Search Documents

By D. Rahal, I. Shah, B. Hines, S. Reddick

"The development of stirred milling technology has proven to be an enabling technology in fine grained metallic deposits. The increased grinding efficiency allows for the economic size reduction required to produce saleable concentrates. This improvement has opened the door to retreatment projects such as reprocessing of historic gold mine tailings. Recently a VXP2500 stirred mill was installed at Casmyn Mining Zimbabwe Turk mine to allow the reprocessing of historic mine tailings. After commissioning, an extensive joint research project between Casmyn and FLSmidth Minerals was conducted to optimize the performance of the mill. This paper outlines the challenges encountered in the optimization program and the outcomes from the joint research project.INTRODUCTION The development of stirred mills has opened up many new opportunities to recover valuable minerals from refractory ores. Stirred mills have already been reported as an enabling technology for the processing of fine grained ores in lead/zinc applications (Pease, Young and Curry, 2005). Historically, conventional ball mills were unable to economically grind material fine enough to produce the required recovery or concentrate grade. This often resulted in the loss of valuable materials to the final plant tailings. The advent of stirred milling has now provided an opportunity which has both economic and environmental benefits, the reprocessing of historic mine tailings. Many tailings stockpiles now contain economic concentrations of gold because of the ability of stirred mills to efficiently and economically grind the tailings to liberation sizes. The primary advantage of reprocessing historic tailings is very low cost processing. Recovering the tailings is usually as simple as pumping the material into the reprocessing circuit. The disadvantage of reprocessing tailings is the low feed grade and greater difficulty in recovery valuable minerals at the same size distribution. The material has already been previously processed so fresh surface area or an increase in metal exposure is required. These are both provided by regrinding the tailings material."

Jan 1, 2014

By Joseph U. Messenger

A method of applying ammonium hydroxide to a drill stem containing hydrogen sulphide gas and hydro-gen-sulphide-bearing oil and water so as to remove the hydrogen sulphide as ammonium sulphide is de-scribed. It was concluded that the method allowed trips with drill pipe filled with sour fluids to be made more safely, more rapidly, and more economically. INTRODUCTION THE DANGER of poisoning rig crews during drill stem tests of sour wells by exposing them to hydrogen sulphide gas is well known in Canada. Concentrations of hydro-gen sulphide in sour gases range from 0 to 40 volume percent. There have been numerous casualties from hydrogen sulphide exposure during drill stem testing. Some have resulted in permanent injuries; in some cases the injuries have been fatal. For example, in Alberta in 1957 there were four deaths and 100 exposures which required medical treatment. These fatalities and injuries indicate the need for a more effective way of handling hydrogen sulphide in oilfield operations.

Jan 1, 1958

By P. Kostuik

"Recent interest in the pipeline transport of solid mate-als has been sparked by the idea that, in an inflationary onomy, pipeline costs are more stable than the costs of any other transport systems. Pipelines have proved over 1e years to be dependable, records indicate that main-•nance and operating costs are comparatively low, and 1eir amenability to automation is of considerable im->rtance.The concept of hydraulic hoisting of solids from under-round, or cross-country hydraulic transportation of con-~ntrates and ores, is currently of prime interest to the ining industry. In general, this paper deals with a preliminary investi-ation of energy losses incurred while transporting a rela-vely heavy, crushed mine product, namely magnetite, oth vertically and horizontally. The basis and validity of everal equations derived by other experimenters to com-ute these losses is discussed. A theoretical analysis of ydraulic hoi:,,ting is made, and, based upon test results, calculation of horsepower requirements for hoisting an re under a given set of conditions is presented. Results indicate that:n vertical transport, the friction head losses are identical to those of clear water, and the total energy requirement for hydraulic hoisting then becomes a function of height of lift, concentration and specific gravity of the solids being hoisted. .-In horizontal transport, friction head losses can be calculated with reasonable accuracy using Bonning-ton's correlation curve. Losses computed utilizing Durand's correlation (K = 85) result in estimates which are too low by 10 to 30 per cent, an<l those calculated using Newitt's correlation (K = 1,100, K = 66) give friction loss values which are too high by 10 to 30 per cent . . -The theoretical horsepower (RMS-hp.) required for hoisting hydraulically is approximately half of that required for conventional double-drum skip hoisting."

