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By J. L. Bowlby

TWO years ago, at the annual meeting of this Society, the writer presented a paper dealing with various aspects of the water problems existing in the Cape Breton coal-mining area. Much of that paper was of a theoretical nature, in that an attempt was made to get at the fundamental principles underlying the problems, in order to establish a sort of background against which future results could be laid. Up to that time, only Central power plant in Glace Bay had been under observation, and, since the work had been in progress only a year, it was recognized that there were strong possibilities that certain variations in procedure might have to be introduced as w; became better acquainted with our problem. We were by no means certain that we had covered all the facts of the case. The present paper will therefore be of the nature of a progress report, rather than a formal contribution to the Society. Subjects of a theoretical or academic nature will be avoided as far as possible, attention being confined chiefly to the results obtained during the past two years. In looking over the work with low-pressure boilers, there are certain general conclusions that can apparently be safely drawn. Wherever possible, proper operating conditions will be expressed in terms of appropriate units. Absence of corrosion in boilers where the alkalinity has been maintained within the pH range 11.0 to 11.4 forces one to the conclusion that these are safe working limits. Previous experience had shown that corrosion would occur if the alkalinity were much reduced. Higher alkalinities are un-necessary, more expensive, and are said to introduce the possibility of pitting. We made no attempt to obtain quantitative data in respect to pitting. It is said to occur at pH 12.0. Consequently, conditions of alkalinity were main-tained well below this figure. In the above paragraph, the pH range 11.0 to 11.4 applies to boiler waters which have received the lime and soda treatment. The general practice, when this treatment is employed, is to use 1.0 lb. lime to 7.0 lb. washing soda. At Central power plant, soda-ash is at present substituted for washing soda, the equivalent proportions being kept the same. But not all the plants are on the lime and soda treatment. At Waterford Lake, a 1: 1 mixture of soda-ash and trisodium phosphate is satisfactorily employed, and in Sydney Mines and at No. 3 power plant in Sydney, good results are being obtained at a somewhat greater cost with a commercial phosphate treatment.

Jan 1, 1936

By S. Yano, S. Takagi, Y. Nagaya, T. Akai, T. Ishida, Z. Bennour, T. Suzuki

"It is often reported that the concept of single-plane-fracture is not sufficient to account for the productivity from shale gas/oil wells. Many researchers introduce the concept of the stimulated reservoir volume to describe observed production behaviour. However, the characteristics of the stimulated reservoir volume are not well understood. In this study, to investigate the characteristics of fractures and surrounding region induced by the hydraulic fracturing, a devised hydraulic fracturing experiment was conducted using the cylindrical shale specimens in the laboratory.Two resins were prepared as the fracturing fluid for the experiments: cyanoacrylate, an instant glue, mixed with a fluorescent paint (resin A), and methyl metaacrylate, a thermosetting acrylic resin, mixed with a fluorescent paint (resin B). These resins were able to fix within the specimen after fracturing. This is because the present study aimed to only detect fractures induced by hydraulic fracturing. Cut sections of the specimens were observed under ultraviolet light irradiation. It is expected that the hydraulically induced fractures and the surrounding regions will be detected because the induced fractures filled with the resin should emit light, while the other parts will not.The specimens, which were collected from the Kushiro Coal Mine in Hokkaido, Japan at the depth around 275 m, were roughly 85 mm in diameter, 170 mm in length, and cored normal to the sedimentary planes. An injection hole with a 10-mm diameter was drilled onto the center of the specimen parallel to the sedimentary plane to simulate hydraulic fracturing in shale gas development. The experiment was conducted under a uniaxial loading condition of 3 MPa and the fracturing fluid was injected into the sealed injection hole at a constant flow rate using a syringe pump.After the fracturing experiment, the resins were fixed in the specimens and then the cut sections were observed under ultraviolet light irradiation. As a result, the hydraulically induced fractures, which were filled with the resin, are clearly observed. Detailed microscopic observations show that the main fractures are tortuous and are accompanied by many thinner ramified fractures. Additionally, fractured regions where the resin penetrates significantly are observed around the main fractures in the specimens fractured by the resin B. These induced fractures and fractured regions are considered to be the stimulated region in which permeability is improved. The necessity of estimation on the features of fracture tortuosity and ramified fractures is indicated for more accurate understanding of the well production and the stimulated reservoir volume."

Jan 1, 2015

By Z. Zhang, X. -G. Chen, F. Breton, A. Arici

"The microstructure of Al-10Si-0.3Mg-0.5Mn alloy with 0.27% Zr and 0.19% V addition was quantitatively analysed and compared to the base material. The addition of Zr and V in the AlSi10Mg alloy exhibit a beneficial effect on the Si modification, resulting in the reduction of the size of eutectic Si particles on both as-cast and T6 conditions. The quantitative image analysis results show that the Zr and V addition and T6 treatment can reduce the size of the intermetallic phases. In addition, the presence of Zr and V is favour to the partial dissolution of the intermetallic phases during the T6 treatment. Different heat treatment procedures were tested to optimize treatment parameters. The mechanical properties of the alloys were evaluated using hardness measurements and tensile tests. The results show that the addition of Zr and V can remarkably improve the mechanical properties on as-cast and T6-treated conditions. The correlation between the microstructure and mechanical properties of the alloys is discussed.INTRODUCTION Al-Si-Mg foundry alloys have been widely used in automotive and aerospace industries due to their high strength-to-weight ratio, heat-treatable capability and excellent castability. The alloying elements such as Cu, Mn, Fe are traditionally added into Al-Si-Mg alloys to improve their mechanical properties (Mohamed, Samuel, Samuel & Doty, 2009). In recent years, the attention has been paid to the effect of the transition elements such as Zr, V, Ti, etc, as minor alloying additives, on the properties of aluminum alloys (Elhadari, Patel, Chen, & Kasprzak, 2011; Kasprzak, Amirkhiz & Niewczas, 2014; Shaha, Czerwinski, Kasprzak, Friedman & Chen, 2015; Shaha, Czerwinski, Kasprzak & Chen, 2014; Mahmudi, Sepehrband, Ghasemi & 2006; Meng, Cui, Zhao & Zuo, 2013). Kasprzak et al., (2014) and Shaha et al., (2015) reported that additions of Zr, V, and Ti in Al-Si-Cu-Mg cast alloys can improve the alloy mechanical properties at high temperature by forming stable intermetallics and precipitates on as-cast and heat treatment conditions. On the other hand, the mechanical properties of near-eutectic and eutectic Al–Si cast alloys depend not only on the chemical composition but also, even more importantly, on microstructural features such as the morphology and size of eutectic silicon and intermetallic particles, which are present on both as-cast and heat treatment conditions (Mohamed et al., 2009). However, there is rather limited information on the effect of Zr and V on the microstructure features and mechanical properties of near-eutectic Al-Si-Mg cast alloys."

