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By J. Bruce See

INTRODUCTION This symposium at Cu 2019 has been organised to honour and celebrate the career and achievements of Dr Phillip John Mackey, an internationally recognised expert in the production of non-ferrous metals, especially copper and nickel smelting and refining. Phillip was born in Epping, a suburb of Sydney, New South Wales, Australia on 19 April 1941 and is one of 4 children of Gilbert and Philomena (nee Bartlett) Mackey. After completing Primary School in Epping, he attended Marist Brothers High School in Eastwood, Sydney and graduated in 1957 as the Dux of the School. Phillip’s father, being on the staff of the Sydney Technical College, in Ultimo, Sydney had the opportunity in the early 1950s to meet the late Professor Sir Rupert Myers, Foundation Professor of the School of Metallurgy at the New South Wales University of Technology (later called the University of New South Wales or UNSW). Gilbert Mackey was impressed by the new young Professor of Metallurgy and encouraged his son to study metallurgy - at the time an expanding industry in post-war Australia. EDUCATION AT THE UNIVERSITY OF NEW SOUTH WALES Phillip commenced at UNSW in 1958, graduating with honours in 1963. His undergraduate education and training included a few years of part-time study while working at British Motor Corporation on sheet metal forming of car bodies. At University Phillip developed his talent for teamwork. He was a contributing writer to the University newspaper, Secretary of the UNSW Students’ Union and was active in the UNSW Metallurgical Society and co-editor in 1963 of “Metallurgical Review,” the annual journal of the UNSW Metallurgical Society. His diverse interests and curiosity about the world around him were likely nourished while at University. Upon completing his Bachelor’s degree, Phillip was awarded a Commonwealth of Australia Postgraduate Research Scholarship. He began his PhD in 1964 on the thesis topic “Studies in Heat Transfer between Dispersed Liquid Metals and Gaseous Atmospheres” working under the supervision of then Associate Professor Noel Warner, an outstanding Professor at the University. Professor Warner is now an Emeritus Professor at the University of Birmingham and is internationally respected for his pioneering theoretical and practical work on established and innovative smelting processes.

Jan 1, 2019

By F. Parada Luna, F. Parada Torres, J. Salvo, A. Reghezza

In the pyrometallurgical processes of copper concentrates, arsenic is preferentially concentrated in the dusts carried by the gases as a result of their volatility and subsequent condensation in the gas cleaning train. In this work results obtained by alkaline leaching of the dusts from Ministro Hales roasting plant, belonging to Codelco, are presented. For the Ministro Hales plant the alkaline leaching would be an alternative to the recirculation of dusts to the roasting furnace. The experimental study considered NaOH and NaOH plus NaHS leaching of two samples of dusts, with 4.24 and 2.55% arsenic respectively. The results show that it is feasible to extract 90% of the arsenic and 50% of the antimony at room temperature using only NaOH, and when the temperature increases to 95ºC, the extractions of As and Sb increase. Copper remains in the solid residue with less than 1% dissolution, and the NaOH net consumption was approximately 140 kg/t of dust. In leaching tests with NaOH + NaHS, two types of methods were tested, baking plus leaching with water and direct leaching with a NaOH+NaHS solution. With the second method arsenic extraction increased to 95% and copper dissolution was less than 0.01% at 85ºC, NaOH 100 g/L, NaHS 90 g/L, 1.5 h and L/S ratio of 5/1 (L/kg). The bismuth remains stable in the solid in all cases. INTRODUCTION The presence of arsenic in copper concentrates has increased with time and the environmental regulations are becoming increasingly more stringent. In concentrate treatment pyrometallurgical plants, most of the arsenic is volatilized condensing in the gas train and is collected along with the entrained solids to form the flue dusts. Under these conditions, in smelters of copper concentrates with high arsenic content, the dusts are being treated in hydrometallurgical plants to extract and stabilize the arsenic, which also means that most of the copper contained in the dust is co-extracted with the arsenic. This affects the recovery of copper in the smelter and an alternative such as the selective alkaline leaching of arsenic, would allow the copper from flue dust to be recovered.

