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By J. Lu, D. B. Dreisinger, G. Houlachi

"Zinc metal is primarily produced by electrowinning from sulfate solution. Zinc electrowinning is very sensitive to impurities such as Co and Ni. These impurities have to be removed to produce high-quality zinc and maintain a high zinc current efficiency. Zinc dusts are used to remove these impurities by cementation. However cobalt cementation is extremely slow and it is typically activated by the addition of Cu and Sb. Even with the use of Cu-Sb activation, cobalt is still difficult to remove. The zinc dust used for cobalt cementation is negligible compared to that consumed for hydrogen evolution. In some plants up to 6-7% of electrolytic zinc is used for cementation. This is a great waste of energy. Cu-Sb activation is not well understood. The objective of this study was to further understand the Cu-Sb activation and then find a way to reduce hydrogen evolution and promote cobalt cementation. The investigation reveals how hydrogen evolution (or zinc dissolution) and cobalt cementation and re-dissolution change with temperature, zinc dust dosage, the concentrations of copper and antimony and time. In the Cu-Sb activation, copper is deposited on a zinc particle to produce a porous shell. The zinc particle as an anode dissolves while the shell as a cathode provides a place where cobalt cementation and hydrogen evolution occur and they are catalyzed by antimony. A certain level of antimony (around 0.1 mg/L) is a key to maintain the catalytic effect of the shell. At a higher antimony concentration, cobalt cementation is accelerated slightly while hydrogen evolution is increased greatly. When a zinc particle dissolves to a certain degree, cobalt loses its cathodic protection and starts to re-dissolve. The change in cobalt concentration is determined by the rates of cobalt cementation and re-dissolution. The operation conditions can be optimized to maximize the cobalt removal and minimize the zinc dust dosage.INTRODUCTIONMost of the world’s zinc metal is produced by electrowinning from sulfate media. Many impurities such as iron, copper, cobalt and nickel are co-extracted into solution with zinc. Iron, mercury, indium, lead and tin are removed by co-precipitation as hydroxides. Copper, nickel, cobalt, aluminum, arsenic and germanium are partially co-precipitated with iron while cadmium, magnesium and chloride remain in solution. Cobalt, copper, cadmium, nickel and etc. are subsequently removed by cementation with zinc dust to achieve an acceptable current efficiency and high purity metal product for the economical production of zinc. Zinc provides a strong thermodynamic driving force for the cobalt cementation:"

Jan 1, 2016

By Daniel Hodouin, Bruno Girard, Claude Bazin

"Mineral processing and smelting operations have the reputation of being complex processes. Process operators of these unit operations have acquired from experience and basic training a superficial understanding of the relationships between process output and manipulated variables. This understanding, which is a very valuable know-how for process control and optimization as well as for operator training, is however difficult to extract. This paper describes a time series based approach that can be used to extract this know-how, which can subsequently be used for operator training. The method is illustrated with an example taken from a smelting process.IntroductionThe emergence of Total Quality Management and Statistical Process Control philosophies in the North American industry (Brown, 1991 ), as well as the increasing popularity of the IS0- 900x standard accreditation, provide a strong incentive to review the training procedures for plant operators. The importance of employees training is acknowledged in two of the 14 points of Deming's road to quality (Deming, 1982) as well as by control engineering consultants (Hendricks et al., 1994). However guidelines for the operation of industrial processes are often transferred from an operator to another one with allowance for flexible interpretation of many operating rules. Such transfer is typical of complex industrial processes to which particulate processes of ore dressing and extractive metallurgy belong.Capturing the knowledge of experienced operators is often a difficult task. Cognition engineers are able to extract this knowledge and transfer it into rules that ultimately are implemented into expert systems. Cognition engineers are however not always available and plant supervisors may prefer to keep training as an internal liability. Another difficulty through interviewing of the process operator is its lack of willingness to transfer what he has learned by hard working to other people, whose ultimate goals are to develop computerized applications that may replace him in a near future, or reduce its importance in the production line.This paper shows that time series can be advantageously used to capture the average process understanding of plant operators and to help reviewing training procedures. The method s illustrated with data collected during normal operation of electric arc furnaces and it is shown how the approach has led to a fruitful discussion between process engineers and operators that ultimately provided savings of more that $200,000 per year per furnace. There are 9 furnaces in the studied foundry. Such savings are not the results of application of advanced control strategies nor of ""miracle press button software», but of an improve awareness by engineers and operators of the importance of properly defining the operating procedures to control a process."

