieved. However, the latest developments also apply to split type cathode stripping. Besides material handling, the modern cathode stripping machine has also an important role to provide information for production and quality monitoring. This paper describes the latest developments from the field of copper cathode stripping based on several different projects implemented in the recent years. INTRODUCTION After introduction of the permanent cathode technology in the late 1970s, it became the prevailing technology in copper refineries over the traditional starting sheet process by the end of 1980s. The main driver was less labour-intensive production due to fewer processing steps required along with more automated electrode handling equipment (cathode stripping machine, anode preparation machine, anode scrap washing machine and tankhouse cranes). Since then the design of the cathode stripping machine has followed the development of permanent cathode technology. The main development steps so far can be summarized as follows: Full Deposit Stripping Machine (FDSM) based on waxed permanent cathodes or cathodes with bottom edge strips resulting in split cathodes (copper sheets separated, see Figure 1) Wax less cathode stripping based on V-groove on bottom of the edge strip resulting in either ‘taco’ cathodes (enveloped copper sheets, see Figure 1) or split cathodes Full Deposit Robotic Stripping Machines (FDRSM) using 1–5 robots for permanent cathode handling and cathode stacking.

Jan 1, 2019

By K. V. Raj, M. McNinch, A. L. Miller, J. Wilson, D. A. Parks

Diesel particulate matter (DPM) has been shown to contribute to various adverse health effects on underground miners. In order to reduce worker exposures, it is critical to measure the levels of DPM in the active work settings and in real time. However, the 5040 method, developed by the National Institute for Occupational Safety and Health (NIOSH), which quantifies the mass of elemental carbon (EC) and organic carbon (OC) in the samples, is based on full-shift samples, and can take days or even weeks for results. Use of real-time monitors could significantly reduce the risk of DPM exposure to miners, and available monitors have been shown to quantify DPM, mainly using the EC and other surrogates. The NIOSH Spokane Mining Research Division is conducting research aimed at developing a field-portable DPM monitor that is able to quantify both EC and OC. One part of this work is to understand the effect of OC on currently available real-time DPM monitors. This paper will present the results from experiments that were designed to observe the effect of OC on real time monitors and compare the results with the NIOSH 5040 method. BACKGROUND Diesel engines remain ubiquitous in mining environments and, as such, are one of the primary sources of ultrafine- and nano-aerosols in the mining environment in the form of diesel particulate matter (DPM). As a result of their small size, DPM has longer residence times in a mine atmosphere than the typically larger mechanically generated dust [1]. In addition, the small size enables DPM to penetrate deeply into the respiratory system [1], and exposure to DPM has been shown to contribute to various adverse health outcomes of the pulmonary system [2] and cardiovascular system [3]. Diesel exhaust has been classified as a carcinogen to humans (Group 1) by the International Agency for Research on Cancer [4]. Due to mounting concern about adverse health outcomes, extensive efforts are being made to reduce workers’ exposure to DPM, with the high exposures that often occur in underground mines being of particular concern [5]. New technologies are being developed to reduce engine emissions and thereby airborne DPM levels, including advanced engine technologies, a host of aftertreatment systems [6], and a variety of alternative fuels. The latter have been shown to markedly decrease respirable DPM and other hazardous components of diesel emissions from mining vehicles [7], yet high exposures persist, especially in underground mines [8]. In order to reduce worker exposures, it is critical to measure the levels of DPM in active work settings. In mine settings, exposures to DPM are measured by collecting air samples on filters (usually full-shift samples) and sending them to a lab for analysis using the Mine Safety and Health Administration (MSHA) P13 method, a variant of the National Institute for Occupational Safety and Health (NIOSH) 5040 method. The 5040 method is designed to quantify the mass of elemental carbon (EC) and organic carbon (OC) on the samples [9], which is then used to determine the concentration in the sampled air. MSHA has set a limit for exposure, which is based on the concentration of total carbon (TC), or the sum of both EC and OC. The allowable TC in the air (time-weighted average over the shift) is 160 μg/m3 of TC [10].