Jan 1, 2019

By Aleksandar N. Nikoloski, Tendekayi Tapera, Glen P. O’Malley

Chalcopyrite (CuFeS2) and enargite (Cu3AsS4) are among the most important copper sulfides. At present, many known deposits of these minerals are uneconomic to process because they are of too low grade, insufficient size or too remote to be cost effectively treated with conventional technologies that can deal with their refractory nature. Therefore, developing an effective, low cost extraction process for recovery of copper from them is important. The research presented in this paper will deal with potentially cost effective strategies involving catalysed processing, studied under conditions relevant to recycle heap leach conditions. The present account summarises some of the main findings from the research program carried out at Murdoch University between 2014 and 2018 that has investigated the effects of the common leaching parameters, major and minor species present in the leach system and additives on the rate and extent of copper extraction from low-grade ores, in which primary copper sulfide minerals such as chalcopyrite and enargite were the dominant copper minerals. The study involved batch leaching tests, detailed structural and mineralogical characterisation and thermochemical modelling under conditions relevant to heap leaching. The results have demonstrated that silver in particular has a strong catalysing effect on the extraction of copper from ore under the leach recycle conditions and, at an adequate dosage, can effectively overcome surface passivation, enabling almost complete extraction of copper from chalcopyrite as well as enargite. The aim of this paper is to present the effect of the studied additives and to briefly elaborate on the mechanism of the silver catalysed leaching process. In addition, the presence of certain secondary agents was shown to lower the critical dosage of silver and promote extraction. The data from the study allowed the development of a theory for the silver catalysed reaction mechanism, which enables further optimisation and cost-effective application of this strategy in practice. INTRODUCTION Researching the extraction of copper from chalcopyrite is important as it is the most abundant copper mineral in the earth’s crust. The majority of ores containing this mineral are processed by flotation to produce a concentrate, followed by smelting and electrorefining. However, most reserves available nowadays are of low grade or limited in size, and therefore not suitable for this conventional approach. Hydrometallurgical processing via heap leaching with or without the aid of bacteria has been shown to be potentially economical. Chalcopyrite is refractory though and suffers from slow leaching kinetics and poor overall extraction, so understanding the mechanism of how it leaches has received significant attention by researchers working in this field over the past 50 years. The refractoriness, which is the main issue with the hydrometallurgical approach, is generally described as passivation of the chalcopyrite surface whereby a surface coating is formed that blocks the reaction interface or wetting of the chalcopyrite, limiting the copper extraction. Exactly what this layer is made of or how the issue could be overcome has been the subject of intense study and debate for many years. One of the limitations of the work to date is that much of the research focus has been on typical initial leaching conditions, while there has been relatively little research into how the leaching is affected under recycle solution conditions where certain metal ions (Fe, Mg, Al) have often accumulated in the liquor. This occurs because the leach solution exiting a heap contains many species from the dissolution of gangue minerals and typically this solution is returned to irrigate the heap, once the copper has been removed, leading to an accumulation of these species.

Jan 1, 2019

By Osamu Inoue, Takashi Yamashita, Takashi Ikemoto

The expansion project in Naoshima smelter and refinery to meet enhancing recycling business of E-scrap was successfully completed in 2016 to have a capacity of 110,000 tpa of E-scrap treatment. Increasing treatment of E-scrap has raised the Ni content of copper anode to 4000 ppm on average, which is double the level of 10 years ago, and to 6000 ppm or more occasionally. The tank house has refined such copper anode to copper cathode with starter sheets. Since most Ni in Copper anode dissolves in copper electrolyte, the Ni content of the electrolyte has increased to 27 g/L and more. Ni accumulated in the electrolyte has increased power consumption during the electrolytic refining of copper, although cathode quality has been controlled satisfactory so far. In order to ensure sustainable operation while growing E- scrap business, the NiSO4 process has been expanded recently so as to recover Ni in the electrolyte more and the pressurized-leaching technology will be introduced to the slime leaching process so as to enhance the capacity of slime treatment and to improve Ni recovery in the near future. INTRODUCTION In recent years, the treatment amount of E-scrap at Naoshima smelter and refinery has increased according to the growth of its recycling business as shown in Figure 1, while impurities from E-scrap have concentrated in copper anode as well. Especially, Ni content in anode has been increasing remarkably since 2012, causing to raise Ni content in the electrolyte as shown in Figure 2. The increment of Ni content in copper anode and electrolyte has adversely affected the electrorefining process, slime leaching process, and NiSO4 crystallization process at the tank house. Such severe conditions provided opportunities to redesign the tank house to establish a robust process to which more Ni is acceptable.

Jan 1, 2019

By S. W. Henley

"Shaft sinking design is typically driven by what has been successful in the past and the availability of sinking plant. This document considers a project where the sinking design and methodology is unconstrained by conventional thinking and removed from the commonly held belief that a sinking methodology should be developed from available winding plant. Schedule was a key driver for the Project and led to an alternative sinking design that aimed to optimise sinking advance rates by specifying new winders that compliment known sinking plant while providing improved safety, redundancy and flexibility within each part of the sinking cycle. KEYWORDS Consistent, Efficient, Flexible, Improved, Innovative, Optimise, Redundancy, Safety, Schedule INTRODUCTION The project is remotely located in the South Gobi region of Mongolia on a brownfield site and the scope of work included the concurrent blind sinking and concrete lining of two (2) ventilation shafts to an approximate depth of 1,150 m at a finished diameter of 10 m and 11 m respectively. All sinking and lining operations to be completed as in-line activities, (no lining whilst other activities are undertaken). The project was schedule driven with a completion date set for the first quarter (calendar) of 2021 and there were no restrictions imposed on the successful Contractor with respect to their equipment selection. Existing sinking plant already onsite from three previous shaft sinking operations would be made available to the Contractor if required on an as is where is basis Sustained advance rates in excess of 2.4 m per day would be required in order to achieve the completion milestone, therefore the design would have to include contingency and demanded flexibility to account for equipment down time. Previous shafts that had been sunk onsite have utilised a single double drum kibble (bucket) winder (hoist) with a single excavator loading on the shaft floor into a large capacity kibble (bucket). DESIGN Mucking The sinking design centred on the mucking cycle given that this is traditionally the most restrictive part of the sinking cycle and typically accounts for one third of the overall sinking cycle. The Vertical Shaft Mucker (VSM) and/or diesel excavator are commonly used plant for shaft sinking that were under consideration. The use of Cactus Grabs was not considered given the higher risk associated with grab ropes and hoist failures."

