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By J. Etcheverry, E. Seguel, R. Díaz, F. Parada, R. Parra, E. Balladares

A thermodynamic and kinetic model is proposed to explain the formation of bornite, chalcopyrite and magnetite during partial roasting of enargite-containing copper concentrates. The proposed mechanism involves solid-solid-gas and solid-gas reactions between chalcocite, pyrrhotite, gaseous sulfur and oxygen. The formation of bornite could be explained by a two-step reaction: the first one involving the formation of chalcopyrite by a reaction between Cu2S, FeS and gaseous sulfur, followed by a second step in which the newly formed chalcopyrite reacts with additional chalcocite to form bornite. Although thermodynamically the direct formation of bornite is more favorable than that of chalcopyrite, the lower diffusion kinetics of bornite formation could explain the preferential formation of chalcopyrite. If the reaction occurs with stoichiometric chalcocite and pyrrhotite in the presence of gaseous sulfur to form chalcopyrite, no bornite is formed, while for the stoichiometric reaction between chalcopyrite and chalcocite in the absence of sulfur, only bornite is formed. For the formation of magnetite and chalcopyrite the mechanism could involve the partial reaction of chalcocite and pyrrhotite in the presence of oxygen. The proposed model is thermodynamically consistent with the experimental results obtained in laboratory and pilot plant. INTRODUCTION The processing of high-arsenic copper concentrates with over 0.7 wt.-% arsenic by conventional smelting-converting normally require an additional arsenic removal step to reduce arsenic content to levels below 0.2–0.3%. One of these high-arsenic copper concentrates is the one produced by the Ministro Hales Division (DMH) of CODELCO, which has the typical chemical composition shown in Table 1. Enargite (Cu3AsS4), chalcocite (Cu2S) and pyrite (FeS2) are the main mineralogical species present in these concentrates, with over 75 wt.-%, as shown in Table 2. Partial or “neutral” roasting of these concentrates under controlled temperature and partial oxygen pressure conditions at 923–993K (650–720ºC) can remove up to 95% of the arsenic as As2S3 (As4S6), which has a high vapor pressure under these conditions, leaving a calcine with low residual arsenic. In copper concentrates, arsenic is present generally as enargite and arsenopyrite. This compounds as well as others such as pyrite and covelite, decompose readily above 770 K (500ºC), releasing, in addition to gaseous arsenic trisulfide, gaseous sulfur which by oxidizing to SO2 generates most of the heat required by the process.

Jan 1, 2019

By R. T. Seaman

There are several hydrometallurgical alternative processing options under development for high enargite and arsenic containing copper concentrates ranging from biological, atmospheric, medium temperature oxidation to total oxidative leaching. The most developed of these options is the CESL process boasting pilot campaigns totaling more than 8760 hours of fully integrated continuous operation, as well as, over two thousand batch scale oxidation tests performed on dozens of concentrates for flowsheet development. This paper reports the observed deportment of arsenic under varying arsenic, iron and acid oxidation conditions; it also describes the impact of iron to arsenic ratios and residue recycle on overall arsenic behavior and its stability in residue. INTRODUCTION The CESL Hydrometallurgical Process (CESL) was originally developed to compete with smelting of clean chalcopyrite concentrates. High Treatment Charges and Refining Charges (TC/RCs) combined with premiums on London Metal Exchange (LME) grade copper from onsite refining made treatment of copper concentrates using CESL a viable option at the time. As world demand for copper increased and a subsidized Chinese refining market developed, smelting contracts became more favourable for the miner and focus was placed on modifying the CESL Hydrometallurgical Process to handle concentrates with elevated deleterious elements. Dozens of pilot campaigns processing concentrates containing elevated concentrations of uranium, fluoride, mercury and nickel all led to further developments of the CESL Hydrometallurgical Process. In 2016 a United States patent was granted for a novel modification of the CESL Hydrometallurgical Process for recovery of copper from arsenic-bearing and/or antimony-bearing copper sulphide concentrates, Henry Salomon de Friedberg (2016). To date, over 2000 bench scale tests along with five fully integrated closed loop pilot campaigns, totaling over a year of online operation time, have been conducted processing arsenic bearing copper concentrates. This paper summarizes the observed behavior of arsenic under varying autoclave and process conditions.