Jan 1, 1996

By P. Hansman

"Place Dome Canada's Detour Lake Mine is a low-grade gold producer located in northeastern Ontario. Production is targeted for 2850 tonnes of ore per day. As reserves and mining areas continued to be found below and far to the west of the shaft (Fig. 1), it became necessary to review and upgrade the materials handling facilities. This paper deals with the selection, design and construction of an underground crusher and conveyor system capable of handling up to 450 tonnes per hour from the lowest existing level in the mine.BackgroundThe Detour Lake mine site is accessed via secondary Highway 652 which runs northeast out of Cochrane, Ontario. The road consists of 200 Ian of hard top followed by 100 Ian of well maintained gravel to the mine site. A 1500 m airstrip is maintained at the mine site for emergency use. Prior to 1991, the workforce was flown to and from the mine site. Crews were now bused to the site from Timmins. Remnants of a winter road exist to the southeast which at one time ended in La Sarre, Quebec. This was the route used to haul bulk items during the construction period and early production years.A town site has not been established at the site for the current operations. Instead, a rotational scheme is being utilized whereby most personnel will work seven days on site followed by seven days off site. For some employees, however, the work week is termed as a four/three schedule where they work 42 hours between Monday at 10:00 a.m. and Thursday at 4:00 p.m. with the following three days off. All surface employees and maintenance employees underground work 12-hour shifts. All underground employees work a lOV2-hour shift. Supervisors work a similar schedule as their crews, however, other staff may overlap their rotation during the week.A five-compartment shaft was sunk in 1984 from surface to the 610 m elevation. Shaft stations were established at the 120 m, 260 m, 360 m, 460 m, and 560 m levels. The cage compartment consists of a single deck, 4O-man cage in balance with a counterweight. The cage hoist is an Ingersoll-Rand 10H and is regulated to travel at 4.1 m/s. Two 9 t skips operate in balance at a travel speed of 7.1 mls. The skip hoist is a 3.05 m drum Nordberg. Both hoists are automated to function without a hoistman during normal skipping and cage service runs."

Jan 1, 1995

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

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

Jan 1, 2019

By V. Panayotov, M. Panayotova

"Finding a reliable relationship between fundamental properties of minerals of non-ferrous metals and their ability to float is extremely important. The approach presented in this work allows us to draw predictive models for an effective flotation, suitable choice of reagents regime and expected technological results before the minerals enter the flotation process. Attempts to adjust the technological regime in already begun flotation are often unsuccessful because the basic physicochemical processes have already taken place. Mineral processing can be optimized on the basis of objective and basic parameters, decisive for the efficiency of the process - the state of the surface of the mineral; assessment of: minerals flotation ability, the need to apply and the type of needed pretreatment, and expected technological results. The theoretical foundations of such an approach are presented in the work, as well as supporting them experimental results, and information on practical implementation in lead-zinc and copper-pyrite ores selection.INTRODUCTIONIncreasing impoverishment of ores of nonferrous metals incoming to the flotation and rising costs for obtaining the final product lead to the inevitable question - will the mineral processing process be beneficial economically, technologically feasible and environmentally safe and what is the way to improve its efficiency. Some of the possible solutions are already in practice - new more efficient and less energy consuming flotation machines and equipment, new - more efficient and selective reagents, new combinations of flotation regimens, and so on.Finding a reliable relationship between fundamental properties of minerals of non-ferrous metals (sulphides and oxides - most of them with semiconductor properties) and their ability to float is extremely important. This is the subject of research, started in the 60s of the previous century and continuing to the present day. Such dependence would enable a realistic assessment of flotation ability (flotability) of minerals before their flotation processing. The basic dependence, displayed by the researchers as Hoberg (1975), Chanturia and Shafeev (1977), Chanturia and Lunin (1983), etc., represents a relationship between bulk properties of minerals and their flotability, considering especially the concentration of the main charge carriers, namely electrons or holes. The bulk properties are practically constant for minerals from a given field. But the 'main arena' of the flotation process is actually the mineral surface (Carta, 1970; Rao, Labonte, & Finch, 1992; Panayotov, 1999; Rao, 2003; Hu, Sun & Wang, 2010). Therefore, we introduce a methodology for studying the surface of minerals, which methodology gives the degree of semiconductor bands bending on the surface and respectively - the charge of the mineral surface and its ability to interact with reagent-collector. The latter can be used to select the most optimal reagent for the corresponding mineral."

Jan 1, 2016

"In conjunction with Banick Mercur , Wright Engineers Limited designed the first multi-stage aqueous pressure oxidation system in an alkaline environment in North America. A three-stage slurry heating and slurry pressure let-down with heat recovery was an innovative approach to the problem of high energy cost with the low sulphur content of the ore. The design of this system required solving a number of problems associated with heating and pumping of slurries at pressures up to 3230 kPa and temperatures in excess of 220°C. Wright received an Award of Merit in October 1989, from the Canadian Consulting Engineering Awards Committee in the Process Category for its design of the Mercur pressure oxidation system.The Mercur Mine is located at the south end of the Oquirrh Mountains in Tooele County about 65 miles by road from Salt Lake City, Utah, U.S.A. The original mill was designed to process 5000 tpd of oxidized gold ore in a carbon-in-leach (CIL) circuit. A new ore treatment system was required to treat the substantial quantities of refractory ore reserves which could not be treated economically in the existing mill. The new pressure oxidation circuit makes the refractory ore amenable to gold recovery by oxidation of the sulphides contained in the ore .Wright Engineers Limited was commissioned by Barrick Mercur Gold Mine Ltd ., a wholly-owned subsidiary of American Barrick Resources, to design the pressure oxidation facility. This followed completion of a feasibility study by Wright which showed that the project was technically and economically viable.Work on the feasibility study began in 1986, detail design commenced in February 1987, and was completed in December 1987. Mechanical installation was completed in January 1988, and slurry feed began in February 1988. The plant has operated since June 1988 without major interruption . Wright was responsible for engineering, procurement and construction management for the addition of the 750 tpd pressure oxidation circuit to the existing mill and also had commissioning responsibility.Major unit processes involved in the pressure oxidation facility are ore thickening, three-stage slurry preheating, pressure oxidation and three-stage slurry cooling with heat recovery, stearn generation, propane and oxygen storage and a video graphic display process control system were included in the scope as process anciliaries.The pressure oxidation increases gold recovery from the refractory ore by conventional CIL from as low as 35% to more than 86070. A patent has been granted for the process."

