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By F. Baerhold, S. Mitterecker

"Acid recovery is an important process step in pickling and leaching processes. Pyrohydrolysis in spray roasters and fluidized beds is a well-established technology to recover hydrochloric acid, but the hydrolysis in the liquid phase is more and more considered as an alternative new process route. The characteristic features of these three processes will be discussed along with select industry trends. For example, hydrochloric acid based leaching processes for ilmenite and nickel laterite ores produce far larger volumes of waste acid streams than the steel industry. Scaling-up of the spray roasting technology to the highest possible capacity is necessary in order to become more cost effective. In addition, process routes based on ferric chloride pyrohydrolysis allow the production of super-azeotropic HCl (up to 30 wt.%). Finally, the new ECOmode concept optimizes the operation of both pickling lines and the spray roast acid recovery plant. Long term results from industrial applications show that fuel savings in the spray roaster of 25% are easily achieved without additional process steps. An attractive feature is also to upgrade existing installations to higher plant capacities.INTRODUCTION Pyrohydrolysis is the conversion of metal halides into metal oxides at elevated temperatures in the presence of water vapor. Reaction products are a hydrogen halide gas and a metal oxide. The process was originally developed and introduced into the market by ANDRITZ Ruthner in the 1960’s. It has found wide technical application in the field of acid recovery and oxide production. The introduction of this process has revolutionized pickling technology. The use of hydrochloric acid has strongly increased over the last decades, since pyrohydrolysis is available for the total recovery of waste pickle liquors from mild steel pickling, as illustrated in Figure 1.Chloride media are also used increasingly in the metallurgical industry. Oxide production processes for Mg, Co, Ni, and Al have been installed. Processes using hydrochloric acid are known for the leaching of ores as well as for the re-extraction of metals from organic solvents. In both cases the recycling of HCl is of substantial importance to achieve highly economic processes and to reduce waste."

Jan 1, 2016

By Dave Rose, David G. Meadows, Mike Westendorf

"Copper Mountain’s concentrator was commissioned in May 2011. The circuit consists of a 34’ x 20’ EGL SAG mill, a Raptor 900 pebble crusher and two 24’ x 39’6” ball mills. Each mill is equipped with a 2 x 8500 HP dual pinion drive system. The plant was designed to process 1585 mtph of SAG feed. From startup, the mill struggled to meet the required production rates. Over time, a major initiative that involved blasting and fragmentation, optimization of the SAG mill in terms of discharge grates, ball size and slurry density, and operation of up to three mobile pre-crushing plants resulted in significant increases in throughput. In parallel, a number of modeling tools, including JK SimMet, were run based on plant benchmarking exercises. Samples collected at the mine site during ramp up confirmed the very competent nature of the ore with drop weight - A*b values of less than 24 being recorded on a variety of plant feeds and lithologies. After due consideration including demonstration of proof of concept, a new pre-crushing circuit incorporating an XL2000 cone crusher and associated screening equipment was designed, constructed and commissioned at the site in August 2014. This paper reviews the history of the original circuit design, the various optimization steps taken and impacts, together with the final pre-crushing circuit and net results.IntroductionThe Copper Mountain Mine is located approximately 20 km south of Princeton, British Columbia, Canada (refer to Figure 1). The region has a long history of copper mining dating back to 1923. The Copper Mountain camp occurs in the 'eastern volcanic belt' of the Nicola Group (Monger, 1989). It is an alkali copper-gold deposit in which the volcanic strata are intruded by early Jurassic dykes, sills, irregular plugs and zoned plutons (Woodsworth et al., 1992). The stratified rocks are relatively fresh with slight metamorphism in the local contact regions."

Jan 1, 2016

By Kristie Tarr

Portland cement is primarily used to solidify backfill in underground mine workings and it represents a major cost in mining operations. CANMET-MMSL has invented an alternative binder technology that does not contain any amount of Portland cement and is made using waste materials, such as development rock and smelter slag. The technology has been successfully tested with many types of base metal and ferrous slag. Research conducted in collaboration with several mining operations in Canada has shown that the CANMET-MMSL binder has considerable potential as a replacement for Portland cement in mine backfill. An integral part of the overall mining operation, backfill is used for two main purposes: safety and environmental remediation. Depending on the mining method used, backfill provides a working surface, stabilizes stope and pillar walls, as well as the surrounding rock mass, and controls caving of stope backs. There are several benefits associated with the replacement of Portland cement in backfill, including reductions in (i) the footprint by placement of slag waste underground, (ii) the environmental effects and long term liability resulting from disposal and containment of slag waste on surface such as acid rock drainage, (iii) GHG as a result of reduced Portland cement usage (the production of 1 ton of Portland cement produces 1 ton of GHG), and (iv) the significant costs associated with reduced Portland cement usage.