Jan 1, 2019

By Esa Peuraniemi, Hanna-Leena Heikkilä

Boliden Harjavalta has been operating a copper flash smelter since 1949. Thus, in 2019, the smelter is celebrating its 70th anniversary of flash smelting. Slag formation is an integral part of the smelting process; currently the processing of slag is based on froth flotation at the slag concentrator. It provides high recovery of copper but restricts slag utilization; the final copper slag is deposited in a landfill. The Harjavalta smelter commissioned a new landfill area for final slag in May 2018. This paper describes the full project from permitting to commissioning of a new landfill area. The permitting process is described comprehensively, including environmental impact assessment together with the environmental and dam safety monitoring. During the environmental impact assessment, observations on rare animals protected on European level were made as well as archaeological findings. Successful completion of the permitting allowed the construction phase to start on time. After six years of intensive work, the new landfill was commissioned in May 2018. The new landfill area provides a sustainable way to store the smelter’s slag for decades to come. INTRODUCTION Slag processing Sulphide based concentrates have always formed the main raw material stream at Harjavalta’s copper smelter. During the smelting and converting steps, the iron included in the concentrate is oxidized and fluxed using a silica-based sand. Originally, the slag was processed by using an electric slag cleaning furnace after which the final slag was water granulated. Froth flotation has been utilized for slag treatment since 1966. In the beginning, the slag was cooled in open ditches. In 1984, controlled slow cooling of slag was licenced to enhance copper recovery at the flotation plant. Today, the slag from both the smelting and converting is tapped/skimmed to slag ladles and treated in the following steps: Slag ladle transportation from the smelter to the cooling area by slag hauler Slow cooling of slag in the ladle to allow mineral precipitation Unloading the slag by tilting the ladle; causing the slag to crush as it drops from the ladle Feeding the crushed slag (-500 mm) by wheel loader to a slag concentrator’s feed funnel followed by a conveyor belt transporting the slag to the concentrator plant Crushing and grinding using lump and pebble mills Froth flotation of the valuables Recycling the slag concentrate back to the smelter feed after dewatering Pumping the final copper slag to the landfill area.

Jan 1, 2019

By K. Kryst, K. Elkayar

The increasing need to process lower grade contaminated copper concentrates and simultaneously produce higher purity copper cathodes necessitates a more diligent control in the electrorefining process. Ion exchange processes can be a powerful tool in controlling electrolyte impurities to ensure trouble free operation within the tankhouse and to help maintain cathode purity. In addition, some impurities can be recovered as a value added product. The use of a short bed ion exchange process provides a unique opportunity for impurity management in copper refineries. Several metals are discussed in this paper, including Sb, Bi, As and Ni and their methods of control and recovery within the copper electrolyte circuit. INTRODUCTION To produce a high purity copper product by electro-refining a copper electrolyte, it is important to control the concentration of impurities such as arsenic (As), bismuth (Bi), antimony (Sb), and nickel (Ni). These impurities are introduced into the electrolyte from the copper anode and can affect the copper product and the production operation. The presence of As in the copper product results in a lower electric conductivity and reduced workability of the copper. High levels of Bi in the electrolyte will result in a copper cathode product containing Bi. High Sb concentration will cause undesired floating slimes. A variety of methods have been developed to control the level of these impurities in the copper electrolyte including precipitation, liquid extraction, and ion exchange. These impurities can be recovered and used. This paper discusses novel methods that utilize short bed ion exchange technology to control these impurities. Also, the use of ion exchange to recover acid from the liberator cells bleed for reuse in the tankhouse is reviewed. SHORT BED ION EXCHANGE Short bed ion exchange technology has been employed extensively in the metal finishing industries for over forty years. Unlike conventional ion exchange, short beds are fixed beds without freeboard space. The application of short bed technology results in significant benefits when compared to conventional ion exchange. The resin used in this technology has a significantly smaller diameter than conventional ion exchange resin. As a result of the smaller resin size, exchange kinetics improves dramatically. This facilitates operation at higher flow rates and the use of smaller beds with a small mass transfer zone length. Smaller beds are especially important for applications that use chelating resins due to the high cost of the resin.