Jan 1, 2019

By Gerardo Fuentes

Arsenic extraction from Chilean copper concentrates up to nearly 99% was achieved, using hydrothermal reactions. Arsenic appears in nature in more than 200 mineral species, being arsenopyrite (FeAsS), enargite (Cu3AsS4), and tennantite (Cu12As4S13) the most common species. In addition, arsenic is present in a large amount of copper concentrates, produced in northern Chile. Medical statistics show that in northern Chile, the amount of cancer to the lung, skin and bladder are the highest in the country. The latter situation, demonstrates the importance of studying metallurgical processes that do not generate atmospheric pollution. The hydrometallurgical processes such as hydrothermal treatment in autoclave reactors, do not produce gaseous emissions. On the other hand, the arsenic remains in solution from where it can be transformed in scorodite by widely known methods. Previous works about hydrothermal treatment of enargite have shown the feasibility of transforming the enargite of the copper concentrates into chalcocite or mineral species of the chalcocite family such as djurleite, digenite and for this reason enrich the concentrate in copper and pass the arsenic to solution. For this research, two Chilean copper concentrates were used, concentrate 1 with 5.5% arsenic and concentrate 3 with 1.8% arsenic. These concentrates were subjected to a complete chemical and mineralogical characterization. Later these concentrates were carry out to batch tests of hydrothermal treatment. Also, these concentrates were subjected to grinding to have different particle sizes. For the hydrothermal tests, a liter capacity Autoclave Engineer reactor was used. Characteristic of this reactor was the maximum temperature 350°C, 5000 PSI pressure, and 900 rpm agitation speed. Kinetic tests were carried out with the two concentrates maintaining the same solid: liquid relationship, the same agitation velocity and the variables studied were treatment temperature, time of reaction, concentration of copper in the solution and grain size of the concentrates. The conclusions of the study indicated arsenic extraction of the order of 93.98% from the concentrate 1 and arsenic extraction of the order of 98.74% from concentrate 3. After the hydrothermal treatment, the solids and solutions were characterized chemically, by electron microscopy and X ray diffraction. Also the solids were characterized by grainsize through a particle size analyzer (MICROTRAC). Results of batch treatment shows that arsenic can be release almost completely from real copper concentrates using this hydrothermal process and the next stage will be a scale up process in order obtain results for pilot plant design. INTRODUCTION Arsenic is an impurity that appears associated with copper minerals forming enargite, tennantite, tetrahedrite, and that due to the pyrometallurgical treatment either in conversion or roasting furnaces, it goes into the atmosphere together with the sulfur dioxide. A great quantity of literature can be found to eliminate arsenic or try to separate arsenic from copper. The methods used are mainly leaching processes with or without pressure (Padilla, Jerez & Ruiz, 2015) with different reagents like chloride media (Herreros, Fuentes, Quiroz & Viñals, 2003) or chloride media with sulphuric acid (Riveros & Dutrizac, 2008) or sodium hypochlorite (Mihajlovic, Strbac, Zivkovic, Kovacevic & Stehernik, 2007). It is known too that Chile is the first copper producer in the world and a great quantity of its minerals are in the north of the country, in Antofagasta.

Jan 1, 2019

By E. Nolasco, J. Ouellet, T. Tikka

"Shaft bottom structural systems are used in underground mining to support transfer of personnel, equipment and ore to and from underground to the surface. This paper first summarizes the evolution of design philosophy from a concrete tower supported on a concrete plug to the current structural steel design arrangement with unlined shaft walls, clear shaft bottom area with only a concrete housekeeping pad. Then the paper reviews the design parameters, analysis and design, and construction methodology for this rather unique shaft bottom structural system consisting of three sets of hitch beams, with each having specific functional requirements. INTRODUCTION Shaft bottom structural systems are used throughout the world in underground mining operations to provide a means to support the transfer of personnel, equipment, materials, ore and waste entering and exiting a vertical shaft ready for transport to and from the surface via the hoisting system. The shaft bottom structural system, for the Service Shaft (SS) at the Jansen Mine, is designed to accommodate two skips, a personnel/equipment cage, counterweight, services, loading pocket, rope changing, arrestor system, crash (or catch) beams, access platforms, stairs and rope divider system. Stantec and BHP have engineered a unique structural steel shaft bottom system for hoisting system at a Potash Mine. The shaft bottom structural system connects to a fixed guide and cantilever bunton system, used in the vertical shaft (shaft barrel), above the shaft bottom. The geometry for the shaft bottom structural system, shown in Figure 1 a), features three sets of hitch beams, beams which span across the shaft excavation or barrel. At the top of the structural system are two upper hitch beams. These hitch beams support a 64 metre suspended rectangular braced tower (upper tower), complete with steel guides for the conveyances. The upper tower passes through the main station (potash or mining) level and then connects to the top of the arrestor (lower) tower structure via a sliding vertical connection. The potash level is supported by two hitch beams spanning the shaft and loading pocket excavations. The arrestor tower, a braced compression tower structure, is carried by a third set of hitch beams. Below the third set of hitch beams is a short suspended tower complete with access platforms, stairs and the rope divider system."