Oct 1, 2010

By R. Webber, D. Averill, L. Whittle, D. Moffeti, G. Kassakhian

"This paper describes the recent pilot-plant work carried out on two alternative processes capable of producing an effluent containing less than 10 pCi/L total radium-226 activity and very close to 3 pCi/L dissolved radium-226. The processes incorporate coprecipitation of radium with barium sulphate by the addition of barium chloride in continuously stirred tank reactors followed by either chemical coagulation with ferric chloride, flocculation and clarification, or polymer-aided dual-media filtration. IntroductionUranium mines in Canada have practiced radium removal from tailings basin effluents since 1965. Treatment consists of addition of barium chloride to precipitate radium as barium- radium sulphate and provision of natural or constructed settling ponds for sedimentation. This process, however, has been identified as being deficient in that it allows particulate or suspended radium-226 releases which are higher than the regulated dissolved values». Also, the large size of some of the settling ponds is believed to represent a real limitation of the present technology because of inferred difficulty in recovery of barium-radium sulphate precipitates.The Federal Government Mining Task Force established a Radioactivity Sub-Group in 1974 to assist in the development of effluent regulations and guidelines for the uranium mining industry. Among the recommendations of the Sub-Group was the need to seek more effective methods of radium removal, including development of physical/chemical treatment systems as an alternative to settling ponds. In response to the SubGroup recommendations, a bench-scale study began in March 1976at the Wastewater Technology Centre (WTC) of Environment Canada . The results of this study indicated that coagulation and flocculation of the precipitate from barium chloride addition is a viable alternative to sedimentation in settling ponds(3). The bench work was expanded and accelerated in 1977, with financial assistance provided by the Atomic Energy Control Board. Subsequently, a research program with a larger scope was developed; this program is jointly managed and funded by government and industry, as shown in Table I.One of the goals of the joint program is to develop, at pilot scale, a physicalI chemical process to reduce the radium-226 content of uranium mining and milling effluents. Target radium-226 activities for treated tailings basin effluents were set: namely, 10 pCil L total and 3 pCil L dissolved radium 226. Extensive bench-scale and pilot-scale work was conducted through 1978 and 1979 and has culminated in the development of two physical/chemical processes. This paper describes work carried out at a pilot plant which led to the selection of the two physical I chemical processes for demonstration. In short-term tests, the processes produced effluents meeting the radium-226 target activities."

Jan 1, 1980

By Jocelyn Bouchard, Claude Bazin, Sami Makni, Rolando Quispe

"Finely disseminated minerals in ores require intensive size reduction to achieve sufficient liberation for the subsequent concentration process. The use of small grinding media in conventional ball mills as a replacement for the usual 2.5 to 3.8 cm balls (1.0 to 1.5 in) produces a finer size distribution at a lower power drawn by the mill motor. However, the conditions ensuring a successful application of using small media for fine grinding are still not fully understood and some plants faced issues to keep the newly introduced steel charge inside the mill during industrial trials. The retention conditions of small media are investigated in this paper using a continuously operated 40 x 40 cm laboratory grinding mill. The tests assess the effect of the steel charge level, the slurry solids concentration and the feed rate on the rate of ejection of the grinding media. The dominant variable is found to be the filling of the mill with the grinding media, followed by the mill feed rate. Results show that increasing the slurry solids concentration within a grinding mill may be favorable for the retention of the media inside the mill.IntroductionProcessing of complex ores often requires the use of fine grinding to liberate finely disseminated minerals. Grinding requirements for a P80 of less than 0.02 mm are not uncommon, and different grinding mills are available for such application (Leonida, 2014; Pease et al., 2006; Davey, 2006; Orford et al., 2006; Nesset et al., 2006). Even conventional horizontal ball mills charged with of 2.5 cm or larger grinding media can be used for fine grinding. However, simulation (Orford and al., 2006) and plant observations (Alves and L-Bouchet, 2005; Brisette, 2010; Brisette and Roman, 2012) demonstrated the advantages of using smaller grinding media in terms of size reduction and energy consumption."