Jan 1, 2018

By Andrew Artzer, Michael Moats, Jack Bender

Antimony and bismuth in copper electrorefining electrolytes can cause cathode contamination, leading to the careful control of their concentrations. This study examined the laboratory scale removal of bismuth and antimony from an electrorefining electrolyte using a solvent extraction mini-pilot plant. The ability of a proprietary phosphonic acid ester extractant (REX-2) in a commercial diluent was studied in a simulated commercial setting. Antimony and bismuth were both extracted effectively from the electrolyte (72% Bi and 70% Sb removed) and Bi was effectively stripped (92% Bi recovered) with 400 g/L sulfuric acid. A final strip solution was obtained from the SX circuit containing 0.36 g/L Bi, 0.043 g/L Sb, 1.3 g/L Cu, and 0.18 g/L As. Unfortunately, most of the extracted Sb remained in the organic phase. Following the pilot study, experiments were performed to examine the precipitation of bismuth from the aqueous strip phase and the removal of antimony loaded on to REX-2. INTRODUCTION For many years, antimony and bismuth have been some of the most problematic impurities in copper electrorefining (ER) (Bardwell & Lapee, 1933). Along with surface roughening, high concentrations of Sb and Bi can contaminate the copper cathode and precipitate at the copper grain boundaries. This drastically reduces the drawing ability of the metal (Bender & Emmerich, 2016). For these reasons, it is necessary to control Sb and Bi concentration in the cathode, which is done by controlling their concentration in the electrolyte. Conventionally, Sb and Bi concentrations are controlled by maintaining a proper As/(Sb+Bi) molar ratio in the anode and arsenic concentration in the electrolyte. Antimony and bismuth are removed during electrolyte purification using liberators (electrowinning cells) and/or ion exchange. Liberators treat an electrolyte bleed stream from the ER circuit and electrodeposit increasingly impure copper as the copper concentration of the solution decreases. Traditionally, first stage liberators reduce the copper concentration from ~45 g/L to ~15 g/L and produce saleable copper. Second stage liberators decrease the copper concentration to ~8 g/L (Biswas & Davenport, 1980) which results in contaminated cathode that is returned to anode production (Biswas & Davenport, 1980; Wang, 2004). Further electrolyte purification is not always required but third stage liberators can be used to reduce copper concentration below 2 g/L which results in significant removal of As, Sb and Bi and very impure deposits which are recycled to a smelter or used for As production. However, liberators do not have very high energy efficiency (Acharya, 2014; Ando & Tsuchida, 1997) and below 0.5 g/L copper, highly toxic arsine gas may form (Biswas & Davenport, 1980).

Jan 1, 2019

By S. Bantubani, I. Mc Dougall, S. Arjun, H. Joubert, R. Hansraj

The requirements determining the design of the Söderberg electrode columns used in base metal applications are different to those of other industries. There is a strong trend to position the contact shoes above the furnace roof in order to reduce the risk of water leakages directly onto the furnace bath. Tenova Pyromet has adapted its proven electrode column design for application on base metal electric furnaces based on these requirements. The design modifications implemented as well as considerations regarding the overall electrode-roof-furnace arrangement are discussed. Equally important is the management of the Söderberg electrodes to prevent over or under baking. Tenova Pyromet has developed the AutoFurn™ furnace controller to not only optimise furnace operations, but to improve electrode management. An overview of the electrode management control philosophy implemented by AutoFurn™ is provided, focusing on intelligent electrode regulation, slipping and baking. INTRODUCTION Tenova Pyromet has developed an electrode system for use in base metal furnaces, which builds on the experience that Tenova has gained during the past 30 years with electrodes in the ferroalloy and base metals industries. In base metals applications, an important safety consideration is the minimization of water within the furnace and on the furnace roof. In addition, direct contact between copper equipment such as contact shoes and pressure rings and the contents of the furnace is discouraged. In a 3 phase AC furnace, three electrode columns transfer the electrical energy to the furnace bath. Both Söderberg and prebaked carbon or UHP graphite electrodes are in use in base metal furnaces. In each column, the electrode is held in position by a hydraulic slipping and regulating device, a pressure ring and contact shoes. The electrodes can operate in a submerged-arc or brush-arc configuration, where the electrode is immersed in the slag or a short distance above the slag (Mc Dougall 2013). The electrode equipment is divided into an upper and a lower electrode assembly. The self-baking Söderberg electrode is made from lengths of steel casings that are added to the top of the electrode column and welded into place. Carbon electrode paste is added to the inside of these casings and is baked to form a solid tip in the region of the contact shoes. The working platform that surrounds the area where new electrode casings are added, is electrically isolated from the building and each electrode is isolated from the others to ensure the safety of the personnel who perform the electrode casing additions.