Jan 1, 2019

By M. Massinaei, S. Abasi, M. Oliazadeh

Reliable measurement and control of the pulp-froth interface (froth depth) in the flotation columns is critical to achieve the desired metallurgical performance. Shallow froths result in a reduction of grade because of particle entrainment and deep froths reduce recovery because of the drop back of entrained and weak hydrophobic particles to the pulp zone. The cleaner circuit at Miduk copper concentrator consists of three parallel MicrocelTM flotation columns 4 m in diameter and 12 m in height. The interface level of the flotation columns is measured by two pressure sensors located at 1.4 and 2.4 m from the cell lip (in the collection zone). In this study the performance of the level detection system at different operating conditions was evaluated. A calibrated float-ultrasonic location system was used for measuring the actual froth depth. The results indicated that the existing level control system was subject to considerable error. The absolute error was found to be 33.27 cm. This was mainly attributed to the constant set-point value assigned to the froth density (ρf=0.3 g/cm3), which actually varies with the operating conditions. The actual froth density at different conditions was calculated and its average value (ρf=0.12 g/cm3) was re-set in the control system. This resulted in substantial reduction in the level-detecting error. Recommendations for more accurate measurement of the interface level were presented. INTRODUCTION Froth flotation is a complicated process because of many involved variables. One of the most significant controlled variables in the flotation process is the pulp-froth interface (froth depth). The froth depth has a profound influence on the metallurgical performance parameters (i.e. grade and recovery). Shallow froths result in grade reduction because of particle entrainment and deep froths reduce recovery resulted from the drop back of entrained and weak hydrophobic particles to the pulp zone (Finch and Dobby, 1990). Various techniques have been developed for measuring the interface level of the flotation cells in the last few decades. These techniques can be classified as: Techniques based on hydrostatic pressure measurements (Finch & Dobby, 1990), Techniques based on the density gradient (float) (Moys & Finch, 1988), Techniques based on conductivity measurements (Uribe-Salas, 1991), Techniques based on temperature measurements (Moys & Finch, 1989).

Jan 1, 2019

By K. Scalise, M. Teixeira, K. C. Kocsis, J. Parrish

"This paper aims to introduce and discuss a newly developed algorithm, which can be uploaded on smart watches and other mobile devices to inform and protect underground personnel working in hot and humid environments. Using this algorithm, the metabolic rate (M) of a worker can be determined in “real-time”. In addition, besides warning signs and alarm signals emitted by the smart watch, the algorithm has the ability to determine work-rest time periods based on the worker’s physiological parameters (e.g. age, height, weight, gender) and the climatic parameters (Td, Tw, flow velocity). This paper also investigates whether a re-designed auxiliary ventilation system can provide appropriate work conditions in the production stopes. INTRODUCTION An underground environment with high temperature and humidity conditions generated from various heat and moisture sources can significantly affect the thermoregulation processes of the human body. This can lead to heat-related illnesses such as heat rash, heat cramps, heat exhaustion, and heat stroke. Heat stroke is a serious illness that carries a high risk of fatality, if the underground workers are not immediately treated and the climatic conditions are not corrected (McPherson, 1993). There are two methods used to control the climatic conditions in underground mines: (1) Using ventilation and cooling systems to reduce the thermal stress on the mine workers, and (2) Reducing the heat transferred to the ventilating air from major sources such as: strata, auto-compression, mining equipment, groundwater, blasting, etc. The heat from strata and auto-compression is transferred to the ventilating air at varying degrees, based on factors such as the virgin rock temperature (VRT), the thermal properties of the surrounding rock formations, and the surface air temperature. However, the use of mining equipment and blasting processes can be managed to reduce their contribution to the mines’ overall heat load. To provide appropriate work conditions, selecting the most suitable cooling method depends on the level of heat to be removed, the location of the problem areas, and economic constraints (Brake & Bates, 2002; McPherson, 1993). Because every mine is unique, it is essential to monitor the climatic conditions to select the most appropriate method of cooling. A majority of the time, simply adjusting or re-designing the auxiliary ventilation system can efficiently remove the heat and dilute the contaminants. This may involve moving the ducting system, changing the size of the duct, changing the size of the auxiliary fan, or switching from forcing to an exhausting fan arrangement. If the required air volumes are higher than the existing primary ventilation system can supply, other measures may need to be taken to increase the airflow delivery. This may include regulating the flow in some areas of the mine to increase airflow to the desired area, adding booster fans, or replacing the primary fans."

Jan 1, 2019

By T. Imgrund, A. Bischoff, S. Spürk

"Specific conditions of underground coal mines at great depth, such as high temperatures, high rates of methane inflow and natural ventilation pressure provide considerable challenges for the reliable design of appropriate mine ventilation systems. For decades coal mining in Germany has coped with these extreme circumstances. Based on these experiences appropriate software tools for simulating mine ventilation networks have been developed and continuously enhanced. This paper examines differences in the simulation results and underlying calculation methods of two software solutions. The analysis is based on the comparison of FlowNetTM and VentSim DesignTM (Version 5.1) regarding the accurate prediction of fluid mechanical quantities and their validation with real measurements. A special focus lies on exceptional ventilation scenarios typical for deep coal mines, like gas inflows, fire events and buoyancy effects. The purpose of the comparison is the future utilization of the experiences gained during mining at up to 1,500 m depth for state of the art ventilation design. INTRODUCTION Ventilation is an essential component of mine operations—as providing fresh air to workplaces, diluting and removing explosive, toxic or asphyxiating pollutants and assuring sufficient climate conditions. On the other hand, the costs of ventilation contribute to a large extend to the overall operating costs of a mine, as the share in the energy consumption of a mine may be in the range of 25–50% (Belle, 2008). Accordingly, proper design of ventilation systems has been a task since early times. Consequently software for modelling of ventilation networks has been developed in order to improve efficiency and accuracy of calculations (Maleki, 2018). Compressible flow has not been considered by initial ventilation modelling software tools. In Germany, the consideration of changes in airflow density—depending on pressure and temperature—in the algorithms of ventilation simulation software has been started in 1974 by Duepre & Hofbauer. This development had been continued by the Pruefstelle fuer Grubenbewetterung (Test Centre for Mine Ventilation) and resulted in the release of a computer ventilation software named “Wetter”. The mine inspectorate of the German state of North Rhine-Westphalia, which supervises the major coal mining district in Germany, declared ventilation simulation based on compressible flow being “state of the art” in 1985 (Pollak, 1985). This pronouncement made allowance for assessing the consequences of underground fires and downtimes of main fans based on advanced software solutions."