Jan 1, 1966

By K. Liu

"In the present work, two-step heat treatments with preheating treatments at different temperatures (175, 250 and 330°C) as the first step followed by the peak precipitation treatment (375°C/48h) as the second step have been performed in Al-Mn-Mg 3004 alloys to study their effects on the formation of dispersoids and evolution of elevated-temperature strength and creep resistance. During the two-step heat treatments, the microhardness is gradually increasing with increasing time until to a plateau after 24 hours when first-treated at 250°C and 330°C while there is a minor decrease with time at 175°C. Results show that both the yield strength and creep resistance at 300°C reach the peak values after the two-step treatment of 250°C/24h+375°C/48h. The formation of dispersoids is greatly related to the type and size of pre-existing Mg2Si precipitated during the preheating treatments, in which that coarsened rod-like ß’-Mg2Si strongly promotes the nucleation of dispersoids while finer needle-like ß’’-Mg2Si has less influence. Under optimized two-step heat treatments and modified alloying elements, the yield strength at 300°C can reach as high as 97 MPa with the minimum creep rate of 2.2×10-9 s-1 at 300°C in Al-Mn-Mg 3004 alloys, enabling them as one promising candidate in lightweight aluminum alloys for elevated-temperature applications.INTRODUCTION Due to the rapid demand from weight-sensitive automotive and aerospace industrials for the light alloys, such as aluminum alloy on the elevated-temperature applications, Al-Mn 3xxx alloys has been developed to obtain the good properties at both room temperature (RT) and elevated temperature, in which the dispersoid-strengthening mechanism plays a significant role (Li & Arnberg, 2003; Li, Muggerud, Olsen & Furu, 2012; Liu & Chen, 2015, 2015). In our previous works (Liu & Chen, 2015, 2015; Liu, Ma & Chen, 2017), the yield strength and creep resistance at 300°C of Al-Mn-Mg 3004 alloy were further improved by modifying the alloying elements, such as Fe and Mo, to optimize the characters of dispersoids, including the size, distribution and the dispersoids free zone (DFZ). For instance, Mo was introduced in aluminum alloy to further improve the volume fraction of dispersoids and reduce the area of DFZ, leading to increasing on both the strength and creep resistance at 300°C (Liu, Ma & Chen, 2017). On the other hand, the mechanical properties became worse at higher Fe content due to its consumption more Mn to form Al6(MnFe) intermetallics, resulting in less available soluted Mn for the precipitation of dispersoids (Liu & Chen, 2015). Therefore, optimizing the characters of dispersoids is always the key to improve the elevated-temperature properties of dispersoid-strengthened aluminum alloys for elevated-temperature applications."

Jan 1, 2017

By C. A. Pickles, D. K. Xia

"In the last few decades, microwave energy has been widely employed in food processing, rubber and plastics curing, and in ceramic sintering. Recently, the application of micro waves as an energy source in extractive metallurgy has received considerable attention from a number of researchers. During the period from the mid-1960s to about 1988, the majority of the research was focussed on the behaviour of materials in microwave radiation. These research activities will continue to be important, since fundamental data on the interaction of microwaves with materials are vital for the successful application of the technology. Since1990, a number of potential applications of microwave processing have been investigated.These include: pretreatment of ores and concentrates, leaching, reduction, preheating of refractories, heating of slags and waste treatment. In this paper, the interaction of micro -waves with materials of interest in extractive metallurgy are reviewed and potential applications are discussed.IntroductionMicrowaves were first used in 1946 for the heating of materials, shortly after radar equipment was invented during the second world war. It was evident that microwaves were very effective for heating water and the technology was quickly adopted for drying and cooking. Since the introduction of the first microwave oven by Raytheon in 1952, many millions of domestic microwave ovens have been sold annually (National Research Council,1994; Atomic Energy of Canada and Voss Associates Engineering, 1990).In extractive metallurgy, microwaves are in the early stages of development (National Research Council, 1994; Atomic Energy of Canada and Voss Associates, 1990; McGill and Walkiewicz, 1987; Bradhurst and Worner, 1990;Peng et al., 1993). However, there are several reasons for the growing interest in the use of this technology in the field of metals extraction. These include the following potential advantages of microwaves: (1) rapid and selective heating of materials; (2) reactions can be catalyzed because the heating occurs on a molecular or atomic level; (3) clean and controllable energy source; (4) the gas volume is reduced and the atmosphere can be controlled as there are no gaseous combustion products; (5) the material is heated internally in comparison to external heating with conventional methods; (6)the temperature of the refractory can be minimized; and (7) the working environment is improved."