Jan 1, 2019

By Michael Klichowicz, Holger Lieberwirth

"For a detailed prognosis of mineral comminution processes, simulation tools like DEM still pose various challenges. Less realistic mineral structures can affect the results of simulation models such as bonded-particle models (BPM). Hence, in order to design mineral processes with these models as tools, representative mineral input data should be used. The article shows, how realistic 2D-structures can be modelled by taking advantage of the parameters identified by the Quantitative Mineralogical Analysis (QMA). Important quantitative parameters can be derived by analyzing three orthogonal thin or polished sections. For this reason point, line, and area analysis are applied. The mentioned parameters precisely describe the texture and structure of minerals. The newly developed software is able for the first time to semi-automatically analyze images of mineral structures using QMA. Moreover, statistically equivalent structures of minerals can be modelled using the QMA- parameters. Generating synthetic, but equivalent mineral structures opens the door for more reliable BPM based simulations. Within future research, the procedure will be adopted to generate 3D-structures. Therefore, QMA-based routines that already have been applied to generate 2D-models shall be utilized as basis. The aim of this procedure is to synthesize realistic microstructures based on certain statistical figures in order to improve comminution simulations.INTRODUCTION Due to its increasing capability, the use of computer technology is permanently intensifying in the design of mineral processes. The recent past has impressively shown that computer models can enhance the understanding of mineral processing. There is in fact a rising number of research activities, which are using simulation approaches to study various aspects of it. Weerasekara et al. (2013) give a comprehensive overview over the application of discrete element modelling (DEM) in the science of comminution.Nonetheless, using simulation tools such as DEM in general still poses a challenge. For a detailed simulation of a mineral comminution process with DEM, bonded particle models (BPM) appeared suitable in the past. The following issues, however, affect an appropriate prediction when modelling with BPM:• Simulations with high numbers of particles may need a lot of time for computation. • Internal modeling parameters like the spring and damper constants in DEM have to be adjusted by comparing its specific crushing behavior in the simulation to reality.• In order to use DEM for prediction of mineral processing, it is essential to validate the correctness of the models and assumptions, which is sometimes difficult. As Weerasekara et al. (2013) have stated it is crucial that the DEM method as well as the DEM code work properly.• It is challenging to implement the complex geometric structures of real minerals into the model."

Jan 1, 2016

By T. C. Chudy, C. C. Hamilton, G. R. Lane

The selection and provision of mineralogical services to assist in developing a GeoMetallurgical program for a Cu Porphyry Mine in South America is covered in this paper and presentation. The objectives of the study were to provide a rapid mineralogical scoping protocol to support and confirm results of sequential Cu leach data and to provide detailed mineralogical analysis of four composites and products on which locked cycle test-work were performed. Results showed good reconciliation was achieved between measured chalcopyrite and secondary Cu mineral contents with leach data using a linear intercept mineralogical procedure by TIMA (Tescan Integrated Mineral Analyser), as well as a generally positive correlation between pyrite and Cu-sulphide grain size. Size-by-size analysis of Composite feeds and test products was used to outline comparisons between feeds on a mineralogical basis and provide liberation and association data with which to diagnose and better explain performance. Identifying key mineralogical drivers to Cu recovery and concentrate grades using these results, a greater understanding of the role of chalcocite and pyrite on metallurgical performance was established. INTRODUCTION As a global supply of copper, porphyry copper deposits are typically very large, often in excess of a billion tons of ore (Keeney, Walters, & Kojovic, 2011; McMillan & Panteleyev, 1980) with grades typically between 0.2 and 1.2% Cu, and up to 2% Cu in places. They are also a source of gold. The complexity of copper porphyries arises from their classically zoned nature, the type of host rock, as well as whether or not they have been overprinted by oxidation, weathering and supergene enrichment (Sillitoe, 2011). Zones of alteration and other geological attributes give rise to varied mineralogy, ore mineral textures and material properties, all of which result in different processing characteristics. Due to the interdependencies of geological and processing characteristics, it is vital to study links between the geology at the macro-scale and mineralogical features down to millimeter and micrometer scale (mineral content and texture) and their influences on mineral processing behaviour (Cropp, Goodall, & Bradshaw, 2013). The growing field of GeoMetallurgy aims to provide an Orebody Knowledge framework by which to gather such information as it pertains to: (1) Understanding and predicting likely flotation response; (2) Guiding relevant test work to investigate and validate links, interdependencies, theories and data interpretation/evaluation; (3) Populate a block model with relevant information and; (4) Implement and refine the model for ongoing optimization and control. Specific questions addressed by using mineralogy often include flotation diagnostics as aids to interpretation, causes of dilution and losses to tailings, as well as guidance to sampling for test work (Cropp et al., 2013).