Jan 1, 1989

By A. Nielsen

"The Polaris mining operation (Fig. I) is located on Little Cornwallis Island in the Canadian Arctic, an area with a very limited shipping season and a hostile climate. The operation consists of the underground works, barge-mounted concentrator facilities, concentrate storage and shipping facilities, a personnel accommodation complex and an airstrip.The electrical energy for the operation is provided by four 2.3 MW diesel-generators installed on the barge. The power is distributed through a 4.16 KV system connected directly to the generators and to the large motors. All other loads are supplied at 600 V or 120 V.The power plant has been designed as a Total Energy System with maximum energy efficiency and thus designed for maximum heat recovery. The waste heat recovered on a year 'round basis is expected to be sufficient for the process heating and drying as well as the space heating for offices and accommodations. The engine cooling system consists of two glycol-water loops, a "" cold"" loop for charge air and lube oil cooling and a ""warm"" loop for jacket water cooling. Each loop is complete with heat exchangers f or heat recovery and out-door forced air radiators to dispose of excess heat. A water loop in the concentrator area transfers the heat from the jacket water loop and the exhaust gas boiler to the space heating equipment, the process and the heating system for the accommodations. Part of the exhaust gas is diverted to the concentrate dryers.IntroductionThe Polaris Mine is a highly mechanized, trackless mine with access through a decline. Drill equipment is electrically powered to minimize fumes and heat. Dry development and production drilling eliminates the need for an underground water supply system.The concentrator, power generators, work shops and auxiliary facilities are located on a barge which is beached in a prepared berth on the shore of Crozier Strait. Oil storage tanks, concentrate storage and accommodation buildings are built on shore. The concentrate storage building ha s a 12-month capacity and is fed from the concentrator by a conveyor system.The accommodation building houses up to 240 people, includes administration office, dining and recreation facilities. The complex is housed in on e interconnected building with public use areas alternating with accommodation units.At full production annual concentrate shipments will amount to about 220,000 wet tonnes. This will be accommodated by about 8 shipments during the regular shipping period-August 15 to October 15. Re-supply of general cargo and fuel will be carried out annually in about 3 shipments within the same period."

Jan 1, 1984

By Brian Hall

For some years, the mining industry has been consistently delivering returns below the market average. One of the causes is a disconnect between what is perceived to drive value creation by many industry analysts and senior corporate executives, and what actually drives the intrinsic value of mining operations. The perverse effect is that strategies that should increase value are perceived to have the potential to drive share prices down, and vice versa. This paper challenges common perceptions of what drives a mining company?s value. Anecdotal evidence suggests that the key drivers of value in the market, and hence in many boardrooms, are such things as increasing the reserves and production rate, and reducing unit operating cost. However, while the quoted reserves may satisfy ore reserves reporting code requirements to be economic, this does not imply that they are optimal. Often, a proportion of the quoted reserve reduces the potential value of the operation. Metal price increases, increased production rates, and operating cost reductions are all seen as conditions or actions that can improve value, and lead to cutoff grade reductions, thereby increasing ore reserves. However, modern cutoff theory and studies conducted by the author and his colleagues show clearly that, while breakeven cutoff grades will fall in these circumstances, optimum cutoffs may be much less volatile, and may even increase. Results from case studies indicate that value is maximised by right-sizing, not maximising, the production capacities, and by optimising the cutoff strategy. Values of many existing underground mining operations can be increased significantly by a substantial cutoff grade increase. Also, an increase in downside risk, and hence reduced returns, can occur using typical life-of-mine planning strategies if prices received are lower than predicted. Since this is often the case, low industry returns may be a direct result of typical strategic mine planning processes. Relatively simple cost-effective techniques are available to provide senior decision makers with information needed to assess the tradeoffs between their many conflicting corporate goals, and the balance between risk and reward. Often this information is not being generated at all. If it is, decisions affecting the value of the company are occurring by default at relatively low levels within the hierarchy, with senior executives apparently unaware of them or their potential impact. The paper concludes that without a re-evaluation of the way mining projects are valued, to demand proven value optimisation from mining plans by boards and mining analysts, the industry will continue to deliver below average returns.