Jan 1, 2019

By C. Chongo, L. Little, Q. Mclennan

Since commissioning in 2005, First Quantum Minerals’ Kansanshi Mine in Zambia has produced over three million tons of copper and 1.8 million ounces of gold from a deposit with highly varied copper mineralization. The processing plant is complex with three independent flotation circuits, gold gravity concentration, atmospheric and high pressure leaching, solvent extraction, and electro-winning. For the first ten years of operating the plant, copper mineralogy was inferred from total copper and acid soluble copper assays, and various misconceptions developed about the mineralogy which impacted operational strategy and project decisions. In 2015, an on-site mineralogy section was set up with an Automated Scanning Electron Microscope, the Zeiss MinScan. This instrument has provided detailed mineralogical data that have debunked some of the mineralogy myths that had developed, highlighting the key underlying issues affecting processing performance. This has enabled the development of projects targeting losses of specific copper minerals in defined size classes, unlocking significant value. This paper describes some of the early misconceptions around mineralogy at Kansanshi processing plant, outlining projects that have benefitted from understanding copper mineral distribution. Learnings from Kansanshi are to be applied at other FQM sites, with a similar mineralogy program being set up at the Cobre Panama project. INTRODUCTION The deposit Kansanshi Mine is located in the North Western Province of Zambia, and it has been the largest copper producer in Africa for over a decade. The deposit occurs within a broad northwest trending antiform that can be traced for 12 km, hosted in a zone of complicated faulting, abundant vein injection, breccia development and down-dropped rock units that lies within Kansanshi’s mining license. The copper mineralization within the deposit is predominantly vein-hosted, although some is strata-bound, occurring in thin bands and veinlets parallel to foliation, and some is disseminated associated with albite-carbonate alteration (Gray, 2015). Supergene enrichment has led to significant variety in the copper mineral speciation. This alteration is complex, and has not followed the text-book principles with a neat layer of supergene mineralization overlying fresh hypogene ore. Rather, the occurrence and extent of oxidation and supergene enrichment is highly irregular throughout the deposit, sometimes occurring at significant depth in localized pockets and along fault lines. Gold mineralization tends to follow the copper within the deposit, but it is dominated by coarse-grained, free-milling native gold. The extent of copper oxidation – measured as the fraction of total copper that assays as acid soluble – is used to classify ore into three main ore types, named Sulphide, Mixed, and Oxide, which are then fed to three different primary crushers for processing in independent circuits.

Jan 1, 2019

By K. Lambrych

Hydrometallurgical equipment used to extract and refine copper and other metals is exposed to extremely corrosive chemical environments. These systems must be fabricated from materials that are durable against corrosion, meet exacting mechanical requirements, and perform economically over the life of the mineral processing plant. In several recent large hydromet projects, fiber reinforced plastic (FRP) has been shown to provide equal if not superior durability relative to alternative materials of construction. This paper will compare corrosion performance and cost of plant equipment fabricated with corrosion resistant alloys, rubber lined steel, and FRP made with Epoxy Vinyl Ester thermoset resins. Case histories for FRP piping, storage tanks, extraction vessels, and electrowinning cells combined with laboratory-based corrosion studies will be reviewed to demonstrate how FRP materials are selected for hydrometallurgical equipment. Specific examples from projects such as El Boleo, Long Harbour and others will be used to demonstrate how the versatility of FRP allows for remote fabrication of durable corrosion-resistant large-scale equipment for hydrometallurgical refinement of copper and other metals. INTRODUCTION Fiber reinforced plastic (FRP) began to emerge as a material of construction for corrosion service in many industrial processes in the late 1950s because of its ease of use and overall performance. Corrosion resistant (CR) FRP, as an alternative construction material to alloys, rubber and tile linings, grew steadily in following decades. The aggressive operating conditions associated with hydrometallurgical processes typically require materials of construction that can withstand acids and acid chlorides at temperatures up to 90°C, sometimes higher. Common metallic materials of construction, such as carbon steel, are not compatible with these conditions. High nickel alloys are capable of meeting performance requirements, but often at a high cost. In many cases FRP mining equipment based on epoxy vinyl ester resin (EVER) offers durability, long service life and lower maintenance costs. Through demonstrated performance and decades of case histories, more and more design engineers are confidently specifying FRP as the material of construction for new mineral processing equipment.