Jan 1, 2015

By K. Nasrallah

"The Dumont deposit is a large (1,179 Mt proven and probable reserve) low grade (0.27% nickel), nickel sulphide project located in the Abitibi region 25km west of Amos, Quebec. Royal Nickel Corporation acquired the project in 2007 and has since taken the project from a scoping study in 2010 through to completing the feasibility study in July 2013.The Dumont deposit in a nickel deposit with the recoverable nickel contained in three minerals; pentlandite, haezlewoodite and awaruite. The minerals are recovered by a combination of flotation and magnetic separation. Over the course of the studies many flowsheet design decisions and changes were made. This paper outlines an overview of the project, flowsheet design including the comminution circuit, desliming and awaruite recovery. The overall resulting flowsheet that was used for the feasibility study is then presented along with the confirmatory locked cycle testing that showed the feasibility design basis was able to produce the predicted concentrate grade and recovery.IntroductionRoyal Nickel Corporation (RNC) is a mineral resource company headquartered in Toronto, Canada, primarily focused on the exploration, evaluation, development and acquisition of base metal and platinum group metal properties. RNC's principal asset is the Dumont Nickel project (Dumont project) located in the established Abitibi mining camp, 25 km northwest of Amos, Quebec. RNC acquired a 100% interest in the Dumont property in 2007. The mineral claims covering the Dumont deposit are currently held 98% by RNC and 2% by Ressources Québec.The Dumont project is located in the province of Quebec in the municipalities of Launay and Trécesson approximately 25 km by road northwest of the city of Amos, 60 km northeast of the industrial and mining city of Rouyn-Noranda and 70 km northwest of the city of Val d’Or."

Jan 1, 2016

"The Key Lake mill was commissioned in 1983 to process the two Key Lake ore deposits. With the depletion of these ore bodies in 1999, the plant was converted to mill the exceptionally rich McArthur River deposit located seventy eight kilometers northeast of the Key Lake mine site. This paper describes in detail the Key Lake milling process. The mill consists of a grinding circuit, a leachkounter current decantation circuit, a solvent extraction circuit, a yellowcake precipitatiodcalciner circuit, an ammonium sulphate crystallization circuit, and a waste treatment circuit. The paper also describes process changes to handle McArthur River ore, including the ore receiving stationLOCATIONThe Key Lake mill is located 78 kilometres south of the McArthur River mine. It is located at latitude 57' N in northern Saskatchewan, within the southern boundary of the Athabasca formation. The nearest town is Pine House, located 240 kilometres to the south. Access to the mine is by an all weather road and by a private airstrip.HISTORYAND DESCRIPTION OF OPERATIONThe Key Lake mill was commissioned in October 1983 to process ore from the Gaertner and Deilmann ore bodies. The mill was designed to produce twelve million pounds of uranium oxide (U308) a year from a feed grade of 2.4% U3O8.Following grinding, the uranium is leached in a two stage leach circuit using oxygen and sulphuric acid. An eight-thickener counter current decantation (CCD) circuit is used to separate the uranium bearing solution from the solid waste residue. The uranium is further purified in a solvent extraction circuit and then precipitated using ammonia. It is then calcined at 840' Celsius, and packaged into drums for shipping to refineries. There is a crystallization circuit to recover ammonium sulphate produced in the process and a waste treatment circuit."