Jan 1, 2019

By A. Abadías Llamas, N. Bartie, M. Stelter

Evaluating the economic viability as well as the sustainability of the Circular Economy (CE) system requires a deep understanding of the distribution of all elements, compounds, alloys, materials etc. in flows. In this paper, the circularity of the copper value chain, including primary and secondary processing, is rigorously evaluated. The studied system comprises the metallic copper production from primary sources (from mineral to metal), copper-containing commodity production (copper is mixed with other metals) and copper recycling through secondary smelting to close the loop. This is linked to photovoltaic (PV) panels and battery storage. A simulation model of this system is created using HSC Sim, considering more than 30 elements (and its various compounds), 180 unit operations and 800 flows. From the mass and energy balances obtained through the simulation, an exergy analysis is conducted to evaluate the resource consumption from a second law (entropy) perspective. Additionally, these results are complemented through a Life Cycle Assessment (LCA), the recovery of technology elements and by-products is discussed, while quantifying the losses through the value chain. Through the digitalization of the complete system, a better CAPEX and OPEX understanding of the metal recovery and losses can be obtained, as well as the associated resource consumption and environmental impacts. New flowsheets and technologies can be evaluated. Several scenarios show how the resource consumption and the environmental impacts are affected by the recovery of different materials to produce different products. INTRODUCTION A key challenge of the circular economy (CE) is the linkage and optimization of both resource and energy efficiency of the (renewable) energy infrastructure and the resource cycle depicted by Figure 1 simultaneously. In addition, the CE requires the linkage of all stakeholders as shown by Figure 1. This has been discussed in some detail by Reuter and Bartie (2019) and Reuter et al. (2019). Copper and all its associated technology elements is one of the key enabling metals that is an economical copper colored thread through CE society (Reuter & Kojo, 2014). The Metal Wheel (Figure 2) shows the linkage of and recovery of various elements associated with copper and other carrier metals. This reveals that copper acts as an important carrier metal because of its properties i.e. the wealth of technology elements it carries from geological ores, but also its capacity to act as solvent of valuable elements and thence a key closer of various complexly connected element and material cycles when these technology elements are released during copper refining. The same applies to all other carrier metals of modern CE society as shown for example recently in a lead policy brief (Blanpain, Reuter & Malfliet, 2019).

Jan 1, 2019

By J. Holder

"A key barrier to vertical mineshaft construction is the construction access required for workers. A suspended work stage is used to provide this access for the mechanical, electrical and structural applications and other tasks. Work stage design is complicated considering the numerous types of construction activities that are executed simultaneously, the confined environment in which the work stage operates, the significant corrosion attack that is present, the weight limitations imposed by the mechanical hoist systems, and the risks associated with worker safety. Also complicating the design, these structures and their construction environment can vary with time. Hence, they can be described as 4D structures. The author has discovered significant synergies and advantages with integrating engineering design with detailed three-dimensional (3D) fabrication design. The significant reduction of minor design errors needing field correction will be underscored. The tender process is also accelerated since the design output is fabrication shop drawings including a detailed quantity takeoff output which lists each component required and its specific quantity. The paper presented will document the numerous advantages to the project using this approach. INTRODUCTION A key barrier to vertical mineshaft construction lies in the access for the workers to execute the construction. Suspended work stages are used to provide this access for the mechanical, electrical and structural applications as well as other tasks. The design of such work stages is complicated considering the numerous types of construction functions that need to be executed simultaneously, the confined environment in which it operates , the significant corrosion attack that is present, the weight limitation imposed by the mechanical systems for suspension, and the risks associated with worker safety. Also complicating the design, these structures and their construction environments can vary with time. They can be described as 4D structures with the fourth dimension being time. This design is a time consuming process. The post tender award is meticulous and requires strong teamwork between the steel fabricator and the consulting engineer. This teamwork may or may not be present and is dependant on the experience and relationship the team members have working together as well as other factors."