Jan 1, 1997

By M. Girard, E. Bourricaudy, M. Ashbury, M. Canizares, A. Mezei

Liquid-solid separation and slurry handleability relate directly to commercial plant ramp-up success or lack thereof. The paper provides a well-proven approach for testing, scale-up and commercial application of difficult processes. Emphasis is placed on practical applications of rheological concepts developed both academically and empirically. The cause-effect relationship between the process chemistry and possible physical implications on the metallurgical plant operability is presented in a simple, direct form. Examples include recent results in areas of mineral processing and extractive metallurgy. Technical process risk factors and their mitigation are included in the discussion.

Jan 1, 2011

By Ron Light, Doug Applett

"HistoryThe Nanisivik mine began operation in October 1976 as the world's most northerly metal mine and the first permanent industrial project in the Canadian High Arctic. Located 700 km north of the Arctic Circle on Baffin Island's Strathcona Sound, exploration by Texasgulf Inc. between 1957 and 1970 had outlined the ore deposit. In 1974, Nanisivik Mines Ltd. was formed with Mineral Resources International, the Government of Canada, Metallgesellschaft and Billiton being shareholders, and Strathcona Mineral Services Ltd. providing project management. In 1996, Breakwater Resources purchased the mine from Alberta Energy Corporation Limited which had recently acquired the assets of then owners Conwest Exploration Ltd.Ore DepositThe main ore deposit is very flat, approximately 80 m in width and varying in thickness from 3 m to 16 m. The length of the deposit is approximately 3000 m with satellite zones located on the perimeter of the main zone. The economic mineralization is sphalerite with minor galena and silver. The sphalerite and galena are coarse grained and associated with pyrite as the main gangue component with minor inclusions pf dolomite and calcite within the deposit. The deposit being hosted in a permafi·ost environment creates a number of challenges. Dry drilling is employed in room and pillar production drilling. The permafrost aids in the stability of both the pillars and the back."

Jan 1, 2000

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 C. Stamberg, W. McGuire, R. C. Carlson, L. W. Vollmer

"The selection of materials and pre-cautionary measures for handling natural gas and condensate containing aqueous hydrogen sulphide are presented. Based on the results of a pilot production test and laboratory studies, together with several years . of operating experience, it was found that the behaviour of materials in sour gas environments can be predicted sufficiently accurately for engineering purposes.The selection of proper materials is essential for preventing sulphide-stress cracking. Where high strength materials are not required, low car-bon steel is used. Components requiring high tensile strength and hardness characteristics should be made from K-monel or properly heat-treated carbon-manganese-molybden-um and AISI 4140 steels. Welded connections are a potential source of sulphide-stress cracking. Consequently high quality welding and strict inspection are suggested as the most economical precautionary measures available.A maintenance program based,. where possible, on visual inspection of equipment and/or test spools located at strategic points should be used to follow weight loss and putting corrosion in sour gas environments. If severe weight loss or pitting corrosion is encountered, inhibitors have been found to be an effective deter-rent."

