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By Henry Smith, Stu Brown

"The Williams Mine is Canada's largest gold producer, with annual production of some 500 000 ounces of gold from two million tonnes of ore. Highly mechanized and automated underground mining practices are utilized at the property in order to help achieve high levels of productivity and the most cost-effective mining possible.Planning and design parameters for the underground ore handling system at the Williams Mine are described in this paper. Particular reference is made to the considerations that were built into the system to provide maximum operating flexibility through the use of multiple orepass systems, portable low profile jaw crushers, underground conveyors and automated hoisting equipment.IntroductionThe Williams Mine is situated in the Hernlo area of northwestern Ontario, approximately 40 Ian east of the town of Marathon (Fig. I). It is operated by the Williams Operating Corporation, which is owned equally by Teck Corporation and Corona Corporation, and is one of three adjacent properties forming the HernIo gold camp.The mine commenced operation in October 1985 at a nominal production rate of 3000 tpcd. Through an ongoing capital development program, winch was completed in late 1989, the production rate at the property was increased to its current level of 6000 tpcd. This paper describes the mine planning process and the evolution of current mining strategies with particular reference to the underground ore handling system .HistoryGold mineralization was discovered in 1944 in the Hemlo area of Northern Ontario, near the north shore of Lake Superior. A year later, 11 claims were staked over the showing and subsequently patented by the late Dr. Jack Kerr Williams.Over the next 35 years, other claims were subsequently staked, explored and abandoned in the immediate area of the 11 patented claims . In 1980 a predecessor of the Corona Corporation acquired claims adjoining the Williams property to the east. Drilling on this claim group started early in 1981 and intersected the Hemlo orebody in May of that year. In December 1981, Corona Corporation entered into an option agreement with Teck Corporation to explore and develop the property.A drilling program by LAC Minerals on the II patented claims discovered the Williams orebody in March 1982. The mine started production in 1985, pouring its first official bar of gold on December 6 of that year. Corona Corporation was awarded ownership of the Williams Mine conclusively by a decision of the Supreme Court of Canada on August 11, 1989. The court's ruling ended a legal dispute with LAC Minerals over the property initiated eight years earlier."

Jan 1, 1991

By M. Oliazadeh, A. Vazirizadeh

"The inclusion of small amounts of silicon, phosphorus and sulphur in iron ores or concentrates can have a profound effect on the steel making process and thus on the final iron and steel products. In nature however, it is inevitable that iron deposits contain unwanted elements. Each element has its own distinct effect on the final steel product. The effect of high silicon is the formation of gray iron. Phosphorus mainly increases hardness and strength, lowers solidus temperature and increases fluidity and cold shortness. Finally sulphur contamination causes cracks in the iron texture when smelting and forms white iron when casting. Overall, the acceptable level of phosphorus, sulphur and silica in the iron concentrate should be less than 0.04%, 0.15 % and 1.5 % respectively.Scarcity of high grade iron ores necessitates using lower grade ores with higher amounts of impurities. The purpose of iron concentrators is to increase iron grades to an appropriate level suitable for steel making as well as removing impurities. Although many impurities can be traced in the ore, the most predominant impurities are limited to one or two elements. In this paper the beneficiation methods for removing SiO2, P and S are presented, including flotation, and the parameters that need to be considered in the design of beneficiation plants are discussed. Finally industrial cases are discussed along with their metallurgical performance.INTRODUCTIONIn metal sector, value of iron ores and concentrates is determined by iron percentage as well ascontent of:• Useful elements (Mn, Ni, Cr, V, Ti): natural alloying elements in steel• Deleterious impurities (S, P, As, Zn, Pb, K, Na): impair metal properties or complicate operational process in blast furnace• Fluxing components (Si, Ca, Mg,, Al)Sulphur and copper result in red-short, phosphorus is the cause of cold brittleness and arsenic and copper decrease welding characteristics. Sulphur should not exceed 0.15 in commercial ore. Acceptable sulphur content in ores and concentrates for producing sinter and pellets is 0.6%max, as sintering and heat hardening of pellets remove sulphur 60-90%. Phosphorous content in ores, sinter and pellets is limited to 0.07%-0.15%. 0.00.1% As max, 0.1-0.2% Zn max and up to 0.2% Cu are allowed in iron-ore part of the blast furnace burden for the production of standard conversion pig iron. Slag-forming components are divided into basic oxides (Ca, Mg) and acidic (Si, Al). Higher ratios of basic to acidic oxides are preferred since this reduces raw fluxing material."