Jan 1, 2019

A Joint Venture between Hope Bay Gold Corporation and Miramar Mining Corporation recently acquired the Hope Bay project in Nunavut, which represents the first discovery of a major new Archean greenstone gold belt in Canada since the discovery of Hemlo in 1982 and the Yellowknife belt in 1938. Located some 685 km North of Yellowknife, the Hope Bay Belt is geologically comparable to the other major gold producing regions of Canada. The initial discovery of gold mineralization in 1992 led BHP to acquire virtually all of the Hope Bay belt. Over the subsequent 6 years, BHP expended C$85 million on exploration, including 180,000 meters of drilling, 2.3km of underground workings and a 27,000 tonne bulk sample, as well as extensive environmental and economic studies. Hope Bay Gold and Miramar acquired the Hope Bay property in December 1999 for a cost of C$27 million, considerably less than the C$85 million spent by BHP in exploring the property. They both hold 50% of the project, and share expenses equally. Hope Bay Gold is the manager of all exploration phases and Miramar will manage feasibility, construction and operations. The Hope Bay Joint Venture controls 1,117 sq. km., covering virtually the entire 80 km long Hope Bay and adjacent Elu greenstone belts. We believe that Hope Bay is one of the best-undeveloped gold projects in Canada. Work to date has defined a global resource in excess of three million ounces of gold in three principal deposits, each of which is open to extension along strike and to depth. A C$17 million work program for year 2000 focused on advancing a significant portion of the gold resource to a measured and indicated category. This will provide the basis for initiation of a feasibility study in 2001. The 2000 drill program was one of the largest undertaken in Canada last year, and is a clear indication of our confidence that Hope Bay will develop into a major gold producing district. The majority of the potential mineable areas known to date lie within 350m of surface and would be accessed through a combination of open pit and declines. Currently, development of a 1,800 to 2,000 tpd gold mine producing in excess of 300,000 ounces of gold per year at cash costs below $US200 per ounce is envisioned. In addition, the exploration potential on the 80 km gold belt at Hope Bay is clearly tremendous.

May 1, 2001

By C. W. Chao, M. Melashvili, F. Hilário

The processing of a bulk concentrate containing copper as chalcocite, enargite, bornite and chalcopyrite along with pyrite and precious metals of silver and gold was studied. A leaching study was undertaken to examine the application of the atmospheric acid leaching process to a bulk copper concentrate for copper recovery followed by cyanidation or thiosulfate leaching for gold and silver recovery. The extraction of copper exceeded 90% under some conditions and the recovery of gold and silver reached a maximum of 44 and 45%, respectively. A sample of bulk concentrate was roasted in air to oxidize the sulfides with release of sulfur dioxide followed by acid leaching of the roaster calcine and cyanidation of the acid leach solution. The copper extraction reached over 92% and the gold extraction exceeded 88%. The sulfur dioxide may be captured for acid production. INTRODUCTION The processing of a large primary copper deposit in Indonesia is under investigation. One of the possible routes for processing involves flotation of a bulk copper-gold concentrate followed by treatment by an atmospheric leach/SX-EW process for copper recovery. The atmospheric leach residue would then be leached for recovery of gold and silver. A second route involves roasting of the concentrate followed by acid leaching and SX-EW for copper recovery. The leach residue from the roaster calcine would then be leached for recovery of gold and silver. The roaster off-gases could be used for sulfuric acid manufacture, enabling the production of a valuable by-product. The Cobre Las Cruces (CLC) process (Munoz, Garcia, & Delgado, 2017) was designed to treat a resource of copper discovered close to Seville, Spain. The process was developed by Outotec with MK Gold (original project owner). INMET Mining subsequently purchased the deposit from MK Gold and built the copper processing facility. The CLC process is similar to the Mt. Gordon Copper process; except that OKTOP reactors have been installed instead of a low temperature, low pressure autoclave process. The plant was designed to treat 1.3 mtpa of ore at a grade of 6.2% Cu to produce 72,000 tpa of cathode copper over a 15 year project life. The main gangue mineral at Cobre Las Cruces was pyrite. The Cobre Las Cruces ore is therefore similar to a rougher flotation concentrate containing copper minerals and pyrite. The roasting of a rougher flotation concentrate can be designed to produce hematite, basic copper oxides and sulfur dioxide off gas. The acid leaching of the basic copper oxides may produce soluble copper for SX-EW recovery of copper. The sulfur dioxide off-gas may be used to make sulfuric acid.

Jan 1, 2019

By W. Petruk

"D. Doyle How large is the beam to which you referred? Are we considering Watts, Griffis the probe of a 1 micron sized particle?& McOuatReply Yes. The beam would be about 1 micron. We could enlarge or reduce it depending on the current in-put.D. Doyle Then chemically bound material we might not see. If you are restricted to 1 micron then you would not be able to separate things smaller than this. Is this not so?Reply We won't be able to analyse it but we will see it as the scanning electron microscope has a resolution down to about 0.1 microns. We should be able to get the general composition of such a grain plus its surrounding grains. We feel we can then make a reasonable estimate if it were as small as 0.1 microns. An accurate analysis would be achieved with grains of 3 and 4 microns.D. Doyle Would this have an application in the hydrocarbon field?Reply The lowest atomic number would be sodium. We wouldn't be able to go below that atomic number. Anything with sodiumor a larger atomic number would enable us to analyse the carbon.M. Fairwether Do you have any policy relating to priorities. For instance areCominco companies going to be limited in the amout of samples permitted to give everyone a chance because quite frankly our company, and I'm sure all base metals companies can hardly wait for the installation of this equipment. I can forsee two years back-log of work in the first month.Reply Quite frankly, I haven't looked at this. At the present time the work is under the START program but I suppose we will have to look at possible changes. Anything under the START program has first line priority.B. Thorndycraft Will the briquetting of samples be automated as well? Pinson MiningReply No, not at this point. Our polish section laboratory is first-class and can turn out 18 polish sections a day which is more thar we can handle.B. Thorndycraft With the capability of this machine to operate in a hands-off manner perhaps it could start to look for the 5 micron gold particle in tailings?Reply I have such a program using a wave-length spectrometer rather than the energy dispersive. The wave-length spectometer is about 10 times as sensitive and would scan the section much faster. A one inch diameter polished section would be scanned in about 7 hours for anything 1 micron or larger."

