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By S. E. Malouf, W. Thorpe

"The Chibougamau Explorers property comprises 35 claims or 1,726 acres in La Dauversiere and Rohault Townships along the west shore of Norhart Lake. The property is serviced by a twelve-mile all-weather high-way leaving the main St. Felicien-Chibougamau highway at mileage 121. This road junction will be a railway siding on the C.N.R. St. Felicien-Chibougamau branch line in early 1957. The property lies 30 miles due south from the town of Chibougamau. The mine-sponsored Chibougamau Explorers town has a population of 250 with 28 residences and a well equipped recreation centre.The original showing was found by Rider Storn, a prospector financed by Val d'Or interests during the 1950 field season. This showing did not stand up to detailed prospecting but it did lead to the discovery of the main zone by A. Oakley during the following winter.The company was incorporated in I 950 under the presidency of E. Ayrhart. Financing following the initial drilling involved an increase of capitalization from 3 to 5 million shares and the advent of financing and management control by Anacon Lead Mines Limited.Exploration and development to Jan. I, 1956, has included 42,331 feet of surface diamond drilling; 31,363 feet of underground drilling; a 3-compartment shaft to 1,245 feet, with 8 levels at 150-foot intervals; and I 1,054 feet of drifts, crosscuts, and raises. Ore reserves to the 900-foot level, proven and partly developed, total 546,725 tons averaging 0·30 ounces gold and 0·93 per cent copper. A mill with capacity of 500 to 750 tons will be installed by January 1956.GeologyThe Chibougamau Explorers property lies along the belt of Keewatin volcanics striking through Bachelor Lake, Opawica Lake, Doda Lake, Coapatina, Norhart Lake, and Rohault Lake, where it is in close proximity to the paragneisses of the Grenville series. The boundary of the paragneisses, a gradational contact, lies within a mile of the Chibougamau Explorers shaft collar."

Jan 1, 1949

By R W. Drybrough

The program to increase production at Sherritt Gordon's Lynn Lake property, which began in 1956, included increasing the rope speed and the horsepower of the ore hoist at the "A" shaft. Because of the limited power sup-ply from the two 7,000-hp. hydroelectric plants, and the ? 3,600 hp. predicted for the combined hoisting requirements of the three production shafts, it was considered desirable to lessen the impact of this type of load on the relatively small power system by equipping all future hoisting arrangements with Ward Leonard direct-current drives. The manually operated 8-by 4-ft. hoist at the "A" shaft had a rope speed of 1000 f.p.m., and was driven by a 385-hp., 380-r.p.m. wound rotor motor with liquid rheostat control. This A.C. drive was replaced by a semi-automatic Ward Leonard direct-current drive with a 515-hp., 566-r.p.m., shunt wound motor, providing a rope speed of 1,500 f.p.m. During several months preceding the con-version, all possible preparations were carried out to keep the changeover period to a minimum, and this resulted in a total shutdown time of twelve days. The capability of the "A" Shaft hoist was increased from 109 tons per hour to 134 tons per hour, and both mechanical and electrical maintenance were reduced considerably. The total cost of the conversion amounted to $130,000. The hoistmen who had been required for the manual A.C. drive were transferred to the new Farley shaft, and the reduction in operating labour costs of the "A" Shaft hoist