Jan 1, 2016

By D. Bronkhorst, J. Rheault

"Backfill practices at the Williams mine have evolved significantly since underground mining began in 1986. The use of conveyors as a means of transporting fill material to stopes has been discontinued in favour of load-haul-dump equipment. The method of adding binder to the fill product has been refined along with the placement technique. Slurry preparation has been fully automated and flyash is now included as a binding agent. Surface preparation of the fill product has evolved substantially with the quarry operation coming into full operation and the construction of pennanent crushing, screening, and storage facilities. Company personnel have replaced all contracted labour involved in backfill operations. Some changes have been made to achieve a higher standard of quality control while others were implemented as a matter of cost reduction. BackgroundThe Williams mine is one of three mines operating at the Hernlo gold deposit, located 40 km east of Marathon, Ontario, along the north shore of Lake Superior. Daily production at the mine is 6000 Ud yielding an annual gold production of approximately 500 000 oz. The orebody extends from surface to a depth of 1300 m dipping north at 70 degrees. The average orebody width is 25 m, but varies from 45 m on the eastern boundary to about 10 m at the western probable limit and at depth. Proven and probable reserves currently stand at 32 million tonnes grading 6 g/t.The Williams orebody has been divided into four separate mining zones: the A-zone, the West-of-Dyke zone, the B-zone, and the C-zone (Fig 1). Ore production originally was from the A-zone open pit operation while the A-zone and B-zone underground workings were being developed. The mining method employed underground is blasthole open stoping with delayed rockfill. Mining block dimensions are nominally 20 m along strike by 25 m high. Stope extraction is sequenced to achieve a relatively short pillar life and 100010 ore recovery. Pillars are extracted as soon as the adjacent primary stopes are two cycles ahead. Primary stopes are filled with consolidated rock fill after mining. This allows pillars to be blasted right against the consolidated fill, but also requires that the fill be self supporting to a height of 30 m.The Williams mine places approximately 1.4 million tonnes of rockfill and 50 000 tonnes of binding agents each year, with about three quarters of the fill placed being consolidated. The consolidated fill placed at the Williams mine contains Portland cement and flyash as binding agents."

Jan 1, 1994

By Lluís Fontboté, François Lieben, Robert Moritz, Massimo Chiaradia

"Abstract - The Las Cañas stratabound Pb-Zn deposit, in northern Chile, is hosted by the Lower Cretaceous back-arc marine carbonate and subordinate andesitic volcanic and volcaniclastic rocks of the Chañarcillo Group, one hundred meters above a thick pile of magmatic arc-related basaltic to andesitic lavas of the Bandurrias Formation. The deposit is not obviously related to local intrusive rocks, which include an Upper Cretaceous granodiorite batholith and dacite dikes emplaced along normal faults. Stratabound Pb-Zn-Ba mineralization is constrained within several porous stratigraphic units, including a 50 cm thick lapilli-tuff and less well mineralized algal-laminated limestone. Veins that locally carry some Pb-Zn occur along normal faults that displace the stratabound orebodies, and this study shows that they represent a different type of mineralization.Fluid inclusion microthermometric data for ore-stage quartz and barite indicate saline fluids (6 to 30 wt% NaCleq) and a formation temperature in the range of 80°C to 160°C for the stratabound mineralization. Two types of fluid inclusions occur in the veins. The type I vein fluid inclusions indicate less saline (average 7 wt% NaCleq) and slightly higher temperature fluids, in the range of 110°C to 220°C. The type II vein fluid inclusions indicate highly saline fluids (25 to 40 wt% NaCleq) which might be related to minor copper mineralization.206Pb/204Pb, 207Pb/204Pb and 208Pb/204Pb ratios for galenas from the stratabound mineralization, with mean values of 18.42, 15.59, and 38.25, respectively, fall within the range of stratabound polymetallic deposits of northern and central Chile, and indicate a Lower Cretaceous magmatic rock source. Galenas from veins have slightly more radiogenic 206Pb/204Pb, 207Pb/204Pb and 208Pb/204Pb ratios, with mean values of 18.47, 15.61, and 38.34, suggesting two distinct episodes of mineralization.?34S values for galena and barite range from negative values (-31.2‰ to -12.5‰ and -7.7‰ to +12‰, respectively) in the mineralized tuff to higher values in the algal-laminated limestone (-2.5‰ to +1.2‰ and +13.2‰ to +17.4‰, respectively), close to Lower Cretaceous marine sulfate values for barite, and homogeneous values for the galena from veins (-12.0‰ ±1‰). Changing redox conditions can partly account for ?34S variations in the tuff. Sulfur sources include magmatic rock sulfur (-1.5‰ to +9.7‰) as well as sedimentary sulfides and marine sulfate."