Jan 1, 1984

By Louise Corriveau

Exploration strategies have traditionally considered felsic gneiss complexes of the Canadian Shield sterile. This paradigm is challenged with the discovery of extensive granulite-facies Cu-Au, Fe oxides-bearing hydrothermal systems in the 1.4 Ga Bondy and 1.5 Ga La Romaine volcano-plutonic complexes of the Grenville Province. Early identification of magnetite-rich units in association with meta-exhalite and pelite-like units was key to targeting the search for mineralization, protolith information and indicators of pre-metamorphic hydrothermal activity during regional surveys of these frontier terrains. Textural evidences for a volcano-plutonic setting were found. There remained, however, numerous impediments to the recognition that the ubiquitous magnetite-laden units and associated aluminous gneiss were hydrothermal in origin. Many metamorphic reactions form magnetite at the amphibolite to granulite facies transition, and hydrothermally-leached rocks resemble metapelite or migmatitic restite, while carbonate alteration zones may look like marble or calc-silicate rock. Non-typical metamorphic mineral assemblages, modes and colours arise when hydrothermally-derived rock types are metamorphosed. These effective markers served as tools for reliable field identification of hydrothermally-altered rocks. In our case-studies, we traced pre-metamorphic hydrothermally-driven Fe-oxide precipitation in association with sericitization, albitization, pyrophyllite alteration, chloritization, and B enrichment by mapping unusually high contents of (1) magnetite or hematite, (2) biotite, (3) albite, (4) sillimanite, (5) orthopyroxene or phlogopite with cordierite, and (6) tourmaline or kornerupine. The unveiled garnetite, tourmalinite, and aluminous gneiss had assemblages that vary according to metamorphic grade and composition. To assess if high mode of Fe-Mg-Al phases resulted from partial melting (restitic residue) or metamorphism of hydrothermal rocks, we compared leucosome mineral assemblages, shape and amount with that expected of in situ partial melting reactions. In both areas there is no compelling evidence for these units to be restites. In contrast, the presence of lapilli in garnet-cordierite-sillimanite gneiss, though extremely rare, constitutes conclusive evidence for extreme pre-metamorphic high-sulphidation alteration of a volcanic protolith, while Fe-oxides or sillimanite-bearing nodules and veins record a less pervasive alteration. Geochemical research on element mobility and mass balance calculations further testified of an hydrothermal origin for these gneisses. The magnetite-rich units, together with calc-silicate rocks, garnetite, hornblendite, albitite among amphibolite, provided effective guides to sites of fluid discharge and led to the discovery of Cu and Cu-Au showings. With the presence of highsulphidation alteration, these features suggest early Cu-Au VHMS-type alteration halos within a predominantly felsic composite volcano-plutonic arc setting for both case-studies. Evidences for subsequent F-driven orthomagmatic fluid activity are also present.

May 1, 2003

By C. A. Biley

"A dedicated iron removal stage was installed at Anglo American Platinum’s Rustenburg Base Metals Refiners (RBMR) during an expansion project which began execution in 2010. The Pressure Iron Removal (PIR) process aims to precipitate iron from a nickel leach solution under medium temperature (145 °C) oxidative conditions using caustic for neutralisation. Since commissioning, the process has been robust in removing iron from the refinery but has shown significant variability in the residues produced. These residues, which are typically comprised of ferrihydrite, hematite or natrojarosite, have varying iron contents and different physical properties which have implications for their filtration and ultimate removal from the refinery. Since over 95% of nickel plated at RBMR passes through the PIR section, volumetric constraints caused by poor filtration performance directly influence nickel production and stabilisation of this process remains central to the continued ramp-up of the refinery to full nameplate capacity. This paper details the development of the PIR process and the journey from concept to implementation. Recent laboratory and plant scale tests have highlighted the importance of the initial phase formed in the autoclave for the ultimate residue requiring filtration and how the feed acid concentration is a critical control parameter. On this basis, a new caustic dosing philosophy was implemented which improved pH control in the autoclave. Further optimisation of the PIR process is on-going.INTRODUCTIONIn 2010, an expansion of Rustenburg Base Metal Refiners (RBMR) to increase nickel production from 21 to 33 ktpa began execution. The management of iron was pivotal in the ultimate configuration in the new circuit and one of the primary differences from the original circuit was the addition of a dedicated iron removal stage: Pressure Iron Removal (PIR). A significant drive for this new unit operation was to facilitate the separation of iron from other leach residues. Not only did this result in a significant decrease in the amount of leach residue but, without the constraints associated with iron precipitation during leaching, has permitted more aggressive leaching conditions and, correspondingly, greater throughput and efficiencies.However, as with many commercial implementations of even mature technologies, the journey from development to implementation has presented challenges for the continued ramp-up of the refinery to full nameplate capacity. Specifically, since all nickel bearing leach solution must pass through the PIR section prior to the nickel purification and recovery circuit, throughput difficulties in the PIR section can directly impact supply of feed to the nickel tank house. Since commissioning, the PIR autoclave has effectively removed iron from the refinery, but has shown significant variation in the solids phases produced in the autoclave. This variation presents difficulties for the filtration of the discharge slurry and represents a key throughput limitation which periodically impacts production. Additionally, the increasing production rates associated with the ramp-up of the refinery have resulted in a significant increase in the required volumetric throughput through the PIR process. This has placed even greater emphasis on improving the stability of the solid phase produced in the autoclave, ultimately targeting an easily filterable residue with high iron content."