Jan 1, 2019

By K. Parameswaran, A. Veek, A. Fernandez, P. Safe

ASARCO LLC (Asarco) Hayden Copper smelter, in Hayden, Arizona, has operated continuously (except for minor labor and economic curtailments) since 1912. The current plant configuration comprises INCO flash smelting, Peirce Smith (P-S) converting and fire refining/anode casting. In order to comply with the U.S. Environmental Protection Agency (EPA) 1-hour National Ambient Air Quality Standards (NAAQS) for sulfur dioxide promulgated in June 2010, the Hayden smelter had to increase overall sulfur fixation from approximately 95 % to 99+ %, requiring a retrofit of the existing converter department and upgrade of associated emissions controls systems. The Converter Retrofit Project (CRP) involved replacing the five existing 13’ diameter Peirce-Smith converters (2 blowing, 2 hot standby) with three new 15’ diameter Peirce- Smith converters (1 blowing, 1 hot standby) to allow for more effective process gas capture using the existing acid plant contact section. Significant upgrades were made to the acid plant wet gas cleaning section and converter primary gas handling system to improve capture performance, reliability, and sludge handling. In addition, a new tertiary gas capture system was installed to further minimize fugitive emissions from the converter aisle. The existing furnace tapping ventilation ESP was replaced with a new baghouse to enable at least 50% SO2 removal using high-surface-area dry lime injection, which was also implemented on the existing converter secondary gas baghouse system. The first two converters and all gas handling system improvements were started up in April 2018, and the third converter was brought online in November 2018. This paper presents the regulatory context; genesis and rationale for CRP; detailed description of the project; challenges during commissioning; system improvements implemented, and design considerations made, and systems performance and process optimization achieved since startup.

Jan 1, 2019

By G. Stokreef, B. R. Davis, M. Kadiyski, T. LeBel, Alvear Montenegro, R. Parra, V. Montenegro

A drop tower and methodology were developed to study the impact of mineralogy on the flash combustion behaviour of copper concentrates. Reaction shaft conditions of a flash furnace were simulated using a drop tower reactor and different molar ratios of oxygen gas to sulphur in the feed. Five feeds were tested in the drop tower including pure minerals chalcopyrite and pyrite as well as two mineralogically distinct copper concentrates and a blend of the two. The behaviour of each of these five feeds were evaluated through analysis of the solid combustion products for elemental sulphur, mineralogy and particle size. Desulphurization curves for each of the feeds were produced. It is thought that the desulphurization behaviour of pure minerals can be used to estimate the behaviour of a concentrate and that a blending recipe could be developed, which can optimize the chemical and mineralogical inputs to the flash furnace and guide operations on the optimum furnace conditions. Overall, the drop tower has proven to be a useful tool for studying flash combustion processes. Experimental results suggest that mineralogy has a significant impact on the conditions required for the production of matte with consistent grade. This finding is especially significant for custom smelters, which may see several different blends in a day. INTRODUCTION Flash furnace smelting is a continuous feed, batch tap, pyrometallurgical process that is used to produce copper matte from copper sulphide concentrates. Dried concentrate(s), silica flux, and secondary materials are blended to a target chemical composition, typically given by the Cu/S ratio, before processing. This fine, dried feed is dispersed throughout the heated furnace reaction shaft with an O2-enriched reaction gas. The particles heat up in the shaft of the furnace, ignite and undergo reactions that pre-oxidize the iron sulphide and some of the impurities in the feed. Oxidation reactions in the shaft of the furnace desulphurize the feed and produce an SO2-containing off gas. The reacted particles fall into the furnace settler and separate into a Cu-rich matte and an iron-silicate slag, while the SO2-containing off gas is carried out of the furnace for heat and dust recovery, as well as H2SO4 production. Some concerns during operation of the furnace are dusting, over or under oxidation of the feed, and a stable flame.