Jan 1, 1999

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 Dan Zou, Ji Chen, Deqian Li

"Bayan Obo mixed RE concentrate is the largest RE source in the world. RE elements occur dominantly in bastnaesite and monazite. This paper describes a clean and green technique for processing Bayan Obo mixed rare earth (RE) concentrate. A two-step oxidation method can lead to the effective recovery of REs, thorium (Th), fluorine (F) and phosphorus (P). Ce has a property of valency change, enabling the separation of it from other RE(III) elements. Making use of the interactions of cerium (IV) (Ce(IV)), F and P, a new extraction system was created to separate Ce(IV) effectively from other RE(III) while recovering F, P and Th by complexing F and P following the oxidation of Ce(III) to Ce(IV). The method includes two parts. Firstly, bastnaesite is oxidized by air roasting and sulfuric acid leaching. And then a first leach liquor containing Ce(IV), F and P is obtained. Secondly, monazite was converted into REhydroxide through alkaline attack and then Ce(III) in the RE-hydroxide was oxidized into Ce(IV) by wetair oxidation. After sulfuric acid leaching the oxidized RE-hydroxide, a second leach liquor is obtained. The two parts of leach liquor containing Ce(IV), F and P are then extracted by Cyanex923 to obtain CeO2, CeF3 and CePO4 based on the complex of Ce(IV) with F and P. The raffinate was further extracted by the primary amine N1923, and Th-product was recovered as ThO2. The RE elements recovery yield during the whole process is greater than 92%. The new technique not only provides a new way for Bayan Obo RE ore utilization, but also avoids resource waste and environment pollution.INTRODUCTIONRare Earths (REs) refer to lanthanides, including scandium and yttrium. Owing to the similar structure and property of RE(III), it is difficult to separate them from each other, especially for the adjacent elements. According to the United States Geological Survey, in 2010, the whole world has about 142 million tonnes of rare earth reserves. In that, China ranks the second with 22% following Brazil which ranks first with 32% (Cheng, 2011). There are three main RE ore types in China with about 80% mixed ores in Bayan Obo, 10% bastnaesite ores in SiChuan and ShanDong province and 2.9% ion adsorption ores in Southern China. Bayan Obo mixed RE concentrate consists of bastnaesite and monazite. The concentrate contains CeO2/_REO of 50% and there are 7–8% F, 4-6% P and 0.2–0.3% Th."

Jan 1, 2016

By J. J. Taute, O. S. Tinkler, G. Zárate, K. C. Sole, J. Steeples

Global practice in copper solvent extraction is reviewed, based on individual plant operating data for 2018. The survey respondents account for almost 40% of global copper cathode production by solvent extraction and electrowinning. Regional trends in operating conditions, as related to the prevailing ore mineralogy, leaching processes, and life of mine are evaluated. The adoption and expansion of this technology in the African Copperbelt and Asia are assessed, together with the associated challenges of specific orebody characteristics. Trends in feed and electrolyte compositions and flow rates, and extractant and diluent selection, together with their impact on copper recovery are reviewed. The evolution of equipment design and associated ancillary processes are discussed, and new projects identified. INTRODUCTION Continuing from similar surveys carried out in 1997, 1999, 2003, 2007, and 2013 (Robinson, Davenport, & Jenkins, 1997; Jenkins, Davenport, Kennedy, & Robinson, 1999; Robinson, Sandoval, & Cook, 2003; Robinson, Moats, Davenport, Karcas, Demetrio, & Domic, 2007; Sole, Zarate, Steeples, Tinkler, & Robinson, 2013), global operating practice in copper solvent extraction (SX) in 2018 is reviewed. Contributing operations include established sites in the USA (8), Chile (9), Peru (2), Spain (1), Mexico (2), and Australia (2), as well as newer operations in Africa (7) and Asia (3). This survey represents 2018 copper cathode production of 1.91 Mt, or about 43% of world production by solvent extraction combined with electrowinning (SX–EW). Trends in operating conditions and performance, as well as equipment design and process technology choices are analysed. The significant expansion in production in the African Copper Belt is also discussed. There are currently about 130 operations worldwide producing some 4.6 Mt copper by SX–EW; of these, about half produced more than 10 000 t/a. Figure 1 shows global copper production by country. As it has for several decades, Chile remains dominant, with 36% of global production of cathode copper. Since 2005, the Democratic Republic of Congo (DRC) has increased production year on year, and now represents 23% of the global total (International Copper Study Group, 2017; Statistica, 2018). Copper production in Asia has grown significantly in the past decade, particularly since the ramp-up of Wanbao in Myanmar. There used to be numerous small copper SX plants in China (Schlesinger, King, Sole, & Davenport, 2011), although most of these have now shut down and Chinese production is lower than it was five years ago. The most important trend, however, is the rise of Chinese-owned mining companies, particularly in Central Africa, but also around the world more generally.