Jan 1, 2016

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

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

Jan 1, 2019

By N. P. Hristov, L. V. Kuzev

"A simulation study of the fill factor of a drum mill was conducted using two types of grinding media: balls and tetraballpebs. Each medium was composed of equal number of grinding bodies having equal weight but different shape. The difference was in the density of the two grinding media.Standard spherical grinding bodies (balls), which were used in practice, and equivalent (by weight) grinding bodies in the shape of spheroidal tetrahedrons (tetraballpebs) were studied. The simulation was conducted in a module of a 1.2 m diameter drum mill, whose length exceeded twice the diameter of the corresponding grinding body. The grinding balls had a few percent (3 to 6) higher fill factor than the grinding tetraballpebs in the same shell.Vertical sections of the grinding medium were taken across the studied shell module along its axis. They showed the location of each grinding body in the fill volume. Defects in the packing of the grinding bodies and the contacts between them and with the inner surface of the shell were also show. Differences in the contact points of the balls and the tetraballpebs were noted. For the tetraballpebs, we observed contacts both between the spherical surfaces and with the edges and vertices of the grinding bodies.The topography of the grinding media is shown visually.INTRODUCTIONThe grinding balls have been studied extensively. It is the traditional grinding medium in ore processing industry. The parameter characterising the grinding medium is its density. It is the mass of the grinding medium per unit of volume. The parameter may have a range of values. It depends on:• The physical properties of the material they are made of;• The shape of the grinding body;• The dimensions of the body;• Size distribution of the grinding bodies in the grinding media.In comminution, the fill factor of the mill shell is the process parameter that is representative of the grinding medium. Some publications have indicated the disadvantages of ball grinding. For example, Wu, Xiang, Glen, and Duan (2008) note in the analysis of the mechanism of grinding that the balls are not the best grinding media and disadvantages include among others:• Easy over-grinding resulting from the point contact between the balls;• Attrition surface is the smallest in the grinding media;• Poor selective exposure of minerals resulting from the accidental and often significantly more powerful impact on the feed;• The grinding efficiency drops as the fineness of grind is increased;In the Handbook of mineral processing, Polkin (1950) noted that the shape of the grinding bodies determines the volume of the void spaces therein and therefore the shape affects the grinding efficiency per unit of volume of the grinding medium."

Jan 1, 2016

By Q. Shi, Q. Feng

"The influence mechanism of chlorite on the flotation of pyrrhotite was investigated through flotation tests, Zeta potential measurements, optical microscope photos, solution ion measurements and absorption measurements. It is observed that, the addition of chlorite influences the flotation of pyrrhotite seriously. The IEP (isoelectric point) of pyrrhotite and chlorite is 7.6 and 4.3 respectively, in the condition of slight alkalinity, both pyrrhotite and chlorite are negatively charged, and the hetero-aggregation will not occur. With the existence of oxygen in solution, the iron ions are dissolved from pyrrhotite surface because of the oxidation of pyrrhotite during pulp conditioning. The dissolved iron ions adsorb onto the surface of chlorite in the forms of iron hydroxyl and iron hydroxide, and the adsorption results in the opposite charge of chlorite surface. As a result, the hetero-aggregation between pyrrhotite and chlorite occurs. The adsorption of chlorite changes the surface property of pyrrhotite and restricts the adsorption of amyl xanthate onto pyrrhotite surface, and then decreases the flotability of pyrrhotite.INTRODUCTIONIn sulfide ores processing, magnesium silicate bearing gangue minerals often report to the concentrate causing downstream processing problems as well as increased smelting costs. Talc,chlorite and serpentine are gangue minerals commonly encountered in complex sulfide ores (Engel, Middlebrook, & Jameson, 1997; Pietrobon et al., 1997; Witney & Yan, 1997; Chen, Grano, Sobieraj, & Ralston, 1999a, 1999b; Bremmell, Fornasiero & Ralston, 2005; Senior & Thomas, 2005; Wang, Somasundaran, & Nagaraj, 2005; WieseHarris & Bradshaw, 2007). All of them are characterized with uniform crystal structures consisting of silicon-oxygen tetrahedron and magnesium-oxygen octahedron. The magnesium-oxygen octahedron layers are broken during the liberation of serpentine to expose the magnesium-oxygen bonds with strong ionicity and hydration, which results in natural hydrophilicity of serpentine. In water, the hydroxyl on serpentine surface ruptures and dissolves into the aqueous phase while the Mg2+ remains, resulting in surface positive charge (Tao, 2012). As serpentine has a positive potential at alkaline condition, it is therefore likely that it will attach through electrostatic attraction to the negatively charged mineral surface and reduce both the flotation rate and recovery of target minerals (Lu et al., 2011). For talc, characterized as good natural hydrophobicity, the liberation comes from the interlaminar fracture to remain molecular bonds on surface with weak hydration. Being naturally hydrophobic, talc easily reports to flotation concentrates, thus reducing concentrate grade (Shortridge, Harris, Bradshaw, & Koopal, 2000; Beattie, Huynh, Kaggwa, & Ralston, 2006). Meanwhile, both molecular bonds and magnesium-oxygen bonds are exposed on chlorite surface when liberation occurs, resulting in moderate natural hydrophobicity between serpentine and talc (Tao, 2012). And thus the mechanisms of influence of chlorite on sulfide ores are more difficult to study."