Jan 1, 1960

By William Petruk

Image analysis studies were performed on mill products from pilot-plant and laboratory tests that had been conducted to determine the effect of grinding media on the selective flotation of two zinc-lead-copper ores. It was found that, when the ores had been ground to the same degree of fineness with flint pebbles, ceramic pebbles or ore fragments (autogenous grinding) as with steel balls, there generally was increased sphalerite, chalcopyrite and galena liberation in the flotation feeds, depending on the textures of the minerals. The increased sphalerite liberation fully accounted for the improved zinc recoveries that were obtained when steel balls were replaced by either pebbles or ore fragments. The increased chalcopyrite and galena liberation accounted for part of the changed technical efficiency of concentration, but other factors, probably changes in oxidation and reduction of iron and in consumption of oxygen, also contributed.

Jan 1, 1977

By Tim Hughes, Steve Gannon, Jennifer Abols

"In the past year, two very innovative gold ore processing plants have been designed, built and commissioned by Gekko Systems Pty Ltd. The first plant in Vietnam utilizes conventional crushing and ball milling followed by gravity concentration (Inline Pressure Jig (IPJ) and Falcon concentrator), flotation, intensive leaching (GFIL), the first industrial scale use of AuRIX®1001 ion exchange resin and a cyanide destruction circuit. The second plant, in Australia, uses a 3 stage crushing circuit to crush the gold ore to minus 1mm, followed by gravity concentration utilising a rougher, scavenger, cleaner circuit of IPJ’s and a Falcon concentrator. A direct smeltable gold concentrate is produced as well as a lower grade concentrate for future intensive leaching or sale. Both plants were designed and built within very tight time frames and budgetary constraints. The experience gained during the plants’ design, commissioning and operation are discussed.INTRODUCTIONGekko Systems (Gekko) is a leader in the development of innovative technology for gravity recovery and intensive cyanidation. It provides low cost and metallurgically optimal solutions for its customers. It has also led the trend away from whole ore cyanidation (CIP/CIL) towards gravity/intensive cyanidation (GIL) and gravity/flotation/intensive cyanidation (GFIL) plants. These types of plants offer low capital costs, low environmental impact and maximum return to investors (Rob J Longley et al, 2002). Key to this type of flowsheet is the application of a continuous concentrate unit, such as the InLine Pressure Jig (IPJ), which is designed to produce high gold gravity recoveries at a coarse grind (Sandy Gray, 2003). When combined with other gravity devices, and/or flotation, and intensive cyanidation, using the Gekko InLine Leach Reactor (ILR), very high overall recoveries can be achieved.In 2004, Gekko launched a new range of products called Gekko Modular Systems. The idea was to provide off the shelf, complete solutions for the separation and recovery of complex and free gold as well as the pre-concentration of diamonds. A series of modules were designed that incorporated Gekko manufactured as well as outsourced equipment. Gekko now offers modules ranging from feeding and conveying to crushing, gravity recovery, flotation, intensive cyanidation, refining and detox (see Figure 1)."

Jan 1, 2007

By K. Peake, A. Dalvi, V. Hernandez, S. Chang, M. Oliazadeh

"Ironstone Resources Ltd. is developing the Clear Hills oolitic iron ore deposit in northwest Alberta, Canada. Sources of oolitic iron ore worldwide are abundant and ubiquitous, but there is currently no commercial technology to extract iron from oolitic iron ores. Mineral processing of the ore to generate a commercial iron ore concentrate is very challenging due to the nature of these ores. Accordingly, Hatch and Ironstone are developing a new pyrometallurgical process (Hatch Ironstone Chloride Segregation) to produce metallic iron from high silica and high phosphorus oolitic iron ores. Following the chloride segregation process, the treated ore is subjected to fine grinding. The iron is then separated by magnetic separation. The iron grade and phosphorus content of the final product shows a strong dependence on the separation methodology and the grind size. The pyrometallurgical process has been developed and tested at a lab-scale pilot plant. The commercialization of this technology could provide an opportunity for the development of a new alternate iron making technology. This paper presents the results of test work done on the raw iron ore to create a concentrate and also on the products of the pyrometallurgical process for the production of a high grade 90% metallic iron product.INTRODUCTION Oolite is a sedimentary rock formed from ooids – spherical particles with concentric layers. There are a number of large oolitic iron deposits where the mineralogical form features a low iron content and high concentration of impurities making them presently impossible to exploit on either a technical or economic basis. The fifteen largest oolitic iron ore deposits in the world are estimated to contain approximately 337 billion tonnes of iron ore, which would be a significant source of supply for worldwide steel production. A viable process that can economically recover iron from these oolitic iron ore deposits could become an important new technology to meet future demand in the steel industry. Hatch and Ironstone have demonstrated that oolitic iron ores can be exploited to produce direct reduced iron (DRI). Direct reduction ironmaking is a method for producing solid metallic iron using different techniques compared to the blast furnace process, which is the more conventional ironmaking route. The demand for the construction of direct reduction ironmaking plants has increased in response to the increasing demand for crude steel. It is viewed as a method of providing low-residual iron sources to scrap-based steelmaking in electric arc furnaces."