Jan 1, 2019

By M. Pelser, H. Gous, J. H. W. M. Brits, K. Voogt

The effect of impurities; iron and selenium, on current efficiency and deposit quality in copper electrowinning was investigated using an at-line cell at the Rustenburg Base Metals Refiners’ Copper Tankhouse. The tankhouse is a direct leach, starter sheet electrowinning operation with a unique electrolyte composition. The effect of iron on the current efficiency in the at-line cell was found to be less severe than reported in literature. However, the distinction between the chemical impact of iron on current efficiency and the physical impact of iron on a starter sheet operation needs to be realised. The effect of selenium on current efficiency and deposit quality is also reported. INTRODUCTION Anglo American Platinum’s Rustenburg Base Metals Refiners’ (RBMR) Copper Tankhouse serves a unique function in the refinery flowsheet in that its primary functions are the removal of copper and production of acid for the leach circuit. This function needs to be performed to specification in order to keep the leach circuit operating optimally. Current efficiency is seen as an indicator of tankhouse performance and efficient energy use. The tankhouse plates starter sheets onto titanium blanks which, after stripping and preparation, return into the production cells for further plating. Approximately 78% of tankhouses surveyed (Robinson et al., 2013) use permanent cathode technology whereas older, smaller operations still use starter sheet technology. Drawbacks of the starter sheet technology are that the starters are thin and bend easily leading to increased shorts, they need to be removed after two days of plating to be manually straightened and are then returned to the cells. In addition, the electrolyte levels in the cells need to be adjusted during the plating cycle to improve loop adhesion to the cathode and reduce loop corrosion. The starter sheet technology is therefore very labour intensive and the final quality of the product is dependent on the quality of the starter sheets produced. In comparison, permanent cathode or direct plating technology lends itself to automation, allows for higher current densities and better cathode quality due to more robust geometric stability in the cell. The present barrier to implementation of permanent cathode plating at the RBMR Copper Tankhouse is the capital cost required to convert the tankhouse, i.e. blanks and stripping machine. At present, alternatives to titanium blanks, such as LDX 2101, are being investigated for the unique electrolyte composition of this tankhouse.

Jan 1, 2019

By X. J. Guo, Q. K. Wang, S. P. Zhong

With depletion of traditional clean copper reserves and fast expansion of the global copper demand and smelting capacity in the past decade, more and more copper smelters have to treat copper concentrate with increasing content of complex arsenic-bearing minerals. Developing economically viable and environmentally reliable methods and approaches for arsenic control in copper smelting has become more critical for sustainable growth of the copper industry. This paper will report the recent effort of Zijin Mining Group Ltd. in developing processes for improving arsenic control and management in the Zijin Copper Smelter. The arsenic distribution and balance data for matte smelting, converting and refining operations in Zijin Copper Smelter is reviewed for understanding of the current challenges in arsenic control. The processes developed for treating the arsenic bearing waste acid bleed stream from the smelting-converting off-gases wet cleaning circuit and the arsenic sludge produced in the refining electrolyte purification circuit for arsenic disposal are discussed. INTRODUCTION With fast expansion of the global copper smelting capacity to meet the world copper demand and depletion of traditional clean copper reserves, more and more copper smelters have to treat copper concentrate with an increasing content of complex arsenic-bearing minerals. China is the world largest copper consumer and also has the largest copper smelting and refining capacity, but majority of its smelting capacity relies on supply of imported copper concentrate. In order to meet the increasing copper demand and ease the tight supply of clean copper concentrate, the Chinese copper industry has to look at the domestic opportunity of developing non-traditional copper resources, specifically those of lower grades, more complex mineralogy, and increasing level of deleterious elements especially arsenic and also process imported copper concentrate with increased arsenic content. Due to a significant and growing environmental issue associated with arsenic contamination, developing economically viable and environmentally reliable methods and approaches for arsenic management and control in unlocking the maximum value of such arsenic bearing copper resources has become more critical for sustainable growth of the industry.