Jan 1, 2001

By I. C. L. Webster, G. E. Ray

"The magnetite-rich Heff Cu-Au skarn lies in the Quesnel Terrane of south-central British Columbia, approximately 26 km northeast of Kamloops. The skarn formed close to the northern margin of the Heffley Creek pluton, and is hosted by Carnian (Late Triassic) Nicola Group limestone. The Late Triassic pluton (208.1 ± 6.1 Ma, U-Pb zircon) is a composite body of alkaline affinity that includes dioritic, gabbroic, and magnetite-rich clinopyroxene ± olivine ultramafic phases.The skarn has been discontinuously traced along strike for more than 1 km, and individual skarn layers are up to 40 m thick. The most common non-opaque minerals are garnet, clinopyroxene, and carbonate, with lesser epidote, biotite, amphibole, and chlorite. Apatite locally forms more than 3 vol.% of the rock. Compositionally the garnet is a low Mn grandite averaging Pyr3-Grs41-Adr56 mol.%.Mineralization consists of lenses, up to 10 m thick, containing massive low Ti-V magnetite (avg. 0.03 wt.% TiO2 and 0.01 wt.% V2O3). Magnetite is associated with up to 20 vol.% pyrrhotite, lesser pyrite, and trace chalcopyrite and gold. There is a moderate enrichment in rare earth elements with assays up to 570 ppm La and 490 ppm Ce. The Heffley Creek pluton belongs to a suite of Late Triassic-Early Jurassic alkalic and Fe-rich intrusions that are widely developed throughout the Quesnel Terrane. Some host Cu-Au porphyry mineralization, but less commonly they are associated with magnetite-apatite veins such as at the Glen Iron mine and Magnet occurrence, or with magnetite-rich Cu-Au skarns such as the Heff occurrence.Although abundant low Ti magnetite is present in the Heff skarn, it does not appear to represent iron oxide copper- gold (IOCG) type mineralization. Microprobe analyses completed during this study show that low Ti magnetite is not a unique feature of IOCG deposits, but also characterizes some other types of hydrothermal mineralization. In British Columbia these include some alkalic porphyry-related magnetite-apatite veins in the Quesnel Terrane, as well as the well-known Wrangellia-hosted Fe skarn deposits. Thus, low Ti magnetite appears to be a common signature of many magmatic hydrothermal deposits in contrast to the high Ti primary magnetite found in most igneous rocks."

Jan 1, 2007

By L. K. Daneshmend

"Competitive pressures dictate that the mining industry moves to larger mobile underground haulage equipment with more advanced equipment mechanization, higher levels of autonomous operation of machines, and longer periods between access by maintenance personnel. Since such new machines will of necessity be more complex than existing designs, machine reliability has become a key issue in progressing toward advanced mobile underground haulage equipment. Reliability and maintainability models of such equipment are practically non-existent in the literature. This paper addresses the issue of reliability and maintainability of mobile underground haulage equipment on two fronts: reliability modelling to enhance decision making for maintenance planning; and reliability simulation to enable evaluation of more rigorous production planning scenarios. A case study from a gold mine in the Chilean Andes is presented, for which reliability models of mobile underground haulage equipment, specifically load-hauldump vehicles and underground haul trucks, are developed, and a reliability-based simulation implemented. IntroductionMaintenance accounts for 20% to 35% of the overall operating budget for a typical underground operation, and hence represents the largest portion of the mine’s controllable operating costs (Cutifani et al., 1996; Campbell, 1997; Goddard, 1998). This, coupled with the fact that industry today is faced with an ever more complex and demanding marketplace, has led to an increasing interest in methods to reduce maintenance costs.Although significant effort has gone into developing effective maintenance strategies for industry in general, application of these to the"

Jan 1, 2003

By Glenn Kosick, Roger Amelunxen, Glenn Dobby

"The variability in flotation response within an orebody is often a significant problem to address during the design phase of a flotation circuit. In addition, for existing operating plants, the reasons for such a variability need to be clearly understood and accounted for if plant performance is to be optimized. MinnovEX Technologies, in collaboration with major mining companies, is developing methodologies and tools to address the issue of variability in ore characteristics (such as ore hardness and flotation response). The tool is used for design and scale-up of flotation circuits from lab flotation data, and for prediction of recovery and final concentrate grades throughout the mine block model for any given circuit design.INTRODUCTIOTNH: EI SSUEB EINGA DDRESSEDThis paper describes the current development of a methodology and software tool for improved flotation plant design and enhanced prediction of froth flotation performance in an operating plant. The need for this tool is attributed to two primary factors: variability of flotation response properties throughout an orebody and related difficulties with scaleup from laboratory tests. Consider these in more detail.It is well recognized that separation performance via froth flotation will vary significantly within a single mineral deposit. This has been enhanced through the wide application of SAG milling. Throughput with SAG mills is more susceptible to ore hardness changes than is the case with crushing-rod-ball circuits; it is not uncommon to see SAG mill throughput to vary by up to 100% on a regular basis, e.g. 1000 tph to 2000 tph.The implication of this on laboratory testing for prediction of metallurgical performance is as follows: it is unlikely that either the plant grind (size distribution) feeding froth flotation or the flotation residence time will be known at the time of laboratory testing for a given block of ore. Hence, the grind used for the laboratory test is not likely to be that which actually occurs once the ore is processed, and the laboratory retention time may not represent the plant conditions."