Jan 1, 2019

By R. Grau

"After commissioning and start-up of the first 500m3 TankCell® flotation cell in 2014, the over 300m3 mechanical flotation cells have earned their place in large-scale flotation operations. These 500 to 630m3 cells, designed for projects with large throughputs, are proven to provide savings in project capital and operational expenditure through savings in foot print, energy consumption, and less equipment to install and maintain. Addition to savings in different phases of a project, these cells provides operational flexibility and has shown improvements in metallurgical performance. In this paper a case study from Kevitsa TankCell® e500 installation and Buenavista del Cobre TankCell® e630 installations will be given.INTRODUCTION After development (Grönstrand and Bourke, 2008), characterization (Morales et al. 2009; Yanez et al. 2009; Grönstrand at al. 2009), validation (Elgueta et al. 2009; Coleman and Dixon, 2010), and more than 350 industrial installations of the Outotec TankCell® e300 flotation cell, the well proven energy efficient TankCell® product family introduced the TankCell® e500 forced-air flotation cell late June 2012 (Outotec.com, 2012). The introduction of the 500 m3 flotation cell followed by introduction of a 630 m3 flotation cell, TankCell® e630, in October 2014 (Outotec.com, 2014). These large flotation cells are designed for projects with high throughputs where the number of cells needed for operation with smaller equipment would become large. The benefits of this size of flotation cells comes from the lower capital expenditures due to lower equipment costs, savings in plant footprint, less installation work with less auxiliary equipment, lower energy consumption, and less units per installation resulting in less components, spare parts and maintenance.The first TankCell® e500 was commissioned and taken into operation in October 2014 at FQML Kevitsa Cu-Ni-PGE concentrator, nowadays owned by New Boliden. The first TankCell e630 was taken into operation in January 2018 at GrupoMexico’s Buenavista del Cobre Cu-Mo concentrator. The start-ups of the first TankCell® e500 and TankCell® e630 followed with test work where the hydrodynamic performances of the cells were characterized and metallurgical performance studied. The results of the TankCell® e500 test work were presented in detail by Mattsson et al (2015a; 2015b) and TankCell® e630 by Grau et al. 2018. In this paper the metallurgical improvements achieved with the large cell installations will be presented."

Jan 1, 2019

By Bing Li, Kai Liu, Yongcheng. Zhao

With the ongoing development of copper smelting technology, the improvement of environmental emission control and achieving an automated continuous copper pyrometallurgical process have become the key development goals for copper smelting. The double bottom blowing continuous copper smelting process combine bottom blowing smelting (SKS) and bottom blowing converting (BCC) process and fire refining as “three-connected furnaces” arrangement is a pyrometallurgical process independently developed by China’s ENFI which can continuously produce copper from raw material through to anodes. This technology can improve energy efficiency and reduce environmental emissions for the overall copper smelting and converting process. This paper discusses the engineering characteristics and production practice of the SDIC Jincheng Metallurgical Project, which operates SKS smelting and BCC converting process with “three-connected furnace” arrangement for the first time in the world. Advanced engineering concepts and tools were used to optimize the configuration during the project engineering stage. The operation shows that the process has good environmental performance, reduced energy consumption and high level of automation, which is a new choice for copper continuous smelting process. INTRODUCTION One of the most important developing trends for the overall copper smelting process is to develop a continuous operation. Continuous copper smelting can maximise energy efficiency in copper concentrate and make full use of the rich heat treatment system. Cold charge and a part of low-heat-generating materials, which are processing with additional equipment in traditional copper smelting process, can be processed with rich reaction heat in continuous copper smelting and converting process to achieve good economic benefits. Yuan and Wang (2017) noted that SKS+BCC copper smelting process can reduce the secondary SO2 emissions compared with discontinuous copper smelting process, which required transportation of melted matte by matte ladle and may cause secondary pollution. The continuous copper smelting process can also reduce the generation of intermediate cold materials.

Jan 1, 2019

By M. Takebayashi, K. Takenaka, S. Matsubara, T. Yoneyama

The cathode production of Toyo Smelter and Refinery, owned by Sumitomo Metal Mining Co., Ltd. reached 450,000 tons per year in 2016 with three tank houses. Niihama tank house commenced operations in 1919 with conventional sheet cathode technology. Approximately 100 years of operation experience and the technical development enabled that the current density of Niihama tank house and the Toyo tank house #1 increased to 320 A/m2. These operation techniques were utilized at the Toyo tank house #2 which started in 2005, and its current density reached up to 350 A/m2 with permanent cathode technology. In order to achieve high current efficiency under condition of 350 A/m2 operation at Toyo tank house #2, continued efforts had been made to verticality of anodes and permanent cathodes. To improve verticality of anodes and cathodes, reduction of short circuit ratio by changing the shape of the anode, optimizing the anode pressing machine and off-line process for rectifying the distortion of permanent cathodes was carried out. In addition, concentration control of electrolyte and expansion of Ni removing ability from the electrolyte prevented passivation. Finally, more than 96% of the current efficiency was stably confirmed. INTRODUCTION Today, Toyo Smelter and Refinery of Sumitomo Metal Mining operates three tank houses in Japan, namely, Toyo tank house #1, Toyo tank house #2, and Niihama tank house with a total capacity of 450,000 metric tons of copper cathodes. Figure 1 shows the trend of cathode copper production and the current density at three tank houses over the last 10 years. Cathode production by Toyo Smelter and Refinery reached 450,000 tons per year in 2016 with three tank houses.