May 1, 2003

By N. Sarmady, Mei Li, N. C. Parson, W. J. Poole

"The mechanical behaviour of AA6082 is a function of the extrusion conditions and in particular, the quench rate after extrusion. It is well known that precipitation of Mg-Si phases on grain boundaries during quenching can negatively affect the ductility of the alloy. In this study, AA6082 alloys containing 0.5 wt. % Mn and 0.15 wt. % Cr were direct chill (DC) cast, homogenized and then extruded to produce 3 mm x 42 mm strips. The microstructure in the as-extruded strips was unrecrystallized due to the Zener-Smith drag from the Mn/Cr dispersoids. It was shown that a solution treatment of 5 minutes at 560°C did not significantly affecting the as-extruded microstructure but if the as-extruded strip was cold rolled prior to heating, recrystallization occurred concurrently with the solution treatment. As such, 3 initial microstructures were produced, i.e. unrecrystallized and recrystallized with a grain size of 9 and 40 µm. After the solution treatment, controlled cooling experiments were conducted using the Gleeble 3500 thermomechanical simulator with cooling rates of 10, 25, 80°C/s as well as water quenched. The mechanical properties (e.g. yield stress, fracture stress and fracture strain) were determined from tensile tests for the different combinations of initial microstructure and cooling rate. The relationship between the quench rate and precipitation of Mg-Si phases on grain boundaries was examined by FEG-SEM. It was found that the yield stress decreased as the cooling rate decreased and that there was complex relationship between the fracture properties and the microstructure.INTRODUCTION As the automotive industry strives to reduce fuel consumption and CO2 emissions with the goal of reducing the environmental impact, there is a need for lighter vehicles. As a result, aluminum alloys are gaining popularity to replace steel parts as structural components. An exemplar of this is the Ford F-150 pickup truck which has an aluminum body and is ˜750 pounds lighter than the previous steel version of the vehicle. The structure includes a mix of aluminum sheet, extrusions and cast alloys. The current research is concerned with the performance of aluminum extrusions in applications such as side rails, crash tubes and other structural parts. These components experience a combination of bending, stretching and folding both during the forming of the part and in service, e.g. crash of a vehicle. There is a need to have detailed information on the influence of processing on final material performance, in particular the cooling rate after extrusion."

Jan 1, 2018

By J. L. Jambor, E. J. Reardon, E. C. Appleyard, J. A. Cherry

Since the 1957 initiation of deposition of mill tailings at the Heath Steele mine, New Brunswick, periodic detailed studies of the tailings geochemistry and mineralogy have been conducted, thereby providing a unique opportunity to examine the geochemical evolution of a sulfide rich mine>tailings impoundment. When deposited, the tailings contained about 85 wt% sulfide minerals, 1 wt% to 3 wt% carbonate minerals, and 12 wt% to 14 wt% alumino silicate minerals. Mineralogical examination of samples of the tailings from deep in the impoundment indicates that little oxidation occurred prior to tailings deposition. The mill process-water deposited with the tailings was neutral to slightly basic and contained low concentrations of Cu, Pb, Zn and As. Sampling of the tailings five years after deposition indicated that sulfide-oxidation reactions, occurring near the tailings surface, had decreased the tailings-water pH and increased the concentrations of Fe, S042-, Cu, Zn and Pb (up to 10 000 mglL, 25 000 mg/L, 2.5 mg/L, 200 mg/L, and 4.5 mg/L, respectively). In 1975, ten years after decommissioning, the tailings could be divided into three zones: an upper oxidation zone in which sulfide oxidation is actively occurring, a hardpan layer, 15 cm to 20 cm thick, composed of tailings cemented by secondary Fe minerals, and an underlying reduction zone in which the tailings are relatively tin oxidized (Boorman and Watson, 1976). Chemical analyses indicated the presence of low-pH waters containing very high concentrations of SOi-, Fe, and other metals (up to 63 600 mg/L SO42-, 28 600 mg/L Fe, 10 mg/L Pb, 600 mg/L Cu and 4000 mg/L Zn), suggesting that there was a period of intense sulfide oxidation between five and ten years after tailings deposition. More recent sampling, conducted some 20 years after the end of tailings deposition, indicates that relatively little change has occurred in the last ten years, thereby suggesting that the hardpan layer has provided an impediment to the downward movement of dissolved constituents. Analyses of the tailings solids show increased accumulations of many metals at or near the depth of the hardpan layer. Evaluations of the long-term evolution of the tailings can be enhanced by comparison of descriptions of the tailings with descriptions of the gossan that overlay the massivesulfide orebody prior to mining. At Heath Steele, this comparison suggests that many of the mechanisms controlling the distribution of metals in the tailings will probably persist over geologic time.