Jan 1, 2002

By Peter C. Jones

The paper gives a broad view of the maintenance procedures at the Cassiar Asbestos Corporation property. Comparisons are drawn to indicate the changes in facilities and approach that have evolved over the last 25 years while coping with the northern British Columbia climatic conditions. Special emphasis has been placed on the importance of the support groups providing services for both the mine site and the town.

Jan 1, 1978

By P. D. R. Maltby

"With the narrow vein structure of Nova Scotia gold deposits, selective mining is rarely possible and mill heads will be too low (below 0.10 oz per ton of gold) to support a conventional mill of up to 1000 ton per day capacity at present gold prices. The recent operation of a small 100 t.p.d. mill at Cochrane Hill using fine grinding - leaching - filtration supported this assertion. Milling costs exceeded $25.00/ton of feed mainly due to high capital, labour and maintenance costs.Accordingly, since 1981, efforts have been directed towards the development of a cheap gold extraction process in terms of both capital and operating costs. The main criteria is to produce gold below US $200 per oz including all administration, mining and process stages. This paper describes such a process development for Cochrane Hill. LOCATION AND HISTORY OF THE PROPERTYThe Cochrane Hill property is located about 8 miles North of Sherbrooke, N.S. on highway No. 7. Sherbrooke is about 180 miles east of Halifax. Gold was originally discovered at Cochrane Hill in the 1860's. Early work consisted of trenching. In the 1930's a small shaft was sunk to about 80 feet and some tonnage treated using a stamp mill with gravity concentration of gold.Northumberland Mines Ltd. acquired a 70% interest in the property in 1977. Subsequent drilling indicated a potential zone of gold mineralization with a drill-indicated tonnage of one million tons having a grade of 0.07 ounces per ton of gold (all high assays cut to 1 oz) down to a depth of 300 feet recoverable by open pit mining methods. In broad terms, the goldbearing zone is about 50 feet wide and 800 feet long and dips at about 70°."

Jan 1, 1984

"The Dumont deposit is a large (1066 Mt probable reserve) low grade (0.27% nickel), nickel sulphide project located in the Abitibi region 25km west of Amos, Quebec. Royal Nickel Corporation acquired the project in 2007 and has since taken the deposit through pre-feasibility. And updated, revised pre-feasibility was released in July 2012.The disseminated nickel sulphide and alloy mineralization is hosted within the serpentinized dunite subzone of the Dumont Sill; a differentiated, pervasively serpentinized, ultramafic sill of komatitic affinity. Pentlandite, heazlewoodite and awaruite are the dominant nickel-bearing, metallic phases and can be recovered by conventional methods (flotation, mag separation). Non-recoverable nickel is also present within the silicate matrix of the serpentinized dunite and is locked in serpentine (Fe and Mg-rich variants) and minor olivine.A detailed mineralogical program including 1097 QEMSCAN® samples and 4300 electron microprobe points were undertaken to increase the understanding of the nickel deportment in the various recoverable and non-recoverable phases in a spatial context across the deposit. Eighty-three (83) metallurgical tests performed under a standard procedure have linked the mineralogy to metallurgical performance. This has allowed the creation and modelling of geometallurgical domains on a block by block basis to understand the recovery variation and metallurgical performance throughout the deposit. Understanding the spatial variability of Ni deportment in recoverable and non-recoverable phases through mineralogy is the key to understanding metallurgical performance thus, unlocking the value of the Dumont deposit and other large tonnage low grade ultramafic nickel deposits."

Jan 1, 2013

By G. Bellefleur, K. Stevens, E. Adam, D. Synder, G. Perron, C. Müller

The Downhole Seismic Imaging consortium was initiated by the Geological Survey of Canada and three Canadian mining companies to assess the usefulness of borehole seismic techniques for the exploration of massive sulphide deposits within areas of steeply dipping stratigraphy. Field tests were conducted over two sub-economic massive sulphide deposits located in Norman West, Ontario, and Halfmile Lake, New Brunswick, using various shot configurations and a three-component borehole receiver array. The Norman West deposit, or structures associated with the deposit, produce detectable seismic reflections as far away as 1500 m. However, these reflections are difficult to distinguish from fault reflections and lithological contacts near the base of the Sudbury Igneous Complex. At Halfmile Lake, the rock properties are favourable for direct orebody detection. The Halfmile Lake deposit has a clear but relatively complex seismic signature on the downhole data, showing not only P-wave reflections, but also strong reflected S-waves and modeconverted waves. Such prominent S-waves and converted waves have not been recognized previously in borehole or surface seismic data acquired over massive sulphide deposits. These waves add complexity to the interpretation but provide additional imaging capabilities. Reflected S-waves and converted waves should be considered when designing seismic surveys for mineral exploration.