Jan 1, 2019

By A. Miura, R. Winarko, D. B. Dreisinger

Chalcopyrite leaching in acidic ferric sulfate media at ambient temperature exhibits slow kinetics and poor leaching efficiency, which is typically attributed to surface passivation. Recently, the presence of small amounts of iodide in the leaching media has been reported to significantly accelerate chalcopyrite dissolution. To further explore the behavior of chalcopyrite leaching in the presence of iodide, it is essential to determine the factors that control the leaching kinetics. In this study, we investigated the effect of initial solution potential, iodide concentration, particle size, and temperature on the kinetics of copper extraction from chalcopyrite. The key in the experimental setup was to minimize iodide loss by iodine evaporation and precipitation as cuprous iodide. The former was achieved by performing the leaching tests in fully sealed leaching systems. The latter was achieved by maintaining the solution potential higher than 420 mV (versus Ag/AgCl electrode), where formation of CuI precipitate is possible. The experimental results show that higher solution potential and higher iodide concentration had a pronounced positive effect on the rate of chalcopyrite dissolution. There was no effect associated with particle size, which might be due to the narrow size range tested. The positive effect of increasing temperature only manifested itself when the solution potential was sufficiently high. INTRODUCTION The main challenge of chalcopyrite leaching using ferric ions as the oxidant at ambient temperature is the slow kinetics and poor extraction. The reason for this leaching behavior has been attributed to passivation resulting from formation of insoluble passivating layers on chalcopyrite surfaces, which could be metal-deficient sulfides, elemental sulfur, polysulfides, and iron precipitates (Klauber, 2008). To overcome passivation, technologies using higher temperature and pressure have been developed to leach chalcopyrite concentrates, such as the CESL process (Mayhew, McCoy, Mean, & Woeller, 2013), FLSmidth Rapid Oxidative Leach (ROL) process (Mulligan et al., 2017), Galvanox (Dixon, Mayne, & Baxter, 2008), Albion process (Hourn, Turner, & Hourn, 2012), HydroCopper (Hyvärinen & Hämäläinen, 2005), TPOX (Dreisinger, 2006), and AAC/UBC process (Dreisinger, Dempsey, Steyl, Sole, & Gnoinski, 2003). Even though they are effective in accelerating copper leaching kinetics, those technologies typically have higher capital and operating costs associated with production of concentrates and application of autoclaves. This makes them unsuitable for treating low grade primary copper sulfide ores.

Jan 1, 2019

By A. Belleau, F. Lavoie

"Quebec Iron Ore Limited (QIO), a subsidiary of Champion Iron Limited (Champion), secured assets in early 2016 from the previous Bloom Lake operations in the heart of the Labrador Trough region of Northern Quebec, Canada. A plan was developed to bring the Bloom Lake facility back in operation once the iron ore market conditions improved. The project restart involved several aggressive moves to change operational effectiveness, one of which was the utilization of the existing Phase 1 Plant with an upgraded process flowsheet and equipment. The production target was 7.4 million tonnes per annum of 66% iron ore concentrate with an approximate 10% increase in iron ore recovery. The Phase 1 upgraded flowsheet was initially based on review of historical Phase 1 data and pilot testing data undertaken during the original Phase 1 operation and was finalized once additional metallurgical test work and process modeling was carried out by Mineral Technologies in late 2016. The final upgraded flowsheet is comprised of rougher spirals, rougher middling scavenging spirals, and an up-current classifier (UCC) with overflow scavenging spiral stage. A portion of the gravity circuit tailings treated by a series of wet high intensity magnetic separators (WHIMS) to boost overall iron recovery. This paper will describe the efforts taken to test and design the Bloom Lake Phase 1 upgraded flowsheet and report on the plant performance since commissioning in first quarter of 2018.SITE HISTORY The Bloom Lake iron mine property is located approximately 13 km north of Fermont, Quebec, and 10 km north of the Mont Wright iron ore mining complex, property of ArcelorMittal Mines Canada. The Bloom Lake mine facilities include a 32 km railway to efficiently transport the high-quality iron concentrate to a loading port located in Sept-Iles, Quebec. The Bloom Lake property was first explored in 1951 and went through multiple minor exploration campaigns by different companies up until 1998. A major exploration program was conducted by Quebec Cartier Mining (QCM) in 1998. This program included line cutting, surveying, diamond drilling, geological mapping, mini-bulk sampling, bench-scale preliminary metallurgical test work and preparation of a “mineral resource” estimate."

Jan 1, 2019

By Wei Ye, Lei Li, Xianjian Guo, Feng He

The purpose of the paper is to study the smelting process of copper concentrate with high Cu/S ratio produced in Africa or other areas around the world and theoretically analyze the influence of slag type on the smelting process. The paper has compared the processes used successfully in smelters, including CAPEX and OPEX and other factors, to provide reference for the industry to select the best process under specific conditions. INTRODUCTION At present, a great amount of copper concentrates with high Cu/S ratio, with chalcocite being the major source of copper, have been produced or are about to be produced in the Copperbelt in Africa and other countries in the world. The copper concentrate is usually high in copper (~30–55wt.%) and low in sulfur (~10–15wt.%). The conventional processes of smelting-converting and roasting-leaching-electro-winning for treating this type of concentrate is high energy and materials consumption such as electric power\fuels\flux et al. The direct-to-blister (one-step) smelting process can be considered favorable due to a reduction in processing steps. Up to now, there are four copper smelters in the world operating with the direct-to-blister smelting process. There certainly are problems in the process such as high copper content in the slag, strict operation requirement of slag cleaning electric furnace, etc. Except technical issues, the impact of factors on economical aspect such as power and material cost needs to be considered comprehensively in process selection. In this paper the technical issues of direct-to-blister smelting process, including slag treatment process, were analyzed and the comparison to the conventional processes and economic analysis of direct- to-blister smelting process to treat high Cu/S ratio concentrate were done.