Jan 1, 2019

By A. Hartikainen, M. Rinne, F. Johansson, R. Yorke, L K. T. Uotinen, E. Korpi, J. Antikainen

"In order to determine the in-situ shear strength of rock joints, large scale testing is required. However, this is both expensive and difficult to execute. One possible method to overcome this may be to use photogrammetry to capture large joint surface roughness in-situ and downscale it to replica samples, which could be sheared in laboratory. In this paper, as a first part in such a method, a technique to digitize surface roughness and to produce replica samples for laboratory shear testing from a larger joint sample are presented. First, a thin granitic rock slice with dimensions of 1.75 m x 0.95 m of granitic intact rock was chosen for the study. The joint surface is fresh and created through tensile induced splitting. The large joint sample is digitized using photogrammetry. Then, one fullscale 1.7 m x 0.6 m geometry is cropped from the digitized joint geometry and then subsamples at 10x, 7.5x, 5x, 2.5x and 1x scales. All sub-geometries are scaled down digitally to produce 0.17 m by 0.06 m geometries. The geometries are used to make casting molds both positive and negative to produce samples with perfect matedness. The casting molds are 3D printed in polylactic acid plastic and C60/75 concrete is cast to produce a replica series. In addition to the creation of this replica series, two pilot replicas are also tested using a portable shear box with a 0.5 MPa normal pressure. The results from the pilot rounds are presented and discussed. Finally, suggestions for future research are given.INTRODUCTIONThe rock mass strength is in many cases controlled by the shear strength of rock joints, especially in hard rock masses. The shear strength is generally found to be influenced by several parameters, including normal stress, the basic friction angle of the intact rock, the joint surface roughness, the joint wall compressive strength, the matedness of the joint surfaces, weathering and infilling. The effect of these parameters for the joint shear strength can be evaluated in laboratory scale. However, the shear strength is known to be scale dependent. How the scale effect influences the strength is not fully understood. The general consensus is that the shear strength decreases when sample size increases, i.e. a negative scale effect (Bandis et al. 1981). However, some studies have presented contradictory results, see for example Hencher et al. (1993) and Kutter & Otto (1990). The extent and nature of the scale effect is still a question of debate (Tatone & Grasselli 2013)."

Jan 1, 2015

By R. M. Graves, C. B. McKenzie, D. Barbour, D. W. Desnoyers

"The prospective Tulks volcanic belt hosts five significant massive sulfide deposits, Tulks Prospect, Tulks East, Jack's Pond, Daniel's Pond and Victoria Mine, as well as numerous showings and indications throughout its 70 km outcrop length. Regionally, the 498 Ma sequence consists dominantly of felsic pyroclastic rocks and flows with interbedded mafic volcanic units and finegrained sediments, commonly graphitic. Contemporaneous granitoid rocks occur in the south half of the belt. Rocks usually display a strong NE-SW, steeply dipping, penetrative foliation and other shear-related high-strain fabrics. Facing directions are normally northwest, although there is local evidence of folding. Sulfide bodies and their alteration systems are deformed and shape-modified in the style of ambient strain.The Tulks, Tulks East and Jack's Pond prospects appear to occupy a similar stratigraphic level, on or near the contact between a thick felsic pile and an overlying broad, conductive package of graphitic sedimentary rocks. All are pyritic deposits with well-developed footwall alteration systems. The Tulks prospect comprises four polymetallic lenses totalling 750 000 t. Three lenses constituting more than 6 000 000 t of sulfide mineralization are indicated at Tulks East but only the 210 000 t ""B"" lens contains significant base metals. Jack's Pond consists of a very extensive, zoned alteration complex in which four lenses of massive to semi-massive pyrite + chalcopyrite are known. The central part of the system, characterized by Mg chlorite and carbonate, is enveloped by pervasive sericitization and capped by intensely silicified and pyritized volcanics. The sulfide lenses are interpreted as part of the footwall zone and are beneath that part of the system most likely to hostpolymetallic massive sulfides.Daniel's Pond differs from the other prospects in such fundamental ways as metal content, associated alteration and structural setting. The deposit comprises strongly tectonized, Ag-rich, Zn-Pb mineralization distinguished by a pronounced Ba-As-Sb signature. Mineralization is confined to a narrow, north-south high-strain zone which separates unstrained mafic volcanic rocks to the east from well-foliated felsic rocks. The deposit lacks an extensive alteration halo and at its north end is associated with a lens of barren, undeformed massive pyrite.The Victoria Mine deposit, the most cupriferous of the five, is related to a moderately dipping structure that separates well-foliated Tulks volcanic rocks from an overlying, less deformed volcanic sedimentary sequence that may be as young as 462 Ma. The structure is also the locus of black chlorite, dolomite and quartz alteration. Structurally remobilized precursor sulfide mineralization is one possible source for the deposit."