Jan 1, 2014

By L. H. Van Loon

INTRODUCTION THE property of Madsen Red Lake Gold Mines, Limited lies between Russet and Faulkenham lakes in Baird and Heyson townships, district of Patricia, Ontario. It is about six miles south west of the town of Red Lake, with which it is connected by an all-weather road. All orebodies found to date at the property are silicified lenticular bodies which lie within tuff beds. The ore bas a specific gravity of 2.95. It consists of mica and chlorite schist, silicified and containing a considerable amount of disseminated carbonate, and mineralized with pyrite, pyrrhotite, magnetite, arsenopyrite, and chalcopyrite. Of these, pyrite, in coarse to fine disseminated grains, is by far the most abundant. Pyrrhotite occurs in ?the gangue material as medium to fine disseminated grains. A very small amount of arsenopyrite, in small crystals, is associated with the pyrite, and small grains of chalcopyrite occur in both gangue and pyrite. Although some of the gold present in the ore is quite coarse, a good deal of it is extremely fine.

Jan 1, 1951

By Andrew Artzer, Michael Moats, Jack Bender

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

Jan 1, 2019

By Wei Sun, Xiangsong Meng, Xiaodong Li, Weisheng Huang, Yuehua Hu, Chonggao Cao, Haisheng Han, Wensheng Chen, Runqing Liu, Central South University

"The similar flotability of scheelite, fluorite, and calcite resulted in the difficult separation by flotation. This work revealed the effect of water glass on the flotation of scheelite by FTIR and the wettability change of the usual cleavage plane, and proved that water glass inevitably influenced the flotation of tungsten minerals to a large extend in traditional fatty acid flotation process. A new collector, benzohydroxamic acid complexes, was found to be selective for the separation of scheelite and calcium minerals with little water glass. Hence a novel flotation process was developed for the recovery of tungsten minerals in shizhuyuan mine in China. Normal temperature flotation of tungsten minerals was developed to instead of the classical “Petrov's process” for scheelite-calcite-fluorite type ores. Compared with the traditional process, the recovery was improved by 10% and the grade of scheelite concentrate was above 55%. The disappearance or decreasing of water glass contributed to improving the recovery of tungsten and fluorite and the circulation of water and reagents.INTRODUCTIONChina is rich in tungsten resources and the proved reserves accounted for 60% of the world's reserves (Cai Gaipin et al., 2009). However the wolframite has been dried up and the scheelite resource has become the main resource of the tungsten metal. Scheelite deposits were classified to scheelite-silicate type and scheelite-calcite-fluorite type based on the associated gangue minerals. Scheelite was easy to separate from the silicate minerals. Flotation separation of scheelite from calcium minerals creates difficulties because of the existence of the same cation in the minerals and similar physicochemical characteristics such as solubility, hardness, specific gravity, PZC (point of zero charge) etc.(Arnold & Warren, 1974; Yongxin & Changgen, 1983; Ozcan & Bulutcu, 1993a; Deng et al., 2015, 2016;). The conventional collectors used in the flotation of scheelite ores are generally fatty acids or fatty acid derivatives ( Yongxin & Changgen, 1983; Oliveira & Sampaio, 1988; Sivamohan, 1988; Shi et al., 2014; Bo et al., 2015). It is difficult to separate scheelite especially from calcite with these reagents because the adsorption mechanism of these collectors occurs through chemisorption of the oleate ion onto the mineral surface (Atademir et al., 1979; Changgen and Yongxin, 1983; Hiçyìlmaz et al., 1993; Hu et al., 2011). Therefore, it is almost impossible to separate them by fatty acid flotation without using any selective depressant. Although oleoyl sarcosine was used as a selective collector in the flotation of fluorite and siliceous scheelite ores (Ozcan & Bulutcu, 1993b; Ozcan et al., 1994), this selectivity is not adequate for a low-grade calcitic scheelite ore. Unfortunately, the reagents such as organic colloids (dextrine, starch), quebracho, sodium phosphates, sodium silicate solutions containing polyvalent cations (hydrosols), etc. (Yongxin and Changgen, 1983; Schubert et al., 1990; Shi et al., 2014; Bo et al., 2015) used for calcite depression also depress scheelite to a large extent."