Jan 1, 2004

By R. Hamilton

"Louvicourt mine began operation in 1994. It is located in northwestern Quebec, 20 km east of Val d’Or. The mine is owned by Aur Resources (30%), which operates the mine, Teck/Cominco (25%) and Novicourt (45%). By the end of 2002 there were approximately 280 employees working on the site, including contractors. Production averages 4300 tons per day, working a massive sulphide body located between 415 m and 920 m underground.Maintenance expenses represent 27% to 29% of the operating costs, and the maintenance personnel represents 34% of the total workforce (including electrical, mechanical and concentrator maintenance staff). Considering these figures, in addition to capital costs and indirect costs, such as energy (electricity, fuel, gas, etc.), it is worthwhile to pay attention to how things are done in order to protect such investment. Every task performed to help understand the impact of Louvicourt’s practices must to be seen as an investment in the reliability of its operation."

Jan 1, 2003

By Peter Radziszewski

"Typical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5% to 20% ). Cornminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and linerkharge wear (50 %). Developing the ability of predicting mill wear will allow a more accurate estimation of wear costs. However, total media wear in a given ball mill grinding process is a product of three recognized wear mechanisms those being abrasion, corrosion and impact. Further, the contribution to total media wear of each of these wear mechanisms has not been well established. In this paper, a overview of abrasive, corrosive and impact wear mechanisms is made from which an empirical wear prediction procedure is proposed. Further, future directions in wear prediction research is outlined. INTRODUCTIONTypical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5 % to 20 %). Comminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and lined-charge wear (50 %). With respect to wear, world wide steel consumption as grinding media is estimated at over 600 000 tomes per annum (Malghan, 1982). As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs.Charge media wear prediction is actually estimated using the Bond abrasion test. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2000

By S. H. Ward

An account of a new method of investigating the electrical .properties J the earth's crust is presented. The method, called AFMAG, employs, as a source, natural alternating magnetic fields of audio and sub-audio frequencies. Measurements of the distortion of these fields caused by geological features are made with the aid of search coil detectors at several discrete frequencies. Many results of field surveys made with the method are presented and they illustrate its advantages and limitations when applied to the search for massive sulphide mineralization. The advantages far outweigh the limitations and in fact indicate that a major breakthrough in geophysical exploration technique and instrumentation has been made.

Jan 1, 1958

By N. I. Basuki, E. T. C. Spooner

"The Bongara area, northern Peru, contains Zn-Pb mineralization with characteristics typical of Mississippi Valley-type (MVT) systems. The northwest-southeast-trending late Triassic-Jurassic car¬bonate host rocks of the Pucará Group were locally replaced by early (D1) and late (D2) dolomite. Fluid inclusion microthermometric data from dolomite and sphalerite are typical of MVT ore fluids (78° to 187°C; ~15-23 wt.% CaCl2 equivalent), and do not show apparent covariation between temperature and salinity.Homogenization temperature data from D1 and D2 dolomite show a decrease from northern (mean Th = 138°C) to central (mean Th = 126°C) to southern areas (mean Th = 102°C), suggesting fluid flow from north to south. Assuming a geothermal gradient of ~30°C/km and an average surface temperature of ~20°C, the lower temperature limits might approximate the ambient host-rock temperatures, whereas the upper temperature limits may represent minimum temperatures of a deeper aquifer, probably at a maximum depth of ~6000 m. The lower temperature limits are also similar to the estimated temperatures if dolomitization and mineralization took place in the Late Cretaceous at ~2000-3000 m depth, probably during the formation of the Peruvian fold and thrust belt near the present day Peru-Ecuador border.Fluid inclusion data show that the fluids in each studied location had relatively uniform temperatures and salinities during dolomitization and mineralization. A significant salinity drop and a slight temperature decrease are observed during late calcite precipitation, suggesting mixing of brines with dilute, cooler fluids."