Jan 1, 2016

By Francisco Tamargo

"This paper refers to a new process for the full recovery (above 98%) of zinc and other valuable metals (Pb, Ag, Au) accompanying zinc in its concentrates while iron is separated as a marketable product free of any harmful impurities and suitable for sale to the cement industry. The process has been especially designed to achieve the following goals: maximizing the recovery of zinc and those valuable metals that are present in most zinc concentrates (lead, silver, gold); allowing the recovery of other valuable metals such as indium or germanium that could be present in some zinc concentrates; avoiding or minimizing the need for residue disposal in safe areas as it is the common practice nowadays in the zinc industry. Finally, the process achieves the previously related goals at a competitive cost, in line with the existing more efficient operations while using conventional equipment working under atmospheric conditions. The new process can be easily implemented in those smelters using the jarosite or goethite processes as well as in those smelters using direct leaching as a part of the processing facilities. A study has already been conducted on the use of jarosite in the cement manufacturing process, in the raw mix preparation as well as for a set controller. It shows that use of clean jarosite in the manufacture of ordinary Portland cement can result in improved quality of the clinker. Jarosite showed compatibility to influence the burning characteristics of the cement raw mixes. The presence of SO3 bearing mineral phases such as natrojarosite showed the suitability of clean jarosite for use as a set controller in cement production.INTRODUCTION Iron is by far the most abundant impurity accompanying zinc in its concentrates, and consequently, iron residue disposal is a major environmental concern for many zinc smelters in the world. This is especially true during the last 30 years when environmental restrictions became more and more strict. Much research work has been carried out during the last 30 years to try to find a solution for this important issue, but with little success. Only the jarofix process represents a partial solution to this problem as jarofix is stored nearby the smelters with certain limitations as a non-contaminant residue. However, it does not avoid the need to dispose of an important amount of residues every year, representing an environmental liability. Meanwhile, the number of countries limiting or banning the disposal of iron residues generated by the zinc smelters is growing, so that the solution to this problem is becoming an urgent matter right now. At the same time, the lack of appropriate available land for iron disposal is also a major concern in many cases."

Jan 1, 2016

By G. J. Kipouros, N. Dadvand

"In recent years coatings of alloys have been produced for a wide range of applications including many for wire bonding application. The main characteristic of the baths of electroless plating is the presence of sodium hypophosphite as a reducing agent and also as the supplier of phosphorus. In the electroplating of alloys the Watts bath is predominantly used. This presentation compares the two techniques, electroless plating and electroplating, and discusses the possibilities of cross utilization of sodium hypophosphite in an attempt to formulate baths that eliminate the waste disposal issue related to some components.INTRODUCTIONThe importance of surface characteristics and the linkage of surface coating process to surface characterization has given rise to the evolution of surface engineering. Not only the surface preparation and morphology has become very important to the coating technology but also the knowledge of surface area exposed to the process of coating or corrosion is absolutely essential in ascertaining the behaviour of coatings and their adherence to the substrate. The above concepts are not unknown to researchers involved in the protection of surfaces, particularly metallic, from corrosion and/or erosion corrosion. However, the complexity of the surface defects, roughness, porosity, interconnected porosity and other related factors have contributed to the uncertainty of the true surface area versus the nominal surface area. Therefore critical parameters such as current density and chemical surface intensity although have been studied extensively at the atomistic level by carefully prepared surfaces are very difficult to be determined at the practical scale required during the coating processes.The most pronounced case calling for an understanding the relationship between the surface area and the effective protection is the coating of surfaces of items produced by compression or consolidation of powdery materials. With the advent of nanomaterials and the additive manufacturing through 3D metal printing this issue is becoming more important. Certain experience in this field come from the accumulated knowledge resulted from the attempts to characterize the electrochemical behaviour of powder metallurgy metals and alloys. Powder metallurgy has been a well-established process for half a century on basic metals such as iron, copper and their alloys. Recently the development of powder metallurgy process for aluminum, titanium and attempts to develop powder metallurgy processes for magnesium and its alloys has brought to light another important issue that of the development of coatings on difficult to plate reactive metals such as aluminum and magnesium. Despite of many successes in the development of powder metallurgy alloys the exact characterization of the corrosion behaviour of the coatings and the substrates is difficult to determine because of the unknown morphology of the exposed area i.e. the exact morphology and area of the exposed surface (Samal, 2007)."

Jan 1, 2016

By T. I. . I. Sibbald, D. E. Jiricka

The Goldfields area on the north shore of Lake Athabasca , was active in the thirties and early fortie . In 1934, Tom Box discovered gold mineralization in quartz veins in pyritized 'granite' at the ite of what wa to become the Box Mine of the Consolidated Mining and Smelting Company of Canada Ltd . In the rush of activity that followed numerou other gold occurrences were found. the most significant of the e being the Athona and Frontier prospects which underwent underground development bur did not reach production. Probable reserve identified to date in the Box and Athona deposits amount to some 3.5 million ton grading an average 0.ITT oz gold/ton . Some 10 to 15 million tons of lower grade, po ible reserve are also reported at the Box. From 1939 to 1942, the mine produced, from underground , 64,066 ounces of gold from ore grading 0.0452 oz gold/ton at a cost of between $1.5 and $2.0 a ton. The Goldfield area is underlain by upracrustal rock intruded by mafic bodies and invaded by ·granite'. The upracru tal . thought to be of Lower Proterozoic age are characterized by a helf-type ucce ion metamorpho ed under lowermiddle amphibolire facie conditions into quartzites, quartz-mica chi ts. dolomitic marbles, diop idic calc-silicates and cordierite-anthophyllite-biotite schi t . Amphibolites of presumed igneou origin form concordant to subconcordant units up to a few ten of metre thick in the upracrustals and occur as le regular masses in the 'g ranite·. Compositionally they vary from gabbroic to granodioritic. with the more mafic varieties being most common. Granitic rocks exhibiting a variety of textural , structural and compositional characteristics underlie most of the area. Four main units are recognized: foliated (older) 'granite', 'leucogranite', feldspar augen 'granite· and pyritic mine 'granite', however, in geological mapping the e are somewhat arbitrarily differentiated due to transitional relationships and admixture of types. Although there is no unambiguous age data available for the e 'granite ' adjacent northshore 'granite' of geologically similar age yield Rb-Sr isochron dates around 2000 Ma. The area is characterized by two major deformation events of which the earliest gave rise to the dominant regional foliation and to rare minor and outcrop scale folds. This foliation which in general parallels compositional layering is folded into the major Goldfields syncline, whose axis plunges weakly to moderately outh. Related open minor folds have subvertical axial surfaces which are locally paralleled by a weakly defined schistosity. Most rock display a pronounced rodding fabric, coaxial with second fold axes and apparently the product of this deformation event. However, a note of caution is introduced into this interpretation, as the rodding fabric is also parallel to first fold axe .