Jan 1, 2019

"Yara Siilinjärvi in Finland is the Western Europe's only apatite mine. Annually it produces one million tons of apatite concentrate, which creates 10 million tons of tailings. At the beginning of 2017, Yara replaced their conventional TSF with a new tailings treatment plant, which increased the solids content of the slurry from 45–48% to 66–68%, making it possible to deposit as a paste. This extended the life-span of the current TSF from 2020 to 2035. This paper presents the storyline of the Yara paste plant project. The project began with laboratory testing and then progressed to pilot-scale thickening and beaching trials. After successful pilot trials, the final process design was implemented. Process automation played a key role in the project. An advanced control system is needed to maintain process stability and the desired operation point in various operational circumstances. Since commissioning the paste plant has been running consistently with an underflow solids content of 66-68%, with average beach slope angles of 3.5 degrees in the tailings disposal area. Robust thickener operation also leads to the trouble-free high-density pumping and high availability of the plan. During first year of operation the plant’s availability has been very good, over 97%, and currently, the plant is operated remotely from the main concentrator plant.INTRODUCTION The Yara Siilinjärvi Plant is owned by Yara International, a Norwegian fertiliser company. Yara Siilinjärvi is Western Europe's only apatite mine, and it primarily produces apatite concentrate for its Siilinjärvi phosphoric acid plant. A portion of the concentrate is also dried and shipped to other Yara fertiliser plants. The site annually produces one million tons of apatite concentrate, which creates approximately 10 million tons of tailings. Extending the lifespan of the 1150-hectare tailings storage facility has been studied in Siilinjärvi for a long time. The aim has been to use the same tailings area footprint, without any new land purchases or major dam increases. An important part of the process has been to study the possibilities for using new technologies, especially thickened tailings technology (Ruhanen, 2014)."

Jan 1, 2019

By J. K. Armstrong, C. Feyerabend

"As mines are constantly growing and expanding, so do the challenges for emergency response teams. They have to keep up with the pace set by current and future production techniques. Although mines have now reached depths of several kilometres and their dimensions and complexity are that of an entire city, mines rescue still must be able to reach the most remote areas within the mine in a swift and safe manner. The duration of breathing apparatus and physical human capabilities are limitations. Therefore, alternative methods have to be explored to extend the time under oxygen for mine rescue teams without putting additional stress on the team members. To address this global trend, Dräger, Goldcorp and PAUS have partnered to develop a mine rescue vehicle that allows mine rescue teams to travel for an extended period of time through a contaminated atmosphere without the use of breathing apparatus, thus extending the range of their mission. By applying currently available techniques, the safe expansion of mine production can be ensured, giving mine rescue teams the ability to respond to accidents in a safer and faster way. INTRODUCTION With the reduction of easily accessible natural resources, there is a global trend toward searching for higher ore grades underground and/or continually expanding mines underground to extend their lifetime. While in some areas this trend has just begun, in other countries mines are reaching depths or horizontal dimensions that pose extreme challenges (Codelco, 2016). These challenges include safety relevant topics, such as rock stability, heat and seismicity, but also more basic issues such as distances to the working face, which can become crucial considering the cost to develop a new shaft or decline. To overcome this, there is a growing interest in collaboration among various stakeholders to drive the development and support of deeper mines and tackle the complexity of underground mining. This is evidenced by the fact that several industry and research groups and networks have formed, e.g. the I2mine project in Europe and the Ultra-Deep Mining Network by the Centre for Excellence in Mining Innovation in Canada. Due to commodity pricing and the trend toward renewable energy, metal/non-metal mining will be the driving force behind this development."

Jan 1, 2019

By F. Prescott, D. Hiam-Galvez, J. Hiam

Mines are frequently located in remote areas with little conventional employment and few opportunities for the existing population. The development and operation of the mines results in several years of intense activity followed by a near-complete reduction in employment and opportunities after the mines are closed. Designing Sustainable Prosperity (DSP) is a method for rectifying this situation by designing for long-term growth in areas that host mines. This process involves the participation of local and national governments, local community, mining companies, investors, academics and those with sector expertise. The mines will be the catalyst for regional sustainable development. If successful, long-term economic and environmental prosperity should be the result for the areas affected by mining and, in some instances, promoting the regions as centres of excellence for a particular industry. This paper will describe how the concept works using the copper producing region of Peru and Chile as an example. Designing sustainable prosperity starts by looking at regions based on the natural resources and skills available, the infrastructure and possible energy sources. Integrated natural resource models and innovative market studies, followed by education and skills requirements, are then established to determine the potential for the region and what needs to be done to realise the possibilities. INTRODUCTION Industry 4.0, or i4.0, dubs the fourth and current industrial revolution. Where previously they have been founded on ever-increasing consumption and production, i4.0 distinguishes itself by its partial focus on reducing resource intensity; lightweighting, energy efficiency, sensors and automation to improve resource efficiency. Industry 4.0 is an era of artificial intelligence, genome editing, biometrics, renewable energy, 3D printing, autonomous vehicles and the Internet of Things (Lu, 2017). These drivers represent key opportunities for the economic future of nations and regions.