Jan 1, 1992

By R. Saulters, M. Garshasb

The Transportable Moisture Limit (TML) is the copper concentrate moisture content beyond which liquefaction may occur during shipping. This study indicates that TML is a function of mineralogical characteristics of concentrate, which in turn may directly affect moisture retention properties. A comprehensive mineralogical and chemical analysis was conducted to determine the impact of composition and mineral content on concentrate’s TML. The analysis indicates that the high TML value is associated with particular minerals such as gypsum, swelling clay, muscovite and pyrite. Since higher TML values allow higher moisture contents for shipping, this property is desirable for concentrate inventory management and shipping flexibility. Scarcity of minerals that may increase TML contributes to limitations of shippable inventory. INTRODUCTION TML is the maximum moisture content of copper concentrate beyond which liquefaction may occur. Moisture content above the TML is considered a shipping hazard as it may result in loss of lives and cargo. The details of TML testing are given in ISO 12742:2007 (ISO 12742). In principle, TML is 90% of the Flow Moisture Point. The method of testing for Flow Moisture Point and TML is based on procedures that provide pulsations that simulate the cyclic force of ocean waves to determine moisture levels at which the onset of concentrate flow may occur. International Maritime Organization (IMO) provides guidelines for developing and approving TML testing and moisture determination in the context of its Moisture Management Plan (MMP) for shipment of copper concentrate cargoes (IMO 2013). The IMO regulations are specifically provided in MSC.1/Circ. 1454 “Guidelines for developing and approving procedures for sampling, testing and controlling the moisture content for solid bulk cargoes which may liquefy”. Since each port has its own operational characteristics, resource availability, peculiarities of governmental structure, and desired methods of interaction with the authorities, it is necessary to adopt the IMO guidelines to the local requirements. In order to establish a TML and MMP measurement system, it is required to implement and follow proper guidelines for the relevant procedures. For example, sampling procedures are developed in accordance of ISO 12743, “Copper, lead, zinc, and nickel concentrates – Sampling procedures for determination of metal and moisture content”. Moisture content of copper concentrate samples are determined in accordance with ISO 10251, “Copper, lead, zinc, and nickel concentrates – Determination of mass loss of bulk material on drying”. To assure strict adherence to the MMP procedures, the overall system needs to be integrated into the site QA/QC system, which would make them auditable and certified in accordance with standards such as ISO 9001, ISO 17025, and ISO 17020.

Jan 1, 2019

By T. Crivits, I. Bellemans, V. Cnockaert, K. Verbeken

It is well known that slag foaming can significantly complicate copper smelting processes. This is especially true for unpredictable events that generate large amounts of slag foam during a small period of time. These events are here referred to as eruptions and are typical for certain types of liquid bath furnaces such as modern top submerged lance (TSL) furnaces. A few authors already made the link between over- oxidation of the slag phase and slag foaming. In this work, the mechanism is studied in more detail and explained using the smelter chemistry and the behaviour of dynamic foam formation. Finally, the most influential parameters on the foaming behaviour of the process are presented. INTRODUCTION Copper remains one of the most important metals used in our modern society. In 2017, the annual amount of copper produced worldwide was about 23.5 million metric tons. This makes copper the third most produced metal in the world after steel and aluminium. The majority of this copper was produced by copper smelting, about 19 million metric tons, both from primary concentrate and secondary feed material as reported in ICSG’s World Copper Factbook 2018. In the past, copper smelting was mainly performed using Flash smelting, a technology developed before 1949 (Kojo, Jokilaakso & Hanniala, 2000). Nowadays, this type of smelter is being complemented by other smelter types such as ISASMELT and Ausmelt technologies. These are two types op smelter furnaces, based on the concept of a top submerged lance (TSL) furnace. These two technologies were both commercialized around 1990 (Arthur & Hunt, 2005). It is estimated by Wang et al. (2016) that at this time, both flash and TSL smelters contribute equally to the total amount of annually smelted copper. Flash smelting stays the most important technology for extracting copper from primary sources. TSL furnaces, on the other hand, are more suited than Flash smelters for the recovery of copper from secondary materials. This is because of the different ways that the feed material is inserted in both furnaces. In Flash smelters, the ore and fluxes needs to be fine and completely dry in order to be able to blow the mixture inside the burner area of the furnace. The feed material for TSL smelters show less constrains concerning humidity and particle size. The material is dropped from the top inside the molten bath. As a result, a more complex and flexible feed stream can be processed using TSL smelting (Wood, Hoang & Hughes, 2017). This feature will become more important in the future as primary copper sources gets more scarce and the need for a fully sustainable material cycle increases.

Jan 1, 2019

By R. Bruce Ferguson

"The Kettle River Operation of Echo Bay Minerals Company is located about thirty kilometers north of the town of Republic in northeastern Washington State. In tum, Republic is about a hundred and seventy kilometers north and a little west of the city of Spokane, Washington. It is also about fifty kilometers south of the town of Grand Forks, BC, Canada. At this time it is the only gold mining and milling operation in the State.Tailings are stored in a double lined, zero discharge facility. It became evident in mid 1996 that the rate of free water accumulation in the tailings pond was going to result in the level in the pond reaching its permitted freeboard elevation long before the solids plus pore water capacity was reached.We then embarked on a mission to manage tailings pond water on several fronts: reduce fresh water usage in the process, divert spring seeps and surface runoff and install a system to enhance evaporation.IntroductionThe Republic area is an historic gold mining district. Hecla operated its Republic unit from 1892 until closure in 1994. The Kettle River Operation processing plant is located in the San Poi! river valley at an elevation of just under a thousand meters. The facility consists of a ninety thousand ton ore storage pad as the mines are remote and truck haulage is subject to interruption due to weather and road restrictions, jaw and cone crushing with screening, rod and ball milling to an eighty percent passing size of seventy five microns. This is followed by preaeration with high purity oxygen supplied by a vacuum pressure swing adsorption plant plus lead nitrate addition to achieve passivation of the sulfide mineralization such as chalcopyrite, pyrite and pyrrhotite. Conventional cyanide leaching and carbon in pulp adsorption completes the loop. Tailings are treated with air/S02 to achieve cyanide destruction. Plant capacity is 1,815 tons per day. Our license requires that there be no detectable free cyanide in the tailings pond water and a maximum of forty parts per million weak acid dissociable cyanide. The operation started in 1989 with production from the Overlook Mine located about five kilometers further up the same valley at an elevation of twelve hundred meters, hence the location of the mill. Current mine production comes from Lamefoot in the Curlew Lake valley, about ten kilometers distant, and K2 in the Kettle River valley about fifty kilometers away. Both are at an elevation of about 800 meters. Lamefoot produces 1,360 tons per day and K2, 455. Both are underground operations."