Jan 1, 2016

By J. J. Brummer

The possibility of diamonds occurring in Canada was first raised by W. H. Hobbs in 1899, who concluded: "that the apex of the fan of diamond distribution probably lies somewhere in the strip of the territory bordering James Bay on the east." This was based on the occurrence of diamonds in surficial sediments at various locations south of the Great Lakes. There are two recorded discoveries of drift diamonds in Ontario: (a) the Peterborough Diamond, 33 carats in size, poorly documented and found sometime before 1920;and (b) the Jarvi Diamond, 0.255 carat, found during 1971 in Sheraton Township near Timmins . Both discoveries were in surficial sediments and their source(s) have not been located. During 1960, Selco Exploration Company Ltd. started a heavy-mineral sample program of river gravels in the James Bay Lowlands of Ontario. This was continued by Canadian Rock Company Ltd. (1962-1963) and the Ontario Division of Mines (1973). The surveys discovered the presence of kimberlite minerals in gravel samples from streams draining the Moose River Basin. A total of 53 pyrope grains, four magnesian ilmenites and one grain of chrome diopside were identified from 32 sample locations. A study of the distribution of these minerals suggested the kimberlite source rocks to be located along the north margin or within the Kapuskasing gravity "high" in the area immediately south of James Bay. The Jarvi Diamond, found on an esker east of Timmins, could well have originated from this area. During 1965, an esker was sampled by the Geological Survey of Canada, and the results attracted attention to the area south of Lake Abitibi. At 9 out of 34 sites sampled along the Munro esker, Lee recorded the presence of pyrope garnets and very minor chrome diopside.

Jan 1, 1978

By Brad Thrall

"Viceroy Resource Corporation of Vancouver, B.C. has recently developed the Brewery Creek Mine, an open pit heap leach gold mine 55km East of Dawson City in Yukon, Canada. Mineable reserves are 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797000oz) of gold. Ore processing employs the heap leach technology using run of mine ore. 1996 highlights included construction of 158,000 m2 of leach pad, placing 1.9 million tonnes of ore at a grade of 1.46 gpt on to the leach pad, construction and successful commissioning of the ADR plant, and pouring of the first gold bar on November 15th.Permitting this mine has drawn significant attention because the heap leach technology, while proven in other jurisdictions, has not been used in Yukon where winter site temperatures routinely reach -45 C for extended periods.IntroductionFollowing a comprehensive governmental and public review of the proposed project, a Class A Yukon Water license was issued in August, 1995. Construction started in 1995 and leaching commenced in October 1996. The Brewery Creek Mine is in Yukon, the home of the Klondike Gold Rush nearly 100 years ago. This is the only part of Canada not glaciated in the Wisconsin advance so there is oxide ore resulting from about 80,000 years of weathering -unique in Canada. Discontinuous permafrost is present on site and has required special consideration in the design. Winter temperatures dip below minus 50 C, demanding an approach to mining and leaching that is different than in more temperate climates.Project DescriptionThe mine comprises 8 shallow, largely oxidized gold deposits along a 7km corridor. An 8 year mining program will extract 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797,000 oz) of gold. The strip ratio will be 1.35: 1. Overall gold recovery is projected at 78%. There will be 2.3 million tonnes of ore placed on the pad annually, with an average of 2.57 tonnes (83, 000 oz) of gold recovered per year. Because of the cold winters, the mine will operate approximately 210 days per year, while the leaching process will operate on a year round basis. Run of mine ore is placed on the heap 150 days per year. The recovery process utilizes conventional sodium cyanide leaching , carbon adsorption, caustic desorption, and on-site smelting. Peak employment during mining is 145 and winter staff is reduced to 45."