Jan 1, 2019

By Kenta Kusunose, Dadang Sundana, Ahmad Munib Fikry, Yosuke Unno

PT Smelting has operated the Mitsubishi Process in which feed materials are charged into the S- furnace through top lances using a pneumatic conveying system comprising 6 feeding tanks (S-Feeding Tanks). Each S-Feeding Tank was designed as a batch machine and had been operated in a batch mode so that feed materials are conveyed 70% of the operating time (conveying time). An S-Feeding Tank was revamped in August 2016 in order to attempt to achieve 100% conveying time during operation, aiming to reduce the furnace draft fluctuations and melt temperature fluctuations leading to reduced copper losses to slag and refractory wear on the sidewall in the melt zone. After two-year operation with some adjustments and modifications, it was concluded that the new system works as expected. During scheduled major shutdown in November 2018, the same modification was applied to all other 5 feeding tanks as the first one, while several modifications other than the S-Feeding system were also conducted for the purpose of expanding the capacity of the smelter. INTRODUCTION PT. Smelting (PTS) Gresik Copper Smelter and Refinery is located 30 kilometers north of the city of Surabaya, East Java major port. PT Smelting equity partners are Mitsubishi Materials with 60.5%, PT. Freeport Indonesia with 25%, Mitsubishi Corporation RtM Japan Ltd. with 9.5% and JX Nippon Mining and Metals corporation with 5.0%. PTS was established in February 1996 and started commercial operation of the Mitsubishi Process in 1999. The Mitsubishi Process is a unique technology which connects 3 stationary furnaces, namely the Smelting furnace (S-furnace), Slag-Cleaning furnace (CL-furnace) and Converting furnace (C-furnace) by covered launders. These features result in savings in manpower and energy, and make the process more environmentally friendly since there is no ladle transfer of melt. Figure 1 illustrates the outline of the Mitsubishi Process.

Jan 1, 2019

By Hajime Miki, Gde Pandhe Wisnu Suyantara, Shigeto Kuroiwa, Eri Takida, Tsuyoshi Hirajima, Masashi Yamane, Yuji Imaizumi, Mitsuru Sawada, Keiko Sasaki

Selective flotation of molybdenite and chalcopyrite using various oxidation treatments (i.e., ozone treatment, hydrogen peroxide, and Fenton-like reagent) was studied in this work. It was found that ozone treatment produced a depressing effect on both minerals, resulting in a lower separation of both minerals. On the other hand, oxidation treatment using hydrogen peroxide (H2O2) and Fenton-like reagent could deliver selective oxidation on chalcopyrite surface and depress its floatability. However, bench-scale copper- molybdenum (Cu-Mo) flotation tests show that oxidation treatment using H2O2 required a higher dosage of H2O2 and a longer treatment time in order to deliver comparable flotation results to that of conventional Cu- Mo flotation using NaHS treatment. On the contrary, Fenton-like reagent could deliver selective Cu-Mo flotation results comparable to that of NaHS treatment at a lower dosage of H2O2 and a shorter conditioning time. Contact angle measurements and X-ray photoelectron spectroscopy (XPS) analysis were employed to understand the depressing mechanism of these oxidation treatments. It was found that the chalcopyrite surface covered by various oxidation products that render the surface hydrophilic. Meanwhile, the surface of molybdenite was slightly oxidized, making the surface remained hydrophobic. INTRODUCTION Molybdenite (MoS2), the main source of molybdenum, is associated with copper sulfide minerals, mainly chalcopyrite (CuFeS2), as a trace mineral. Molybdenite is a valuable mineral and plays an important role in making the copper-molybdenum (Cu-Mo) processing plant economically viable (Laskowski, Castro, & Ramos, 2013). Therefore, it is important to recover molybdenite efficiently and effectively and separate it from copper minerals. Conventional separation of copper and molybdenum minerals is usually carried out in selective Cu- Mo flotation process using a copper depressant (i.e., sodium hydrosulfide (NaHS), Noke’s reagent, or cyanides) (Ansari & Pawlik, 2007; Bulatovic, 2007). However, this process still suffers due to imperfect molybdenum recovery. Moreover, the copper depressants use in this process are toxic and harmful to the environment. As alternative, various methods have been developed to replace these toxic and hazardous depressants, for example, by synthesizing new non-toxic or low-toxic depressants or by using various oxidation methods such as sodium hypochlorite in an alkaline medium, manganese dioxide, oxygen, and plasma pre-treatment (United States Patent Office Patent No. US 2559104 A, 1951; Patent No. US2238250 A, 1941; Barzyk, Malysa, & Pomianowski, 1981; Hirajima et al., 2014). However, the oxidation phenomenon and mechanisms are still not clear since detailed surface analyses have not done and the plasma pre-treatment method is facing technical issues for industrial scale application.