Jan 1, 1993

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 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. 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 G. Kordosky

The experience of one plant regarding degradation of copper solvent extraction reagents in the presence of 20 to 30 g/1 nitrate in the leach solution was described at Cobre 2003. However, little information is available about the degradation of copper solvent extraction reagents where the nitrate concentration of leach solution is much lower. Analysis of plant organic samples taken from an operating copper SX plant over a long period of time shows that small amounts of nitrate in the leach liquor can be a significant contributor to reagent degradation provided certain other conditions are present. These conditions are discussed and cautionary suggestions on copper SX plant operations in the presence of even low levels of nitrate in the pregnant leach are made based on industry experience with copper solvent extraction plants and copper solvent extraction reagents.

Jan 1, 2007

By D. Zagidulin

Short term electrodeposition experiments (up to 80 min.) at two current densities 220 and 275 A/m2 with varying concentrations of thiourea and glue were performed to determine the optimum levels of glue, thiourea and current density for industrial copper electrorefining. White Light Interference Microscopy was used to acquire digital images of the morphology of the electrodeposited copper. Scaling analysis was employed to pa-rameterize the morphological information encoded in the images. The limiting rough¬ness, *, the critical scaling length, Lc, and the aspect ratio, 4*/ Lc, were determined as a function of the deposition time and the amount of organic additives. These parameters were used to calculate the dynamic exponent, B, the roughness exponent, 1/z, and the as¬pect ratio exponent, y, plotted as a function of glue and thiourea concentrations. These 3D plots allowed us to find conditions corresponding to the minimum of B and y, at which the deposited copper is characterized by a small roughness and a dense structure.

Jan 1, 2007

By G. B. Harris

A new process for the recovery of copper and associated base metals from a polymetallic massive sulphide ore is described. The massive sulphide contains copper and especially nickel sulphide in a mineralogical form which negates both good recovery and high grades in concentrates using conventional flotation methods. Flotation concentrate processing and smelting from this ore is therefore probably not an economically viable option. Utilising the properties of high concentration chloride brines, a flow sheet has been developed and tested whereby the recovery of all of the metals and some of the power requirements for the plant itself can be derived from the ore. The new process overcomes the less-amenable-to-flotation ore characteristics, and provides the opportunity of economic processing in remote locations where the dependency on fossil fuels is both a logistical and an economic concern. The Ferguson Lake deposit of Starfield Resources in Nunavut, Canada, is the subject of this paper, which focuses on the leaching and copper recovery unit operations of the proposed flow sheet.

Jan 1, 2007

By Shoma Sinha, Nirmal K. Sinha

Deformation in a material at any temperature is known to consist of two major components - a reversible (elastic) part and an irreversible part (often called plastic). What is not universally known is the fact that at high temperatures greater than about 0.4Tm, where Tm, is the melting point in Kelvin, the reversible part includes not only elastic but also a delayed elastic (recoverable but time dependent) component. A very simple computational model is presented here for a constant-strain stress relaxation test (SRT) on the basis of a three-component constant-stress rheological equation derived from strain relaxation and recovery test (SRRT) data. It is shown that the rapid stress reduction during the early periods of relaxation when the stresses are high is governed primarily by delayed elasticity irrespective of the imposed strain level. This is in contradiction to the popular concept prevailing today. The asymptotic approach to a quasi-stable stress rate at longer times and lower stresses is, however, controlled by viscous flow (power-law creep). The validity of the theory is examined by comparing the predictions with experimental results obtained for a titanium-base alloy Ti-6246 at 0.45Tm.