Jan 1, 1997

By P. E. Cavanagh

Introduction A comprehensive study of the economics of processes for smelting iron ores has been made by the Ontario Research Foundation on behalf of the Ontario Research Commission. This survey has included investigations of all processes in commercial use, and proposed processes which are still in the experimental stage. The two main objects of this work were to discover (l) whether any known processes would be particularly suited to present and future Canadian conditions, and (2) whether any new process would be of assistance in solving the present scrap shortage in Canada. The future prospects of the Canadian iron smelting industry have changed radically in the past ten years. For nearly twenty years before 1939 no iron ore was mined in Canada. The ore used was imported, mainly from the United States and Newfoundland. The economics of iron blast-furnace smelting always force the location of blast furnaces near the source of coal for making high-grade metallurgical coke, and near the market for iron and steel, rather than near the iron ore. Blast-furnace installations in Canada are confined to the north shores of the Great Lakes with ready access to United States coal and ore and the major Canadian markets, and to Nova Scotia near the coalfields of that Province. This distribution of primary smelting capacity naturally results in relatively high prices for iron and steel in the West, the North, and in Quebec.

Jan 1, 1948

By A. D. Hunt

THIS PAPER will review Canada&apos;s fossil fuel situation as we understand it today and indicate some of the options available to meet our energy requirements during the next 15 years. It would therefore be helpful to examine Canada&apos;s total energy requirements and discuss the un certainties of the data base on which we must base future energy investment decisions. On the world scene, Canada is at the top of the energy-consuming nations, using about 49 bbls of oil equivalent per caput/ year. This appears to be slightly ahead of the U.S. and of course much higher than that of many countries. It is generally accepted that there is a more or less direct relationship between energy consumption and material standard of living. However , as we know, Canada, when compared with other industrialized countries on a GNP, an individual-income or some other standard basis, is lower in the list. This situation may be explained by our more severe climate and a relatively small population distributed over a land mass second only to that of the Soviet Union.

Jan 1, 1977

By R. F. Favreau, L. C. Lang

"In modern blasting technology blasts are designed and analysed on an energy-mass-time relationship. The energy of the explosive is derived by computer analysis and the work potential of the available energy is expressed in numerical values. The mass involved in the blast is determined by the geometry of the blast and by the rock density. Time is also a relevant parameter, because time is required to complete the three basic stages of the breakage process.For successful application of the blasting parameters the basic mechanism of the breakage process must be understood.The physical-mechanical properties of the rock, local structural geology and the thermodynamic properties of the explosives can be brought together to create a computer model of a blast with which it is possible to predict separately the results of each parameter and to express the predicted degree of change in any parameter by actual numbers.A long-range project has started to create a ""bank"" of rock properties for as many Canadian and foreign mines as possible. A computer model has been successfully applied to predicting the effect of varying the ratio of spacing to burden at constant rock volume and significantly improved results are achieved, as the S/B ratio is increased up to about 2 to 2.5.High-speed photography has been applied to determine the time that has elapsed between zero time and the beginning of movement of the mass. The results assist in determining minimum and optimum delay time required for multiple-row blasting.The approach as described in the paper is now well proven and is clearly superior to previous methods based on powder factors and human assessment."

Jan 1, 1972