Jan 1, 1998

By M. Romana

"RMR geomechanics classification is being used all over the World as a useful tool for tunnel predesign and construction. Bieniawski (1989) provided guidelines for tunnel excavation and support, according RMR numerical values; guidelines which are now included in almost all tunnelling books, regulations, and Internet portals. But in the last 25 years tunnelling techniques have changed and some of the guidelines have become obsolete.. The conditions of tunnels for the 1989 guidelines application were: horseshoe shape; 10 m width; vertical stress < 25MPa; construction by drilling and blasting. Actually a substantial number of tunnels are mechanically excavated. The tunnel themselves are now bigger: a span of 14 m. is very common for transportation tunnels. The guidelines included the following concepts: excavation method, rock bolts (20 mm diameter, fully grouted), shotcrete, steel sets. There is an increasing gap between the 1989 guidelines and the practice: in many tunnels different to-the-case excavation and support requirements are established. Bieniawski himself has published a paper (Lowson & Bieniawski, 2013) with different guidelines based in Lowson’s practice, with no explanations and/or justification.In this paper a new set of RMR tunnelling guidelines is proposed taking account of the actual industry conditions, and increased safety regulations. The 5 classes’ limit RMR:values (80,60,40,20) are no more useful actually: Values of RMR (70, 50, 30) are used in practice to define quite different tunnelling techniques. Consequently a 10 subclasses system is proposed. Each subclass covers a span of 10 RMR points. As the number of mechanically excavated tunnels has increased a new correction factor is necessary, according the excavation method.The proposed 2014 RMR guidelines are for tunnels 10 to 14 meters wide, the world practice for new roads/railways tunnels. Many tunnels for water/ services are smaller with smaller support needs. Wider than 16 m caverns are out of the RMR system.It is not possible, nor desirable, to propose only a value for support in each subclass. The new guidelines propose an interval of values for certain types of support. More pessimistic values must be chosen for wider tunnels, tunnels with lower RMR values in each class and cases when a higher safety factor is asked for. More optimist ones are chosen for narrower tunnels, and higher RMR values. Some alternatives are almost complementary, others are contradictory. Sometimes several different techniques are proposed Guidelines are not automatic options. Good engineering judgment must be used when choosing between the different alternatives. Guidelines are not mandatory suggestions."

Jan 1, 2015

By T. Patching

"MINING operations at the Star Key Mine provided an example of a small-scale prairie-type coal mine that employed several interesting procedures and achieved good productivity with low capital expenditures. The mine was located about 11 miles north of the city of Edmonton, on the north side of the Sturgeon River valley. It opened in 1945 and operated until 1979. Output in the winters reached about 500 tons a day in the late 1940s, but in the spring and summer periods activity was usually limited to only servicing and development work. In the later years of the mine's life sales fell to less than 100 tons per day. Most of its output was bought for domestic use in furnaces and stoves so production of large sizes was desired. A modest market for stoker coal did exist for a time. Screenings (-5 / 8 in.) were sold to the city's power plant until it converted to natural gas and were also used for thawing frozen ground for work on pipelines. These markets were comparatively small and the small sizes did not command as high a price as did the larger sizes. However the extension of natural gas services to rural markets cut into the domestic heating market and finally led to closure of the mine. The coal was strong when freshly mined, but because of its high inherent moisture content, about 25%, it did not store well on surface but tended to disintegrate. It was of subbituminous Crank with a heating value of about 8000 Btu per pound and burned with a clean flame. The seam mined was in the Edmonton Formation, of Upper Cretaceous age, at a depth of 60 to 80 ft below the surface. The seam in the mine area was about 11 feet in thickness but thinned gradually to the north and south and split to the west (1). It was approximately horizontal but gentle rolls in the floor of several feet in height necessitated brushing in places to produce suitable haulage grades. The floor material was of bone and clay, overlying a thin layer of coal. Above the seam there were only a few feet of mudstones and weak sandy shales while the remainder of the overburden was clay and unconsolidated glacial till and surface soil. Because of the soft floor and weak roof material the miners usually left about a foot of coal in the floor to provide a firm working surface and about a foot of coal above the rooms and two feet of coal above the entries to provide a competent roof."