Jan 1, 1965

By R. K. Kandasami

"Naturally occurring geomaterials such as sandstone, limestone and other soft rocks, residual soils are composed of a particulates held together through a cohesive matrix, usually composed of argillaceous or siliceous cementing agents. Geo-materials such as these have been reconstituted in the laboratory using ordinary Portland cement in lieu of the naturally occurring cohesive matrix, in addition to using sand particulates. This cohesive frictional granular ensemble is an ideal model material for understanding the mechanical behaviour of transitional geomaterials such as soft rocks. An important consequence of this cohesion that pervades the inter particulate system is the enhanced tensile strength when compared to a purely granular ensemble. The assessment of strength of these soft rocks and cemented sands is often understood in the framework of continuum mechanics, i.e. considering the cemented sand to be a continuum, and performing simple elemental tests such as uniaxial compression or tension tests. A simple model to describe the behaviour of these cemented sands has been to use elements such as springs, sliders, and dashpots in parallel to describe the existence of a cohesive and frictional component of strength (Jiang et al. 2007). In the framework of continuum mechanics, the mechanical response of these cemented sands have been modeled using traditional plasticity theory, for example utilizing variants of the traditional Mohr Coulomb model, Modified cam clay (Sun and Matsuoka 1999) etc. The typical sand models are amended to include the effects of cohesion, as proposed by Lagoia and Nova (1995).The mechanical response of cemented sands is understood to be transitional to the behaviour of a typical particulate assemblage and a quasi-brittle material. However, the effect of length scale in these material systems has not been investigated. This paper presents a series of uniaxial compression experiments conducted on weakly cemented sand specimens of various sizes. The effect of length scale in these weakly cemented sands is investigated through unconfined compression tests."

Jan 1, 2015

"Since the publication of a literature review on this topic (McIvor and Finch. 1986), the author has been involved in more than thirty rod milling efficiency tests at seven different milling operations. There have also been several interesting additional articles on rod milling found in the literature. This update describes the most important findings from investigations carried out since the previous review, as well as developments in the techniques used to evaluate the grinding efficiency of rod mills.Rod Mill Power DrawA series of power draw versus charge level measurements were taken on the rod mills at Leg Mines Selbaie and Kidd Creek Mines Limited (McIvor. 1989). These revealed a significant (about 10070) loss in power draw with the installation of new liners, a common observation among rod mill operators. This loss in power draw depends on the liner design, particularly the average thickness and lifting face profile of the new liner. The theoretical power usage of both mills was approximately 5070 lower than that given by the Allis-Chalmers power draw formula (Rowland. 1982).The shape of the power draw versus charge level curve suggested an opportunity for reducing media consumption. By increasing mill speed, power draw could be maintained with a smaller mill load, most significantly for mills operating with high charge levels. Analysis of data from 62 installations confirmed this (McIvor, 1989). On average, a 10 percentage point increase in critical speed results in a 15070 in specific media consumption (Le. kg of steel consumed per kWh of energy consumed). Table 1 lists a number of relatively high speed rod milling installations. No negative effects on the drive train, liner life, or the tendency for rods to tangle have been reported. A reduced lifting angle of the new liner leading face (e.g. 45 versus 90 degrees from the mill perimeter) is recommended for higher speed rod mills (Leclerq, 1990)."

Jan 1, 1995

Adsorption of guar gum was investigated on model minerals typically found in water-insoluble slimes of potash ores. The ionic strength of the adsorption mixtures was varied from distilled water to saturated NaC1 and KC1. The tests were also performed using a saturated brine. The adsorption density of guar gum on illite, dolomite, and kaolinite was found to be not only a function of total ionic strength but also of the type of electrolyte present in solution. Generally, the adsorption of the polymer on the clay minerals was higher in KC1 solutions than in NaCI, but the differences gradually disappeared when the electrolyte concentration approached saturation. The results were analyzed in terms of chaotropic (K ) or kosmotropic (Na') properties of background counter-ions and their different behaviors at the hydrated mineral-water interface. Guar gum adsorption on dolomite was related to the presence of calcium and magnesium sites on the mineral surface, and to the cis-configuration of OH groups of mannose units. Based on viscosity measurements, it was suggested that guar gum macromolecules form entanglements in NaCI solutions (and distilled water), but individual molecules exist in KC1 solutions thus creating different modes of adsorption of guar gum from concentrated NaCl and KCl solutions.