Jan 1, 1986

By Richard R. Klimpel, Simon Isherwood

"The industrial processing of ores containing gold and/or platinum minerals is generally mining site specific, e.g. McQuiston and Shoemaker (1985). The variety of various mineral processing techniques employed is extensive and includes size classification, gravity concentration, amalgamation, roasting, cyanidation, chemical leaching, ion ex-change/solvent extraction and froth flotation. Because of the relatively small quantities of Au and Pt that exist in the earth’ s crust, the economical recovery of Pt and Au often requires that the Au and Pt be associated with other industrially important minerals. This association of Au and Pt with other minerals can occur in very complicated mineralogical forms. It is probably safe to say that Au and Pt mineral concentration from natural ores represents some of the most complex metallurgical challenges facing the mining industry.Froth flotation is a technique that is receiving increasing industrial attention for the treatment of finely disseminated and/or associated Au and Pt minerals. The high throughput at relatively low capital cost, the potentially high selectivity, and the existence of a well-established technology for treating more abundant associated minerals such as base metal sulfides makes froth flotation an attractive processing alternative. Broadly speaking, the industrial oriented froth flotation experience with Au and Pt minerals exists in two major categories: 1) the flotation of Au and Pt associated with pyrite (or pyrrhotite) and 2) the flotation of Au and Pt associated with primarily Cu, Zn, or Cu/Pb/Zn containing sulfide ores.In the case of the precious metal association with pyrite (or pyrrhotite), much of the published froth flotation experience comes from South Africa. Typically, such processes are done using an acid side flotation with a collector such as mercaptobenzothiazole. The approach obviously is to float pyrite and hence recover associated gold. Generally the higher the total sulfur recovery the higher the gold recovery. Similar experiences of Pt associated with pyrite are a lot less common mineralogically and hence less well documented industrially. For example, in Canada, Pt is often associated with pyrrhotite but is not really processed as a straight pyrrhotite recovery.For the situation of Au and/or Pt associated with other base metal sulfide minerals, the industrial experience base is quite varied and is well documented from both North and South American operations. The nature of most of these flotation processes is to design, operate, and optimize the installed concentrator to get the Cu, Pb, Zn, Ni, etc. that is the bulk of the valuable sulfide ore and then take whatever precious metal recovery comes. Typically, such flotations are on the basic side using standard thiol collectors such as xanthates, dithiophosphates, thionocarbamates, etc. It is very common with such ores to use mixtures of collectors in order to get ""everything"". Often in such plants, there is an inherent balancing act occuring so as to operate in such a manner to reject pyrite (and pyrrhotite) in order to maintain high base metal mineral concentrate grades and yet recover sufficient precious metals such as Au and Pt minerals which often are associated with the iron sulfides being rejected. Obviously in such an environment, the smelter operation often ends up telling the concentrator what the precious metal recovery actually was in a given production period."

Jan 1, 1994

By R. J. McLeod

"A new backfill plant was constructed in 1966 adjacent to the mill of Giant Yellowknife Mines. The reasons for the addition were to increase the storage capacity for sand fill, provide facilities for mixing cement with the fill for mine stope floors and provide a system for pumping fill to B -3 raise, a horizontal distance of 4700 feet from the mill. At the same time it was hoped to increase the density of the fill pumped underground.The plant consists of two only steel storage tanks of approximately 400 tons capacity of sand fill each. They are 17 feet in diameter and have conical bottoms. The cylindrical section is 28 feet high and the conical section 14 feet high. Each tank is fitted with water nozzles around the conical sections to remove solids at the required density. The water pressure on the nozzles is 150 lbs per square inch. The cyclones used to produce the sand back fill are located on top of the storage tanks. Five model Dl0B Kreb cyclones are used for this purpose. The vortex finders are 2 1/ 4 inch and the apex 1 1/4 inch. The cement silo or storage tank is located in the same building. It is a steel tank 17 feet diameter by 41 feet high with a conical bottom. The silo holds 250 tons of bulk cement, which is unloaded by air from a bulk highway carrier using 12 to 15 lbs air pressure. The cement is conveyed to the top of the silo, a distance of 70 feet from the ground by 450 lbs. of air.The mixing tank is a rubber lined steel tank 14 feet in diameter by 14 feet high. It is equipped with a Lightning Mixer supplied by Greey Mixing Equipment Ltd. The mechanism is driven by a 40 HP motor. The lower shaft is unsupported and is equipped with two 36 inch diameter axial flow turbine type rubber covered impellers. The backfill is discharged from the storage tank into the mixing tank by a 5"" x 4"" SRL-C pump powered by a 20 HP motor. A level control in the mixing tank regulates the rate of discharge of this pump through a Grinnell-Saunders automatic control valve. The cement silo is situated at such a height above the mixing tank to allow for a Rota-Lock, a small hopper and a cement feeder. The feeder is an Omega gravimetric belt type dry feeder with a Sensogram controller. The capacity of the cement feeder is 20 tons of cement per hour. The cement and backfill are mixed in the mixing tank at whatever ratio is required for the mine stope floors."

Jan 1, 1968