Jan 1, 1985

By Chris D. Veldhoen

If you ask most diamond drillers if they have ever encountered methane gas, the answer would probably be yes. This paper explores a serious methane gas explosion that the author experienced while diamond drilling at a northerly Ontario mine. It is hoped that by documenting this event, people may be alerted to the hidden danger posed by confined methane in underground workings. Methane gas tends to be a problem the deeper you get in any Canadian gold mine. Common procedure when encountering methane is to take readings at the collar of the hole and at different lengths from the source. Monitoring continues until the readings drop below 0.25 ppm. Methane is a tasteless odorless gas that is usually locked up in the seams of the rock or is a product of decaying organic material. Occasionally, the gas is released abruptly over a short period of time. Sometimes the gas can continue to vent for months or even years. In small concentrations, methane is relatively safe. In concentrations under 0.25 ppm methane gas is considered safe to work around. Concentrations over 0.25 ppm can become dangerous. During the summer of 2001, Chris Veldhoen was drilling at a gold mine in Ontario. There were three electric hydraulic diamond drills on the 39 meter level, approximately 5200 feet deep. Two Longyear LM 55 drills were about 40 feet apart and another LM 75 was about 2000 feet away. The two LM 55?s were set up to drill holes no deeper than 350 meters. All three drills had been set up on the same spot for more than a month. None of the three drills had ever encountered methane gas in the entire time they had been drilling on the level. On the day shift of July 13 2001, the driller on Chris? cross shift hit a small pocket of gas. He tested the hole and got a reading above 0.25 ppm. He proceeded to rope off the area with a ?DO NOT ENTER - METHANE GAS? sign and a rope denying access beyond the sign. On Saturday, July 14, Chris was instructed to install a Margo plug in the hole that was emitting methane gas. Methane is usually odorless, but occasionally when encountered underground, is accompanied by a rotten egg smell. As he walked to the drill, he could not smell methane gas. He proceeded to test the hole and got a reading of 0.22 ppm methane. This number was in the acceptable range for drilling, but he plugged the hole anyway.

May 1, 2004

By J. D. Bateman, A. S. Dadson

"The Giant claims are on the west side of Yellowknife bay about 1 ½ miles north of the settlement of Yellowknife. They were staked in 1935 by prospectors in the employ of Burwash Yellowknife Gold Mines, Limited, and were transferred to Giant Yellowknife Gold Mines, Limited, in 1937, on the incorporation of that Company. During the period 1936 to 1940 a number of surface discoveries were investigated by trenching, test pitting, and diamond drilling. A shallow inclined shaft was sunk on the Brock zone, a system of narrow and irregular quartz veins, and some 75 tons of high-grade ore was mined and shipped to the Trail smelter. In 1943 management control was acquired by Frobisher Exploration Company, Limited. An intensive campaign of diamond drilling was commenced in January 1944 and carried on continuously for 2 years. This drilling indicated the presence of substantial gold orebodies lying within well-developed shear zones. Two shafts, 1 ¼ miles apart, were sunk during 1946 and underground development started on two of the orebodies.The claims are underlain largely by rocks of the Yellowknife greenstone belt, which extends from the shore of Great Slave lake along the west side of Yell ow knife bay and northward . The most conspicuous structural features are the steeply dipping late faults. The chief of these is the West Bay, which bisects the greenstone belt at an acute angle and crosses the southern part of the Giant group. At the 'big bend' near the west boundary of the claims there is a remarkably sudden change in the strike of the fault, from about 12° N.W. to about 3° N.E. The apparent horizontal displacement along this fault is of the order of 3 miles, with the east side moved relatively north. There are also a large number of lesser faults with apparent horizontal displacements of a few feet up to l,70J feet for the Townsite fault. Most of these faults trend north to northwest. Surface examination indicates that, at least for most of them, the movement corresponds to that along the West Bay fault, that is, the east side moved relatively north. According to the available evidence, all or most•of the movem:nt along the West Bay fault and along the lesser faults was post-ore. Late diabase dyk:s that cut through ore-bearing shear zones, with chilled margins in contact with the mineralized material, are apparently displaced to the same extent as earlier geological markers. ( see Frontispiece. )"

Jan 1, 1954

"The United Nations, through the secretariat of the United Nations Economic Commission for Europe (UNECE) and its Ad Hoc Group of Experts on Harmonization of Fossil Energy and Mineral Resources Terminology, developed the Framework Classification.The UN Classification applies to fossil energy and mineral reserves and resources located on or below the Earth's surface. It aims to simplify reporting of energy and mineral reserves and resources in a globalized world and is intended to improve global communications regarding remaining quantities of non-renewable fossil energy and mineral resources. The UN Classification is not recognized as a Foreign Code under NI 43-101.SECTION I1.,The United Nations Framework Classification (UNFC-2009) applies to fossil energy and mineral reserves and resources located on or below the Earth' s surface. It serves the needs for classification at a global level (for energy and mineral supply studies), for governments (resources management and policy formulation), for industry (business process management) and for financial reporting.SECTION II2.,The UNFC is a generic system in which quantities are classified on the basis of the three fundamental criteria of economic and social viability (E), field project status and feasibility (F), and geological knowledge (G), using a numerical coding system. Combinations of these criteria create a three-dimensional system. Categories (e.g. El, E2, E3) and, in some cases, sub-categories (e.g. E1.1) are defined for each of the three criteria as set out and defined in Annexes I and II.3 The first set of categories (the E axis) designates the degree of favourability of social and economic conditions in establishing the commercial viability of the project, including consideration of market prices and relevant legal, regulatory, environmental and contractual conditions. The second set (the F axis) designates the maturity of studies and commitments necessary to implement mining plans or development projects. These extend from early exploration efforts before a deposit or accumulation has been confirmed to exist through to a project that is extracting and selling a commodity, and reflect standard value chain management principles. The third set of categories (the G axis) designates the level of confidence in the geological knowledge and potential recoverability of the quantities.4. The categories and sub-categories are the building blocks of the system, and are combined in the form of ""classes"". The UNFC can be visualised in three dimensions, as shown in Figure 1, or represented in a practical two-dimensional abbreviated version as shown in Figure 2."