Jan 1, 2019

By A. Iizuka

This study was aimed at separating and recovering arsenic, antimony, and tin from refining byproducts generated in a nonferrous smelter. The target byproduct was a silicate slag containing the above elements. The silicate structure of the slag was destroyed by alkali roasting, which then enabled leaching of arsenic, antimony and tin in solution effectively. Arsenic was selectively leached out by hot water after alkali roasting of the slag, while remained antimony and tin in the residue were leached out by a hot hydrochloric acid solution. However, when the solid-liquid ratio in the leaching solution was increased, hydrous gelatinous silica (gel-like silica) was formed during the hydrochloric acid treatment and filtration became difficult. Therefore, high temperature and high pressure leaching was attempted to improve the filterability of the silica leaching residue. In the high temperature and high pressure treatment, the hydrous gel-like silica was transformed into easily filterable aggregated silica, and thus, filtration became easy. By investigating the neutralization and sulfurization treatment for the obtained hydrochloric acid leachate (second leachate), it was possible to separate antimony and tin. INTRODUCTION This study aims to separate arsenic and recover valuable metals such as antimony and tin from refining byproducts generated in a nonferrous smelter. The target smelting byproduct is antimony-containing slag in this study. In many cases, alkali roasting was adopted as pretreatment for extraction and recovering valuable metals from smelting residues and wastes (Rao, 2006). For antimony-containing slag in this study, one effective method for separating arsenic is to convert arsenic to sodium arsenate (Na3AsO4) by roasting with a sodium compound and then to remove arsenic by hot water leaching. In addition, by alkali roasting, the silicate structure of the slag is destroyed, and effectively alkalizing tin and antimony is accomplished. Then, precipitating and recovering tin and antimony leached into the acid solution are considered possible by sulfurization and neutralization treatments. In this research, we conducted basic studies on the following. 1) Destruction of silicate structure in slag by alkali roasting and leachability of arsenic. 2) Conversion of hydrous gelatinous silica to aggregated silica by hydrochloric acid treatment of residues after arsenic leaching under high temperature and high pressure, as well as leachability of antimony and tin. 3) Separation and recovery of antimony and tin by sulfurization and neutralization treatments of hydrochloric acid leachate. These topics of investigation can support to consider the separation and recovery process of arsenic, antimony, and tin from antimony containing slag.

Jan 1, 2019

By V. Barranco, K. Yliniemi, A. Hubin, B. P. Wilson, D. Janas, M. Lundström, P. Hannula

Typical copper electrorefining electrolyte contains tens of grams of copper, some nickel and a wide variety of impurities. Some of these impurities comprise trace amounts of dissolved precious metals like Ag, Pd and Pt that currently are not fully exploited. This study presents an electrochemical method to recover precious metals via a combined electrodeposition-redox replacement (EDRR) process, where energy is used only to deposit copper during the ED-step, while the RR-step relies on the spontaneous replacement of copper by the precious metals (i.e. cementation). In order to conduct the recovery successfully, process parameters such as deposition potential (EDE), deposition time (tDE), redox replacement time (tRR) and the number of cycles (n) was adjusted. Investigations performed in synthetic electrolyte (40 g/L Cu, Ag = 10 ppm) proved the possibility to recover a high purity silver. Short deposition times (tDE) were found to result in small Cu particles that provided a suitably high surface area for silver cementation. In addition, experiments performed in real industrial electrolytes containing 44 g/L Cu and 54 ppb Ag (Boliden Harjavalta, Pori), demonstrated that EDRR could attain Ag recoveries of between 46–67% in an energy efficient, additive-free manner. INTRODUCTION The raw materials used in the state-of-art primary and secondary copper smelters contain mainly base metals, but also notable amounts of Ag, Pt and other precious metals. These valuable metals distribute mostly into the copper matte during the pyrometallurgical treatment (Tirronen et al., 2017) and end up into the cast anodes. During the electrorefining stage, metals more noble than copper such as gold and platinum, are not dissolved but enrich into the anode slimes. In contrast, silver may dissolve slightly and thus, minor chloride addition is used to retain silver in the slime as sparingly soluble AgCl. Moreover, other metals less noble than copper may also dissolve, such as nickel. Therefore the electrolytes used in electrorefining typically contain substantial amounts of copper, nickel and even arsenic as well as minor metals such as antimony and bismuth. In the Boliden electrorefinery these were reported to be in the magnitude of 51.2– 60.4 g/L of Cu, 10.9–15.8 g/L Ni, 7.2–15.9 As, 0.2–0.4 Sb and 0.1–0.4 Bi with an associated sulfuric acid concentration of 139–155 g/L (Lehtiniemi et al., 2018). Copper is the most noble of the dominating base metals in the electrolyte and is thus deposited on the cathode as 99.99% purity LME-A grade. Anode slimes are collected and subjected to precious metals treatment e.g. via pressure leaching—selenium recovery— Doré smelting—Ag refining—Au/Pt/Pd leaching and reduction process (Ludvigsson & Larsson, 2003). The Pori refinery, operating in Finland since 1941, not only produces 133 000 t Cu cathode and 2300 t nickel sulphate annually but also 66 t silver, 4.6 t gold and palladium and platinum (in year 2017).