Jan 1, 2019

By Z. X. Cui, X. D. Ma, L. Contreras, X. Jiang

The bath smelting process is one of the key techniques applied in the current copper smelting production, which includes bottom-blown and side-blown furnaces differently from the direction of gas injection. Liquid matte and slag form the bath in the smelting furnace. To better understand the fluid dynamic of the bath and further optimise the operational efficiencies in the smelting furnace, physical simulation of the bath in the horizontal cylindrical reactor was investigated. The mixing behaviour and wave motion of the bath in the bottom-blown and side-blown furnaces have been studied systematically. The mixing time is reduced with increasing gas flowrate and height of the lower layer, while it increases significantly with increasing the height of the upper layer. The wave motion in the furnaces can be suppressed with the presence of the upper layer, but the bath with a lower density generates higher wave amplitude. Overall, in comparison to bottom-blown mixing, side-blown mixing shows better mixing efficiency and lower wave amplitude. In addition, the plume eye and splash phenomena occurred in the bottom-blown furnace have also been studied. INTRODUCTION The bath smelting process is one of the principal techniques used in the current copper making industry. The bath smelting techniques have some advantages of high smelting efficiencies, low capital costs, low energy consumptions and reduced dust generations. From the direction of gas injection, the bath smelting techniques can be classified into five general types: (1) top-submerged blown, e.g. Ausmelt and Isasmelt; (2) top-suspended blown, such as Mitsubishi smelting; (3) top direct-to-copper smelting, such as Outokumpu smelting; (4) submerged side-blown, such as Noranda and Teniente smelting; (5) bottom-blown, such as Dongying Fangyuan bottom-blown copper smelting furnace (Schlesinger et al., 2002; Zhao, Cui, & Wang, 2013; Perez-Tello et al., 2018). In many metallurgical processes involving bath smelting situations, the objectives of agitation are to accelerate the bath homogenization inside the bath which makes a significant contribution to process efficiency and product quality. Extensive mixing studies have been performed for ladle refining and tundish associated to steelmaking processes through water physical models. Mixing time is the parameter used to quantify the efficiency of mixing in the liquid bath and the first empirical equation predicting mixing time was proposed by Nakanishi, Fujii and Szekely (1975). Later on, more systematic studies have been carried out on ladle mixing and researchers realized that the geometry of the vessel could also have a significant impact on mixing (Khajaviand & Barati, 2010). However, the investigations using ladles which are vertical cylindrical vessels, cannot be directly used in the study of mixing behaviour in horizontal cylindrical reactors. Understanding the fluid dynamics inside horizontal cylindrical bath smelter is useful from the industrial perspective. Recently fluid dynamics of horizontal cylindrical smelter related to copper smelting have been studied by the authors’ group (Shui et al., 2015a, 2015b, 2018a; 2018b; Xu et al., 2018, 2019).

Jan 1, 2019

By M. Silva, G. Olivares

In a leach pile, the drainage layer or commonly known as the cover layer has been a matter of discussion due to the multiple phenomena or events that occur within it which impacts negatively the leaching process. All these events are directly related to the loss of draining capacity and irregular flow net due to an excessive compaction and entrapment of fines reducing the flow velocity which in turn cause more fines to get trapped. Over time the problem will continue to worsen rapidly and accelerating continuously like a geometric progression. When the cover layer gets to a certain level of deterioration, the only alternative to continue operating safely is to replace the entire cover material, a huge cost for any mining operation. TESRA has developed a solution to counteract the accelerated advance of the deterioration of the drainage layer. The solution consists of using large agricultural tractors coupled to a proprietary design of a heavy- duty subsoiler-ripper for deep tillage for loosening and breaking up the ground at various depth and with different legs spacing from 0.4 m or above. Also, if the ground still requires additional breaking, TESRA proposes the use of a rototiller that breaks any lumps to homogenize the ground to a certain particle size. With the subsoiler-ripper/rototiller combination equipment a layer of 0.5 m deep is properly treated to match closely the original conditions of the ground thus regaining the flow net as close as to the original seepage condition. Once the ripping is completed the subsoiler-ripper tool is replaced with another that is similar to a plow and thus continue with the installation of drainage pipes. The tool digs a Y-type profile trench allowing the pipe to sit precisely at the bottom of the trench, contributing additional lateral and lower support. This tool controls the installation depth of the drainage pipe at different inclination grades and also eliminating any irregularities that may block the flow of the leaching solution. Sometimes the obstructions block completely the flow of the solution. The main result of this methodology is the reduction of groundwater levels and at least 2% less humidity in the leaching gravel due to an increase in the draining capacity of cover. The methodology developed by TESRA applied during a couple of cycles of a dynamic leach pile has proven to be a real alternative to remedy and recuperate the cover close to its original conditions. The methodology requires only one large tractor and two accessories operated in different stages while replacing the more traditional and less efficient machinery such as the Bulldozer, Motor-grader, and backhoes.