Jan 1, 2006

By W. G. Q. Johnston, J. S. Fox

"A semi-continuous belt of Aphebian magnesian extrusive and related sub volcanic rocks has been delineated by the authors over a distance of 150 kilometres in the central La Ronge lithostructural domain in Saskatchewan. They range in composition from peridotite to basalt and attain true thicknesses of up to 800 metres. For the most part, they are confined to a basal position in the La Ronge/Wasekwan metavolcanic sequence, and appear to have been extruded onto sialic crust. However, at Lynn Lake, Manitoba, in the eastern continuation of the domain, similar rocks seem to be found only at a relatively high stratigraphic level.Although metamorphism seems to have obliterated any quench textures that may have existed in these rocks, ultramafic pyroclastics, pillows and amygdules as well as intercalated, fine-grained clastic sediment attest to the volcanic nature of these rocks. Bulk chemical analyses indicate that these rocks are heterogeneous, with komatiites, tholeiitic picrites and even rocks with some of the characteristics of boninite all being present. The close temporal and spatial relationship of ""boninites"" and komatiites in the belt may suggest a high degree of Aphebian mantle inhomogeneity in the domain.Nickel-copper mineralization has been found in differentiated tholeiitic stocks within the La Range domain (e.g., at Lynn Lake), but none has yet been found in the ultramafic volcanics. However, the tholeiitic picrites do host one economic gold deposit - the Agassiz Mine at Lynn Lake .Because of the depleted nature of their mantle source, the ""boninites"" of the domain are not considered to be likely hosts for mantle-derived nickel sulphide mineralization. However, the opposite is true of the komatiites, and this points to the need for whole-rock geochemistry in setting priorities for nickel exploration in the belt. In contrast, all of the magnesian rocks of the domain probably have the same gold potential. Inasmuch as the geophysical response to gold mineralization in the domain is likely to be unpredictable, lake-sediment geochemistry, followed by base-of-till and/or alteration oriented lithogeochemistry, should be emphasized during gold exploration."

Jan 1, 1981

By Xueyi Guo

The total production of nonferrous metals in China is 31.35 million tons in 2010, which ranks first in the world for the past ten years. Large quantities of solid wastes including slag, residue, ash, and dust are produced every year in the production of nonferrous metals. These secondary resources are renewable, not only contain abundant valuable metals which can be recycled by metallurgical processing, but also could eliminate the pollution to the environment. In this paper, the status quo of the main nonferrous metals recycling in metallurgical industries in China was summarized, and the future development of waste processing and resource recycling was prospected.

Jan 1, 2011

By G. Vibert Douglas

It has long been recognized that there are considerable difficulties connected with the physical and chemical explanation of shoots of ore. The literature of Mining Geology has frequent references to this subject; but the reasons given for the actual location of primary ore-shoots have been very tentative. In practice, the geologist or miner depends to a great extent on sampling and assaying to determine the boundaries of profitable ore within the mineralized zone. An ore-shoot is a portion of a mineralized zone which has a sufficient concentration of valuable minerals to make it ore. Ore deposits have been classified according to many schemes. There are classifications based on morphology and on genesis. However, it is also possible to classify hypo-gene hydrothermal ore-bodies in terms of ore-shoots or vice versa. A.-Ore-deposits in which the whole of the mineralized zone is ore, which is the same as saying that the ore-shoot extends throughout the mineralized zone. An example of this type of deposit would be the pyritic mineralization of Rio Tinto. B.-The second type would comprise those deposits in which the ore-shoots bear some relation to the geological structure of the area: all deposits in which 'bonanzas' occur in folds in the zone, at the intersection of veins along certain fault (or joint) features and other such controls. There are a great many examples of this type: (a) The Rhodesian copper lodes, where Bancroft and Gray have pointed out the association of the shoots of ore with drag folds. (b) A similar association has been pointed out by Faribault for some of the gold veins of Nova Scotia, e.g., Mount Uniacke. (c) The great bonanza of the Comstock lode in Nevada. C.-Deposits in which the localization of ore has been effected by the chemical character of the wall-rock. Examples of this type are the Fahlbands of Scandinavia, and the 'indicator' deposits of Ballarat, Australia. D.-Lastly, deposits in which there is apparently no structural and no chemical control.