Jan 1, 2019

By F. L. Pirkle, W. C. Bagby

The International Mineral Valuation (IMVAL) Committee, founded in 2012, develops strategies for conducting mineral valuations. These strategies have been developed into a mineral asset valuation template called the International Mineral Property Valuation Standards Template and is part of IMVAL's work to establish consistent international standards for mineral property valuation and reporting. This template is a living document that is updated from time to time and is based on three core values or pillars (Fundamental Principles): Competence, Materiality, and Transparency. Objectivity, Independence, and Reasonableness may be applied as required by national codes or standards. The Template harmonizes international standards for valuation of mineral assets as real property and represents a consensus of current good practices. It is not a stand-alone reporting code and does not supersede existing national reporting standards. The Template deals with mineral-property valuation as distinct from mineral-project evaluation. INTRODUCTION In his 1989 book, “The New Realities”, Peter F. Drucker, Clark Professor of Social Science at Claremont Graduate School, California, stated: “Information is data endowed with relevance and purpose. Converting data into information thus requires knowledge. And knowledge, by definition, is specialized.” (p. 209) The discipline of mineral valuation is knowledge-based and requires a mineral valuer to collect, analyze, and interpret data that are material to a mineral valuation assignment. In so doing, the valuer turns the data into information (data come first, followed by information as determined by the mineral valuer) for use by the client (from whom the valuer obtained the mineral valuation assignment) and other intended users. The valuer uses his or her experience and education (each valuer’s knowledge base) supported by a foundation built upon competency, materiality, and transparency, the three pillars of mineral valuation (“minerals valuation”, “mineral-property valuation”). The International Mineral Valuation Standards Template (“IMVAL Template”, IMVAL”, “Template”), third edition dated May 2018, defines mineral valuation as: “The estimation of the Value of a Mineral Property in money or monetary equivalent. The word ‘valuation’ can be used to refer to the estimated value (the Valuation conclusion) or to refer to the preparation of the estimated Value (the act of valuing).” (p. 14) Furthermore, the IMVAL Template clearly separates valuation from evaluation: “The Template deals with Valuation, which is distinct from Evaluation. The distinction inherent in these defined terms is that Valuation addresses the estimation of value of a Mineral Property, whereas Evaluation addresses the broader assessment of a Mineral Property for an investment decision.” (p. 2)