Jan 1, 2019

By E. Landry, W. Naudts

"The amenability theory was developed by Alex Naudts and considers grouting operations as continuous permeability tests where Newtonian test fluid water, is replaced by a cementitious or solution grout. It utilizes hydraulic conductivity and grouting data to establish the amenability coefficient of a particular grout zone. The amenability coefficient reflects the quantity of apertures or voids accessible to the selected grout, divided by the number of apertures accessible to water, in the zone intersected by the grout hole. By calculating, in real time, the effectiveness of a particular grout at the start of a grouting operation, the most appropriate grout to match design criteria can be targeted. By incorporating the amenability theory in every grouting operation, schedule savings and an increase in productivity are typically achieved. Three short case studies illustrate different enhancements in productivity: At an underground mine during the construction of declines, a cement based suspension grouting program was highly ineffective and did not appear to affect the inflow into the mine workings. The amenability theory was utilized to evaluate the grouting program and based on this evaluation, changes to the grouting program were made which allowed the grouting to meet the project’s objectives. On another project, a pre-excavation grouting program implemented during shaft sinking operations utilized the amenability theory to select the most suitable grout for the program. Schedule constraints also influenced the selection of the grout. In the third case history presented, the amenability theory was used to select the grouting sequence of every individual grout hole for a grouting project in a highly heterogeneous soil formation. INTRODUCTION The amenability of a given soil, civil structure or rock formation for a particular grout formulation is a measure of the ability of that grout to permeate the pores, fissures and apertures water will flow through. The amenability coefficient of a particular grout for a specific zone equals the percentage of apertures, pores and fissures accessible by water that are also accessible by grout. Grouting should be viewed as a continuous permeability test, where the apparent in-situ hydraulic conductivity value for a particular grout zone is continuously evaluated, using grout instead of water as test fluid."

Jan 1, 2019

By Phillip J. Mackey, Susumu Okabe

Copper mining and smelting has been carried out in Japan for over 13 centuries and in fact stretching back into antiquity. In 1889, the first electrolytic copper refinery in Japan based on Elkington’s principles was started by Furukawa Mining at Honjyo in downtown Tokyo based on the multiple or parallel system. Within two decades, a number of other copper refineries had begun operation. Interestingly, but perhaps not surprising given Japan’s long history of copper production, this first refinery in Japan commenced just ten years after the first refinery in the USA (1879), and not long after the first plants in Europe. By 1935, as a result of growth in copper demand, there were six domestic copper electro-refineries in Japan with an installed capacity of 97,800 tonnes/a, while actual production was 71,200 tonnes/a. Since about 1932, the majority of these plants were based on the more common parallel or multiple system. In 1946, just after World War II, domestic production capacity had dropped considerably to 29,000 tonnes/a, however with reconstruction, capacity and production expanded rapidly, with production reaching 81,000 tonnes/a in 1955, 187,000 tonnes/a in 1960 and 307,000 tonnes/a in 1965. The introduction of the first refinery and the early ones that followed and subsequent developments are discussed in this paper. The major technical achievements during the high growth period from 1945 to 1963 included: the use of fire- refined anodes; the introduction of newly-developed rectifiers; the introduction of lead-lined or vinyl-lined concrete cells; improvements in the electrolyte heating system; application of improved anti-corrosion materials and the introduction of heat-insulating surface coverings. In 2018, Japan’s refined copper production totaled 1,595,000 tonnes of copper, with Japan ranking third in the world after China and Chile. INTRODUCTION In 2018, Japan’s refined copper production totalled 1,595,000 tonnes of copper and Japan ranked third in the world for refined copper after China and Chile (Wood Mackenzie, 2019). There are now eight copper refineries in Japan which provide Japan’s production of refined copper. These plants are listed in Table 1 (Wood Mackenzie, 2019), also showing the production trend since 2013; in general, production over this period has been relatively steady. The Toyo plant is now the 7th largest copper refinery in the world after several large plants as rated by capacity in China, Russia and India (ICSG, 2019).

Jan 1, 2019

By Carricajo T., S. Nazari

This paper discusses a unique dynamic solution for optimizing comminution area in real time. The proposed technology is a combination of model predictive controller, optimization methods and real-time digital twin for mineral processing. For grinding area, the ultimate goal of an effective control strategy is to maximize the mineral processing, reduce energy consumption and stabilized size classification. To have a continuous and effective comminution operation, a reliable control system that minimizes the need for human intervention is necessary. The implemented advanced control strategy for comminution area consists of two layers in hierarchical structure. The lower layer objective is to stabilize the process stabilization using model based predictive controller (MPC) technique. The unique MPC used in this layer is called BrainWave™ chosen for its robustness, convergence and delay handling; furthermore, the developed platform offers stable and repeatable solution and has short implementation time. The upper layer is a supervisory control, which sends the preferred process operation points to the lower layer to drive the process towards optimal performance. The goal of this layer is to optimize the process targets, as well as making decision. To increase process performance, automatic ore characteristic adaptations are implemented in this layer. A digital twin of comminution area is built in the second layer that runs amongst the process in parallel. By communicating, the process data in real time, a virtual densitometer is built that is being used in our advance process control strategy. Our control strategy has been developed, tested and implemented in various areas of a copper concentration plant. The targets for implementing advance process control are usually increased throughput, decrease energy consumption and variability reduction in particle size distribution. Overall, the results showed in different mine sites are increase throughput between 2 and 5%, specific energy consumption reduction about 15% and process variability reduction up to 60%. In this paper we discuss the results of the recent implementation of this technology on Minera Los Pelambres, Minera Collahuasi and CAP Minería. Their concentrator plants have installed control systems with the most updates optimization features, for SAG mill, ball mills and thickeners.

Jan 1, 2019

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

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

Jan 1, 2019