Jan 1, 2019

By L. Iacopetti, A. Calderara, M. Perego, P. Perrone, T. Furusawa

De Nora, a global leader in the development, manufacture and sale of all DSA® anodes based on titanium coated with mixed metal oxide (MMO) for electrochemical processes will present key advancements made to Self-Protected Anodes (SPA) for metal electrowinning. (DSA® is a registered trademark of Industrie De Nora S.p.A., Italy). In 2017 pre-commercial demonstration cells were operated in industrial environment. The program’s focus was to fully demonstrate the SPA anode’s superior mechanical and electrochemical robustness during normal and induced dendrite shorting occurrences while providing quality of plating in commercial operation. De Nora specialists focus on key components, in particular the engineering of the self-protection system, DSA® anodes and catalyst formulation capable to cope with high levels of pollutants while providing long service life in the meantime. Results of field trials confirmed the expected anode performance and enabled the selection of the most suitable “SPA” structure, based on resistive protective elements. Summarizing the economic benefits to copper electrowinning tankhouse operation with De Nora’s self- protected all DSA® anodes are; Double-digits power savings Cobalt-free electrolyte operations Minimal dendrite shorting damage incurred at the SPA anode Smoother, uniform cathode deposits with tangible improvement on current efficiency Tankhouse operations producing lead-free copper quality cathodes, in addition to elimination of lead waste for improved operator health and environment. INTRODUCTION De Nora’s engineered “Value Technology Package” (“VTP”) for electrowinning, consisting of our DSA® anodes coated with mixed metal oxide specifically formulated for copper electrowinning in conjunction with an advanced self-protected full titanium anode (“SPA”), underwent an intense field qualification program, inclusive of the behavior under massive short-circuits as part of the field investigation, with the goal to select and appoint the ultimate design configuration of “SPA”. Optimized wireless-based monitoring system was included in the field program, demonstrating good functionality for cell parameters monitoring and remote shorts detection.

Jan 1, 2019

By L. Hanschar, C. Bwalya, E. Mokola, K. Musonda, A. Makayi

This paper describes the installation, commissioning and operation of porous plugs featuring nitrogen gas purging in both Anode Furnaces at the Kansanshi Copper Smelter (KCS), a greenfield copper smelter which commenced operation in March 2015. The decision to install porous plugs was taken in November 2016 and the facilities were successfully commissioned on both anode furnaces in October 2017. The project was successful and has demonstrated the following process improvements: 12.2% reduction in main burner fuel consumption Larger average charge size, due to fewer accretions; 17 minute decrease in average refining cycle times Improvements in slag skimming operations Improved casting rates INTRODUCTION Porous plugs The application of porous plug technology in copper smelter anode furnaces is well established around the world. Vukovic and Gamweger (2017) and Wallner (2004) noted that the stirring and mixing action induced by the injection of inert nitrogen gas into the molten copper via porous plugs has potential benefits in the following areas: Homogenization of the temperature and composition of the molten copper resulting in Reduced fuel consumption Less tendency to overheat slag (reduced risk of refractory attack) Less slag produced with a lower degree of oxidation (lower rate of refractory wear) Less tendency for buildups to form Improved environmental performance (produce low sulphur blister at Peirce Smith converters) Elimination of buildups inside the furnace and potential for larger charges per cycle Reduced processing times and potential for shorter cycle times More efficient skimming of anode slag Shorter skimming time Less copper carry over to Peirce Smith converters Very little slag remaining in the furnace (reduced potential for buildups to form)

Jan 1, 2019

By S. Hunt, D. de Vries

Kansanshi Copper Smelter began operation in March 2015. The smelter features a single Isasmelt furnace for smelting, one 6-in-line electric furnace for matte-settling and slag-cleaning duties, four Peirce- Smith converters and two rotary anode furnaces. A 1,450 t/day oxygen plant and a 4,085 t/day sulphuric acid plant are included in the smelter operations. This paper gives an overview of the operational performance to date with summary data for each smelter section. Challenges such as an unforeseen refractory failure in the Isasmelt furnace in 2016 are discussed. Design improvements around the electric furnace are mentioned. Plans for potential future improvements are put into context. A consistent focus on operations, maintenance and continuous improvement has seen the smelter easily achieve its nameplate capacity of 1.2 million tonnes per annum of copper concentrate throughput. The smelter now operates consistently at a concentrate throughput rate of 1.38 million tonnes per annum of copper concentrate. INTRODUCTION The smelter was commissioned in March 2015. Design and commissioning of the smelter have been discussed by de Vries, Hunt, Hanschar and Louie (2016). The smelter rapidly ramped up to achieve nameplate production (1.2 million tonnes per year concentrate) and in 2018 operated consistently at the rate of 1.38 million tonnes per year concentrate. This paper presents major operational data for each section of the plant and describes significant events and improvements during the first four years of operation. The smelter flowsheet is shown in Figure 1.