Jan 1, 2019

By F. Calizaya, S. Bhattacharyya

"This study presents the results of a set of sound level measurements carried out in three underground mines where auxiliary fans are used regularly. Sound levels as high as 100 dB(A) by the workings and 115 dB(A) by the fans were measured in these mines. These levels are quite high even for intermittent noise. Ventilation layouts and best practices used to overcome potential problems related to overexposure to fan noise in underground hard-rock mines are presented here. Factors such as fan type, rotational speed, and silencer-duct arrangements and their effects on noise attenuation are discussed. INTRODUCTION Auxiliary fans are used in hard-rock mines extensively. They are used to ventilate development headings, crusher rooms, conveyor transfer points, and underground shops. They are also used in developing long tunnels and ramps. Depending on the flow requirements and space restrictions, single or double stage axial fans with rigid or flexible ducts are used. Usually, they are high pressure (3.75 kPa) fans with flow rates ranging from 7.1 m3/s to 30 m3/s. The input power varies from 15 kW to 113 kW (Mutama, 2004; O’Connor, 2008). Typically, a 113 kW fan can produce sound levels as high a 113 dB(A) (Howes, 2005). Silencers are often omitted for cost reduction purposes or when used, they are not well maintained. Furthermore, silencers provide limited attenuation depending on their length and type. Typically, the sound attenuation ranges from 10 dB(A) for a 1 × D silencer to 14 dB(A) for a 2 × D silencer, where D is the fan diameter. In a mine, the silencer is subject to wear and tear, especially when the silencer is used in harsh environments. The interior lining can get clogged and the acoustic fill can get covered by dust and diesel soot, thus reducing the silencer performance. Periodic cleaning of the interior casing can increase the attenuation capacity of a silencer by 6 to 9 dB(A) (Dabson & Acuña, 2017). However, this requires a special silencer design in which the inner cartridge can be removed, the sound absorbing material replaced, the inner casing cleaned and the whole set retrofitted. A number of techniques can be used to reduce auxiliary fan noise in underground mines. These range from selecting the right fan for the job to installing adequate silencers. The choice depends on the air quantity requirement at the working place, cross-sectional area of the opening, duct diameter, fan type (axial or centrifugal) and local restrictions. Auxiliary ventilation layouts and best practices used to overcome potential problems related to overexposure to fan noise in underground hard rock mines are presented here. Factors such as fan type, rotational speed, duct diameter, and silencer-duct arrangements and their effects on noise attenuation are discussed."

Jan 1, 2019

By R. D. Williams

INTRODUCTION In a very real sense, Butte, Montana is where the copper came from that won two world wars. The price for that unrestricted mining and smelting of copper came due in 1983 when Butte was declared a Superfund site. One of the most complex remnants of this mining history is the Parrot Tailings site. The Parrot Smelter operated from 1881 to 1910 moving Upper Silver Bow Creek to the south and depositing approximately 370,000 cubic yards of waste material adjacent to Upper Silver Bow Creek. Additional mixed waste from other processing sites deposited below the Parrot Tailings at the Northside Tailings, Diggings East and Blacktail Creek Berm total approximately 224,000 cubic yards of contaminated material. As part of the Superfund remedy, EPA proposed leaving all these wastes in place. This proposal prompted a detailed criticism by several geoscientists familiar with the site. The EPA’s 2006 Record of Decision (ROD) did in fact leave these wastes in place which set the stage for a series of studies which continue to refine how best to protect extensive completed Superfund Remediation downstream of Upper Silver Bow Creek. HISTORY OF THE PARROTT MINE AND TAILINGS Placer gold was discovered in Silver Bow Creek, the headwaters of the Clark Fork River only a few miles from the Continental Divide in 1864. This discovery lead to a stampede of placer and lode claims. One of the first lode claims located was the “Parrot” lode, located in the fall of 1864. Progress for developing a mine at the Parrott was slow and it wasn’t until 1876 that one of the original locators encountered a rich vein of chalcocite. Butte had its first copper mine! The Parrott mine was patented in 1879 and by 1881 the Parrot, having lost a “t”, concentrator and smelter were built. The smelter and concentrator were an immediate success with a considerable market for processing the District’s rich copper ore. In 1885 the employees of the Parrot Smelter bought the Butte Reduction Works, another smelter. In 1885 dams were built on Silver Bow Creek to contain various tailings and related processing byproducts (Figure 1).