Jan 1, 2009

By H. McSorley

"I note that many of you are involved in the production of zinc, lead and copper concentrates and metal, which I am the most familiar with among the non-ferrous metals. This afternoon, I will describe briefly the responsibilities of a commercial manager within a mining/metallurgical company, the basic framework of a non-ferrous metals concentrate agreement and lastly, a brief discussion of the copper, zinc and lead concentrate and metal markets.I hope this talk will contribute to a better understanding of the business between all of us and look forward to a discussion after the presentation.First of all, I work for Noranda Sales, which is the marketing and sales agent for Noranda Minerals' three operating units: Noranda Copper, Noranda Zinc and Mining Corporation. In addition, Noranda Sales acts for other mining companies in the marketing of their raw materials and metals. Noranda Sales is responsible for procuring all raw materials for Noranda's metallurgical plants, such as the Copper Group's Forne and Gaspe smelters, as well as the sale of surplus concentrates and by-products and metal products. In addition, Noranda Sales trades metals and concentrates and owns a metal brokerage business, Rudolf Wolff, a founding member of the London Metal Exchange. It does $3 billion of business in 60 countries, exporting some 80% of Noranda Minerals' output to the world.What are the responsibilities of commercial management in resource marketing to the mine and metal producers?Resource marketing and commercial administration involves assistance in establishing exploration priorities, contributing to feasibility studies, pre-production planning, product promotion and realization of income for many properties and metallurgical plants through completion of supply and sale agreements with other companie"

Jan 1, 1988

By Rob Sloan, Simon Yu, Rob Robison, Mark Schmidt

"In 2000-2009, Highland Valley Copper (HVC) tested new generation 660 mm cyclones supplied by various manufacturers in an aim to replace existing first generation 762 mm cyclones without success. Due to plant infrastructure limitations, new generation larger (838 and 800 mm) cyclones were not feasible. In 2010, an approach of testing hybrid 800 mm diameter cyclones was initiated by HVC. The hybrid cyclone consists of an 800 mm cyclone upper head section and the use of existing 762 mm cyclone conical sections. Three hybrid 800 mm cyclones were engineered and fabricated, targeting a throughput of 850-950 mtph and were installed in the E line cyclone cluster. Since starting operation on September 20, 2011, three hybrid 800 mm cyclones have been running at combined throughput ranges of 750 to 1050 mtph. Based on plant Daily Met Reports, the test hybrid cyclones demonstrated a decrease in misplacement of coarse copper bearing material to the cyclone overflow and better cyclone efficiency when compared to the existing 762 mm cyclones at the same throughput. The hybrid 800 mm cyclones also demonstrated significantly higher capacity per unit when compared to the existing cyclones. Based on the improvement demonstrated in the full plant trials, HVC decided to install the new hybrid 800 mm cyclones in all five grinding lines starting in July, 2012. A more detailed discussion of the full plant trials and initial results of the first complete retrofit of E line will be presented. INTRODUCTIONHighland Valley Copper Partnership (HVC) is Canada’s largest base metals mine and one of the world’s largest producers of Cu-in-concentrate. The mine, located in the central interior of British Columbia, is owned (97.5% interest) and operated by Teck Resources Ltd. In 2011, HVC produced about 210 million lbs. of Cu and 8 million lbs. of Mo. HVC’s Mill Modernization Project (MOP), which was budgeted for $475 M, is currently under intensive construction and is scheduled for commissioning in Q1 2014. The following description was valid at the time of writing this paper.The operation has been expanded a number of times since the startup of the original Lornex mill in 1972. The plant started up with two 9.75 m (32 foot) diameter primary SAG mills. It was expanded in 1981 with the addition of a 10.36 m (34 foot) diameter SAG mill. In 1988, the Highmont concentrator grinding mills and flotation section were moved and attached to the Lornex mill site. This added two10.36 m (34 foot) diameter AG primary mills to the Highland mill. This brought the total number of primary mills to five referred to as Line A, B, C, D and E: three SAG mills (A, B and C) each with two ball mills in parallel and two AG mills (D and E) each with one ball mill. Each AG mill line is equipped with one short head pebble crusher. Ball mill circuits are classified using 8 clusters of ten Linatex 762 mm cyclones to reach a target P80 330 micron (equivalent 43% +100 mesh), before the cyclone overflow product advances to the bulk flotation circuit."

Jan 1, 2013