Jan 1, 2019

By G. Stokreef, B. R. Davis, M. Kadiyski, Alvear Montenegro, H-P. Loock, V. Montenegro

An optical probe for burner monitoring in copper flash smelting, first developed at the University of Concepcion, is being industrially tested in a collaboration between Kingston Process Metallurgy, Aurubis AG, and Queen’s University. The sensor uses fibre-optics and optical emission spectroscopy to analyze the light emitted from combusting particles in the furnace reaction shaft. Information that can be extracted from the emission spectrum includes the temperature and brightness of the flame, as well as emission lines or bands from excited atomic and/or molecular species. Laboratory proof of concept testing of the sensor was first completed in a continuously fed drop tube reactor using O2/S stoichiometric ratios between 0.8–4.0. These O2/S stoichiometries capture the industrial operating condition, which is typically less than 1. The results from lab testing justified a plant trial of the optical probe, which was done in the Aurubis Hamburg flash furnace in September 2018. From this trial, approximately 31 hours of spectral data were collected and synchronized with furnace operating data. The results from both drop tower test work and the plant trial are discussed with implications for process monitoring in the commercial furnace. INTRODUCTION Flash furnace smelting is a continuous feed, batch tap, pyrometallurgical process. It is used to treat copper sulphide concentrates and small amounts of secondary materials, such as dust generated by the process. O2 enriched air and the concentrate blend are fed into the furnace reaction shaft through the concentrate burner. As the particles fall through the reaction shaft, they heat up and ignite, causing rapid, exothermic reactions that oxidize the iron sulphide, some copper sulphide and impurities in the feed. Reacted particles fall into the settler of the furnace, melt and separate into a matte and slag, while the SO2 enriched off-gas is carried out of the furnace for heat recovery and H2SO4 production. Operation of the furnace is achieved by adjusting a limited number of control variables. The main variables are the blast feed rate, blast oxygen enrichment, SiO2 flux feed rate and CH4 combustion rate (CH4 is used when the chemical energy of the concentrate is low) all measured relative to a set feed throughput. (Davenport & Partelpoeg, 1987; Davenport, King, Schlesinger & Biswas, 2002) The batch tapping nature of flash furnaces makes operation difficult because these control variables impact each other, and product quality can only be measured when matte and slag are tapped from the furnace. This means that operation of a furnace is based on feedback control from matte and slag assays. Because of the time-delay associated with understanding product quality, there is interest in using technology to monitor the combustion reactions that happen in the furnace reaction shaft.

Jan 1, 2019

By R. Stauber, B. Friedrich, S. Hilgendorf, G. Homm, C. Gellermann

Due to recent regulations limiting tolerated lead concentrations in copper alloys, new recycling strategies have to be considered. This study investigates the suitability of technical approaches in order to recycle leaded copper alloys exemplified by CuZn39Pb2 and produce an alloy in compliance with future lead regulations in the European Union. Three pyrometallurgical methods are considered to achieve this goal, namely vacuum distillation, intermetallic precipitation and fractional crystallisation. The impact of essential process parameters i.e. pressure, temperature, time, atmosphere and quantity of additives are identified using thermochemical and kinetic modelling. An experimental implementation reveals a general suitability of all three refining processes. However, the individual efficiency of the different methods strongly differs. Thus, a detailed assessment is carried out in order to provide an evaluation on the efficiencies of these methods from a technological, economical and environmental point of view. Taking a possible industrial realisation into account, the authors recommend the approach of vacuum distillation, although the energy demanded is comparatively high due to the enthalpy of evaporation for zinc, or rather, lead. However, required lead concentrations can be reached quickly without introducing an external element to the material. The prediction of selectivity and evaporation speed, which are immanent variables for a successful implementation, can be validated experimentally. INTRODUCTION In 2016 approximately 1 million tonnes of leaded copper alloy semi-finished goods with an average lead content of 2.1 Ma% have been produced in the European Union (EU). The global production in the same period exceeded 5.5 million tonnes, while the lead concentration has stayed constant at around 2.4 Ma% for the last 50 years. Up to 80 % of the current production volume is covered with end-of-life scraps, which are used for direct re-melting, depending on the alloy composition. Lead has been a popular alloying element in copper alloys in general, and in particular in brass due to its positive impact on various mechanical, physical and chemical properties, especially machinability. Health and environmental concerns were drivers for many governments to implement limitations on lead. In order to further utilise the resource potential of leaded copper scrap, it is inevitable to develop methods to efficiently eliminate lead from circulation as far as possible. (Glöser, Soulier, & Tercero Espinoza, 2013; Soulier, Glöser-Chahoud, Goldmann & Tercero Espinoza, 2018)

Jan 1, 2019