Jan 1, 2019

By C. Beul, R. Schipper, L. Santisteban

The Garfield Mine is a legacy vanadium mine in western Colorado with multiple mine openings, waste rock piles, and adit seepages that Cyprus Climax Metals Company (a Freeport-McMoRan Inc. subsidiary) voluntarily remediated through the Colorado Voluntary Cleanup Program (VCUP). The process to obtain a No Further Action determination from the VCUP is presented. Water treatment of the adit discharge, which is not part of the VCUP, is also presented. The remedial action objectives (RAOs) included preventing direct human or biotic exposure to the waste rock and radiation emitting from the waste rock; reducing metal concentrations and radionuclide activities in adit discharge and keeping the discharge from leaving the site; maintaining the existing undeveloped character of the surrounding landscape; and eliminating the need for a stormwater permit. The design included regrading, installation of an infiltration barrier and rock cover, a diversion channel, and a biologically based passive remediation water treatment system. An initial lab-scale evaluation provided design criteria for the full-scale system, which consists of a sulfate-reducing biochemical reactor and post-treatment aerobic polishing cells. The system has been operational since construction was completed in late 2016. Results indicate that the treatment objective of 90% bulk removal of constituents of potential concern (COPCs) is largely being met. The system also provides year-round operation and non-point discharge of effluent. After 20 plus years of site investigations, the remedial action was completed in 2016 with all RAOs met and a No Further Action determination granted in 2018. INTRODUCTION The Garfield Mine is a legacy vanadium mine located approximately 11 miles north-northeast of Rifle, in Garfield County, Colorado. It was leased and operated intermittently from 1929 to 1975 but was never operated by Cyprus Climax Metals Company (Cyprus) (a Freeport-McMoRan Inc. subsidiary). Prior to reclamation, site features included: • six mine openings—five adits and a vertical opening; • two adits discharging mining-influenced water (MIW)— Adit 1 at less than 4 liters per minute (L/min) and Adit 2 at 7 to 10 L/min; • main waste rock pile (13,500 m3) and small waste rock pile (minimal volume); • debris (small structure, mattress, bottles, shells, etc.) indicating site used for recreational purposes; and, • unimproved road, which has been used to access forest fires.

Jan 1, 2019

By Chandrakala Kari, Divyang Shah, Lakshmi Kanth Reddy, Kaushik Vakil, Sanjay Sarkar, Bhavin Desai

Gases rich in sulfur dioxide generated as a by-product of smelting and converting operations of copper concentrate contain selenium apart from zinc, lead, copper, cadmium, bismuth, antimony, and arsenic. The gaseous stream is heat exchanged in waste heat boiler, dust collected in electrostatic precipitator followed by cleaning in reverse jet scrubber to remove fumes, halogens and fine dust before passing to DCDA Monsanto catalytic converter to produce commercial grade sulfuric acid. The gas cleaning section of the acid plant uses water to scrub the smelting gases. During scrubbing, precipitated solids, known as sludge, settles at the bottom of the reverse jet scrubber containing 30 to 40 wt% copper and selenium up to 40 wt% selenium. In the existing process, the sludge collected during scrubber cleaning process is directly recycled to the smelter for copper recovery. However, the selenium is expected to again vaporize due to high oxidation potential during smelting and converting, causing accumulation of selenium in sludge. In present investigation, a roasting process has been developed to recover the selenium from acid plant sludge before charging it to smelter for copper recovery. Selenium is associated with copper in acid plant sludge as copper selenide, as determined by X-ray diffraction and electron microscopy. Thermodynamic and thermo-gravimetry study revealed that the copper selenide phase present in the sludge was amenable to oxidation at 600°C forming oxides of copper and selenium (Cu-Se-O). However, the dissociation of selenium from the copper oxide was made possible by sulfatation using sulfur dioxide between 450°C to 600°C, resulting into the formation of CuSO4(s) and SeO2(g). Lab scale experiments were carried out in vertical tubular furnace to determine the optimum roasting conditions with respect to roasting time, temperature and molar ratio of O2:SO2. Using these optimum conditions, selenium up to 90 wt. % in the form of SeO2 vapors could be recovered from the acid plant sludge in a large scale commercial roaster. Roasted sludge free from selenium and containing oxides and sulfates of copper could now be recycled in the smelter for copper recovery. INTRODUCTION There is an increasing demand for minor metals due to their use in innovative technologies. Similarly, the increase in demand for selenium is expected due to disruption in sectors related to renewable energy and photovoltaic battery. Selenium is a chalcogen and a non-metal with black or red crystalline appearance. Selenium is used in the glass industry to decolorize glass and to make red-colored glasses and enamels. It is also used in solar cells and as a photographic toner. Selenium is used as a brasses in stainless steel to improve its machinability. Selenium sulfide is used in anti-dandruff shampoos. In spite of the toxicity of its compounds, selenium is also an essential trace element for humans and animals. Selenium occasionally occurs free in nature, but more often occurs as selenides of iron, lead, silver and copper. There is an urgent need for the development of new methods for selenium recovery from the existing resources. Commercially, selenium is obtained only from anode mud produced during electrolytic refinery of copper.