Jan 1, 1934

By A. J. A. Janse

The Argyle diamond mine is located in the Kimberley region in the northern part of Western Australia at approximately 16°43'S, 128°23'E in subtropical bush/shrub savannah. The 46 ha (113 acres) pipe is situated in the Halls Creek mobile zone, which was cratonized about 1800 Ma ago and adjoins the Kimberley Block, a plateau upheld by Early to Middle Proterozoic rocks underlain by presumably Archean basement. The crater and upper pipe zones of the diatreme have been preserved within tilted and faulted strata of Middle to Late Proterozoic age. Its age of intrusion is about 1178 Ma. The diatreme rock is not kimberlite but olivine lamproite which was not recognized as a primary diamond host rock until 1978. The Kalumburu Joint Venture, a syndicate of local and overseas mining companies, started in 1972 to prospect the Kimberley Block with a modest annual prospecting budget of $125,000. Annual expenditure doubled each year and reached $1 million in 1975 after which CRA took over management of the newly-formed Ashton Joint Venture with an annual expenditure of $2 million (all figures are in current US dollars). The Argyle diatreme was found near the end of the eighth field season and after a total expenditure of $12 million. Annual production is about 34 million carats which places Australia as the world's number one diamond-producing country, although, because of the low value per carat, it is number four in terms of total annual value. The global cost required for a major diamond mine is between $300 and $400 million, with a discovery every 10 to 12 years. Because Canada contains large areas of prospective platform rocks underlain by Archean basement it is likely that a major diamond deposit does exist there. It is probable that it will be found before the end of this decade and exploration expenditure may be in the order of $100 million.

Jan 1, 1992

By Dong-kil Lee

A ventilation improvement index (VII) was developed in an effort to propose an optimal ventilation improvement measure for a designated mine. The index was designed to incorporate both ventilation effectiveness and cost efficiency. To verify the effect of the index, a mine experiencing a ventilation problem was selected, the ventilation assessed and ventilation network analyses were conducted. In the selected mine, four areas with ventilation difficulties were detected and an improvement schemes for each area were proposed. These were simulated through network simulation and evaluated using the ventilation improvement index. As a result, an optimum ventilation improvement scheme was proposed. The improvement scheme with the highest VII value was incorporated in the design.

Aug 1, 2013

By C. A. Pickles

Microwaves are being applied industrially in a number of fields and possess a number of potential advantages that could be applied in extractive metallurgy. Microwave heating has a number of unique characteristics, which distinguish it from the conventional heating methods. In microwave heating, the absorption of energy occurs at the molecular or atomic level and therefore the heat is generated within the material itself. In conventional heating, the exterior temperature is usually higher than the interior, but with microwave heating, the temperature gradient is inverted and thus the interior temperature can be much higher than that of the surface. The energy densities in microwave systems can be relatively high and this in combination with the low thermal conductivities of oxide ores, can lead to very high internal heating rates. In this research, the application of microwave energy for the reduction of a nickeliferous silicate laterite ore has been investigated. The real and imaginary permittivities are the two fundamental properties which determine the microwave absorption characteristics of a material and these were measured for the ore, both with and without carbon additions using the cavity perturbation technique. The microwave heating and reduction experiments were performed in a 0 to 2 kW CTC microwave system. The effects of power input, processing time, sample mass and charcoal additions on the microwave absorption were investigated. Then reduction experiments were performed and the nickel grade of the magnetic concentrate and the nickel recovery in the concentrate were determined.