Jan 1, 2003

By M. Cai

"It is well known that Peak Particle Velocity (PPV) plays a critical role in triggering rockburst damage in burst-prone underground mines. To estimate the PPV distribution around excavations, advanced numerical modeling can be used. In wave propagation numerical modeling, velocity field is often assumed to be uniform around an excavation. However, it is observed that wave velocities near mine openings are not constant, but rather non-uniform. Wave velocity in a rock mass depends on many factors such as rock quality and confinement. In particular, confinement has a large influence on wave velocity. The confinement around an excavation is low at the wall boundary and increases gradually and reaches a constant value far away from the excavation boundary. Hence, stress redistribution due to excavation will have a profound influence on wave propagation in underground mines. A better understanding of the influence of excavation effect on wave propagation helps to interpret the field observation of rockburst damage and improve mine safety. For this purpose, a non-uniform velocity model need to be constructed and the influence excavation effect should be taken into account in numerical wave propagation modeling.In the present study, a non-linear velocity model is developed firstly by considering the influence of confinement on wave velocity. This model links wave velocity to confinement explicitly. A coupled numerical method is then developed and used to study the seismic wave pattern around excavations in underground mines. Two codes, FLAC and SPECFEM2D, are coupled in the analysis. The motivation of developing the FLAC/SPECFEM2D coupled approach is to take merit of each code, such as the stress analysis capability in FLAC and the powerful wave propagation analysis capability in SPECFEM2D. The coupling process is realized through three steps: (1) static stress redistribution around an excavation is analyzed in FLAC; (2) data for coupling analysis, such as non-uniform velocity field (based on the new developed velocity model), mesh, absorbing boundaries, etc., are generated by FISH functions in FLAC and passed to SPECFEM2D; and (3) wave propagation modeling is performed in the SPECFEM2D considering a fault-slip seismic source. Because stress redistribution around the excavation is considered, non-uniform velocity field for the SPECFEM2D model can be obtained and this is a notable feature of this study. Very large difference in the S-wave response in these two modeling approaches with uniform and non-uniform velocity fields can be seen from the modeling results.Using the FLAC/SPECFEM2D coupled modeling approach, accurate velocity models can be constructed and this in turn will help to produce accurate PPV distribution around the excavations. Accurate PPV contours can be used to identify high rockburst risk zones for improved dynamic ground support strategy."

Jan 1, 2015

By Gunther F. Leichtle, James A. Finch, Colin A. Hardie, Cesar O. Gomez

"The recycling of wastepaper has become an important and common process worldwide. Techniques similar to those used in the mineral processing industry are used to produce high quality recycled fibers for use in newsprint. The mineral processing group at McGill University in conjunction with the Pulp and Paper industry (Mechanical and Chemi-mechanical Wood Pulp Network) has adapted one of these techniques, column flotation, for the purpose of deinking wastepaper. This article describes the typical process of recycling wastepaper from a mineral processing point of view. It also presents results from McGill University's flotation column test work at a Quebec wastepaper recycling facility.IntroductionThe recycling of wastepaper has become an important and common process worldwide, including Canada which is now the world's largest importer of wastepaper (CPPA, 1996). Much of the current drive to increase the utilization rate of recovered paper is a reaction to the solid-waste disposal problem that faces many Canadian and U.S. municipalities. The real problem in the disposal of solid waste is not the absolute amount of garbage that must be disposed of each year but this amount in relation to available disposal capacity.Environmental restrictions on the creation of new landfill sites and incinerators (the two main options for disposal of municipal solid waste) have tightened the demand/supply balance in the solid waste disposal market. This shrinking availability of waste disposal sites in the face of increasing population with greater disposal requirements defines the Developed World's solid waste crisis. Not only are there fewer disposal sites available, but the costs of using them is skyrocketing. These factors have increased government regulations mandating recycled paper content in newsprint. Conservation and environmental issues make the current drive for wastepaper recycling a real issue and business opportunity.A key to the success of producing quality recycled fibers from wastepaper is the deinking technology. The production of tissue papers, paper towels and newsprint requires regenerated fibers which are substantially free of ink particles. Of the various options available, mineral processing techniques can offer the most economical, environmentally acceptable and versatile strategies (Yoon, 1993). Today, it is recognized that flotation is one of the most effective processes for removing ink particles in the production of a high quality product. In a collaborative program with the Pulp and Paper industry, the mineral processing group at McGill University has adapted column flotation technology for this purpose. This article will describe the typical process of deinking from a mineral processing point of view (Part A). It will also present some results from McGill University's flotation column test work at a nearby paper recycling facility (Part 8)."

Jan 1, 1998