Jan 1, 2019

By S. K. Mohanty, T. M. Dittrich, M. J. Allen, M. Dardona, A. C. Strzelecki, J. Hovey, Boukhalfa. H.

Rare earth elements (REEs) have many critical uses in advanced technology and the lack of a reliable domestic supply has been identified by the US DOE as a vulnerability to US economic security. Our research focuses on developing an economical process for extracting REEs from coal fly ash and power plant wastewater sludge by coupling a hydrothermal extraction process with the engineering of a ligand-associated media for concentrating REEs from the extraction liquid. Ash samples have been collected from two Detroit-area Power plants that burn 100% low sulfur coal (Trenton Channel) or fuel blends of 75% low sulfur coal and 25% high sulfur coal or 70% low sulfur coal, 15% high sulfur coal, and 15% petcoek (Monroe Power Plant). Ash samples are then subjected to hydrothermal conditions from 100-350 °C for 1-24 h at Los Alamos National Laboratory and leachate concentrations are measured by inductively coupled plasma-mass spectroscopy (ICP-MS) and compared to hydrofluoric acid digestion results. Experiments designed to aid in ligand selection and loading optimization show that di-(2-ethylhexyl)methanediphosphonic acid (H2DEH or DIPEX™) can be successfully loaded to a commercially available swellable organically modified silica (SOMS) media. Preliminary results show that DIPEX-associated SOMS surface loading can exceed 4x on a mass/mass basis and in simplified sorption experiments with Nd solutions (~pH 2), the DIPEX-SOMS media can sorb >90 mg Nd per g of media. Recovery of the sorbed Nd with low pH solutions has had low efficiency with the high (4 g DIPEX/ g dry media); however, near complete recovery was achieved with lower surface loadings (0.05 g DIPEX/ g dry media). Column experiments are being conducted to evaluate the optimal ligand/media ratio for repeated use. Completive sorption with Fe and other major cations are also being evaluated.

Jan 1, 2019

By J. Ji, A. Shi, B. Grcic

Floating copper minerals selectively against iron minerals such as pyrite has been challenging for most copper and copper/gold deposits, and many operations continuously strive to maximise copper recovery whilst meeting target concentrate grades. Maximising grade is achieved by selectively floating copper minerals and rejecting pyrite. This challenge is to minimise copper and gold losses in the roughers if this approach is used. A flotation testing program was conducted by Bureau Veritas Minerals on a high-level pyrite bearing copper/gold ore sample to establish the potential to reject pyrite in the rougher stage, but also maximise copper and gold recoveries, and ultimately final product grades. The test program included feed characterization, rougher and open-circuit cleaner flotation as well as locked cycle flotation tests. The main variables tested for achieving selective flotation of copper and gold were collector type and pH control, through lime addition. Grind sizes used in the rougher and cleaner tests had been established from previous test work. The collectors tested were SIPX, Aerophine® 3418A, Aerophine® 3421, Aerophine® 3422, Reagent S-11192 Promoter, Reagent S-11193 Promoter and FloMin C4150. The two best performing collectors were then tested in open-circuit cleaner tests, with the optimum conditions examined further in locked cycle tests. INTRODUCTION Bureau Veritas Minerals Metallurgical Division was retained by Eldorado Gold Corporation to perform a metallurgical testing program on a high-level pyrite bearing copper/gold ore sample from a gold- copper porphyry deposit to optimize flotation reagent scheme to improve gold and copper recoveries and copper content in the concentrate. The test program included feed characterization, rougher and open-circuit cleaner flotation as well as locked cycle flotation tests. Initially, various collectors were screened in kinetic rougher tests, and over a pH range of 8.5 (i.e the natural ore pH) up to 11.5, using a primary grind size of 80% passing 120 µm. The collectors tested were SIPX (sodium isopropyl xanthate), Aerophine® 3418A, Aerophine® 3421, Aerophine® 3422, Reagent S-11192 Promoter, Reagent S-11193 Promoter and FloMin C4150. The enrichment ratio, which is the ratio of copper to sulfur in the concentrate to that in the feed, was used as proxy for rating selectivity (i.e. degree of pyrite rejection or depression).

Jan 1, 2019