Jan 1, 2004

By M. Okada

Consumption of refined copper in Asia, especially in China and India, has soared, and the copper price is at its historic height. New smelters and refineries have been built in the region with state-of-the-art technology to meet the unprecedented demands. Supply of copper concentrate, on the other hand, has fallen short of the growing demand for concentrate by the smelters and refineries with expanded production capacities. The shortage of copper concentrate is attributable to the delay in the mining development, caused by the prolonged price slump of before 2005. The mismatch between supply of concentrate and smelting and refining capacities may cause disorderly movement of copper price, and hamper the sound development of the industry. The recent high copper price may prompt substitution with alternate materials. With electrical and thermal conductivity and workability, copper is one of the most customer-friendly materials. It is regrettable for both customers and suppliers if the high copper price would become an impediment to promoting copper use. Some outdated and environmentally unfriendly facilities still operate. In view of recent lower-grade copper concentrate with higher sulfur content, these facilities must reform their equipment to environmental-responsive one. The current level of per capita consumption of refined copper in Asia shows that the demand of copper will further grow. It is necessary to realize stable supply of refined copper and stable copper price, to help developing countries including China and India to economically take off. Hence, it is essential that the development of the copper industry be accompanied by environmentally friendly facilities, and harmoniously increased capacities between mining and smelting and refining.

Jan 1, 2007

By Robert B. Parsons

"As the 1981 annual reports of Canadian mining companies one by one see the light of day, it becomes clear that softer metal prices and rising costs are the main problems which beset the industry today. These problems are certainly not lightened by that familiar monster lurking in the shadows-taxation! Recent taxation trends in Canada, and the existing tax regimes for the mining industry, continue to be a major bugaboo.This examination by ""Tax Notes"" of the impact of current taxation trends on the mining industry is based on a speech delivered recently by Robert D. Brown, senior tax partner of Price Waterhouse, Canada, to the annual meeting of the American Institute of Mining, Metallurgical and Petroleum Engineers in Dallas.IntroductionThe purpose of this paper is to tell a story-and it is a story with a moral. The first thing to recognize is that, with depressed metal prices and weakening western economies, high taxes, in and of themselves, are not really the cause of the Canadian mining industry's present doldrums. However, the character and level of mining taxes in Canada has had a profound influence on the development of the industry over the past decade, and the expectations of what the future holds for mining men in Canada will equally play an important role in the future development of the Canadian industry. Our tax system has, in the past, served as a negative and destabilizing influence in the mining industry: it has contributed to a major short-fall in investment commitments and hence in the contribution which the mining industry can make to the Canadian economy, and the adverse effect on the industry of previous high marginal tax rates was made far worse by the confusing, contradictory and perverse nature of the tax system itself."

Jan 1, 1982

By P. M. STILES

Acquired in 1957, the limestone quarry of Ideal Basic Industries on Texada Island north of Vancouver has become a major supplier of high purity limestone and construction stone. Crushed limestone products are shipped by barge to various destinations along the B.C. and northern U.S. coasts.

Jan 1, 1984

By D. M. Baird

Copper mineralization occurs at several places along the shore of Notre Dame bay, Newfoundland, and at one time formed the basis of a considerable mining industry. The sources of the mineralizing solutions and the structural settings of some of the deposits have not been clearly understood. This paper reports new investigations into the origin and structure of the deposits between Bett's cove and Stocking harbour, in a strip about ten miles long and two miles wide along the western shore of Notre Dame bay. The most complete previous descriptions of the area are those of Snelgrove (1929, 1931, 1938), Douglas, Williams and Rove (1940), and the writer (1947). The discovery of the Bett's Cove deposits took place about 1860, soon after the discovery of the well-known Tilt Cove deposits to the north. Several other small prospects were found between Bett's cove and Stocking harbour between 1880 and 1910. At Bett's cove, active mining began in 1875 and a small smelting plant was built near the mine in 1877. .Mining operations came to an abrupt halt in 1883 after a major cave-in. In 1900, and again in 1906, attempts were made to drain the old workings and reopen the mine but were abandoned long before completion. Total production amounted to about 130,000 tons of hand-picked copper ore which averaged about ten per cent copper, and 2,450 tons of pyrite (Snelgrove, 1938, pp. 52-55).

Jan 1, 1948