Jan 1, 2005

By Thomas J. Ferree

"IntroductionThe objective of this chapter is to provide important criteria relating to the operation and maintenance of spiral concentrators and Reichert cones, so that optimum performance can be achieved and maintained in a mill circuit. Little space will be devoted to the theory of the forces of separation or the technical details of the devices, as that in(ormation is available from other sources. The discussions in this chapter will be presented in a manner that should be interesting and useful to anyone who may be exposed to these methods of gravity separation. It is most important for the reader to understand that each application of the equipment should be considered on its own merit, as no two suites of minerals respond exactly alike in mineral processing circuits. The information herein should be useful as a guide toward increasing the metallurgical efficiency of the processes, irrespective of the type of minerals being concentrated.It is highly recommended that certain tools be made available to the personnel responsible for metallurgical performance of the spirals and cones; these include a set of standard type laboratory screens, a medium quality petrographic microscope, and a vanning plaque or gold pan. These devices, when used in conjunction with a basic understanding of the mill circuit, provide immediate knowledge on the characteristics of any sample in the mill circuit. The information gained by visual evaluation of sized fractions of products can solve many of the basic problems of mineral beneficiation, and at the same time increase the interest of the non-technical personnel associated with the plant operation. There are many problems associated with obtaining satisfactory performance from a mill operation. Many of the problems have to be solved using a technical approach. Some of the problems, however, can be solved with practical knowledge and common sense. It is hoped this chapter can provide a guide to both techniques."

Jan 1, 1989

By Rory S. McLaren

"Cost per ton as it relates to mobile underground vehicle maintenance and repair is invariably astronomical. The reason: the mechanic replaced a ten micron filter with a 250 micron filter because it appeared to be physically identical; the purchasing agent bought ""Brand X's"" filter because it was 25% cheaper- the result, a filter life 30% less; the hydraulic pump was sent for repair at a cost of $2000.00 and failed shortly after installation; pumps are failing because there is no testing equipment available in the workshop.This paper will identify a number of the common pitfalls resulting in high costs, and offer solutions to bring these costs to within acceptable limits.The following key issues will be discussed:A. Genuine parts replacement versus ""Brand X""B. Evaluating outside repair shopsC. The influence of environmental conditions on optional equipmentD. Personnel trainingE. Servicing and testing equipmentIntroductionThere is an old English aphorism, ""Look after the pennies, the pounds will look after themselves"".This paper is aimed at the pennies.Genuine Parts Replacement versus ""Brand X""High interest rates and floundering mineral prices are forcing many equipment owners and operators to purchase ""Brand X"" or ""Pirate"" parts in order to survive. ""Pirate"" part manufacturers invariably offer lucrative discounts which are most appealing to the average purchasing agent whose responsibility it is to save money. Unfortunately, in most cases, there is a tradeoff involved; when the price drops so does the quality, which can adversely affect owning and operating costs. Unless totally obvious, ""original"" or ""pirate"" part metamorphosis requires scrupulous analysis by competent individuals.Hydrostatic transmissions have been used successfully for many year s in various industries. They found their way into mobile underground loaders in the early seventies and have since become very unpopular in mining circles due to their apparent ""inherent unreliability"" and near ""prohibitive"" repair costs. The big benefit of the hydrostatic transmission is compact component packaging which enables the machine to work in very small headings."

Jan 1, 1983

By Ian Sherrell, Michael Pirie, Riddhika Jain

"Conventionally, direct reduced iron (DRI) mixed with char from rotary furnace discharge is cooled down before separating the char via magnetic separator. The DRI must then be reheated for further downstream processing (i.e. smelting). Outotec has developed the PyroMagTM specifically to conserve thermal energy during this separation process. DRI is cleaned by feeding hot particles directly after reduction onto a rotating surface just above a stationary magnet assembly. This allows hot feeding of the clean DRI to the downstream equipment which renders the entire process more energy efficient.Continuous testing of the PyroMagTM was performed to verify design and separation. Because the use of DRI/char was not practical for test use, the pilot plant testing was conducted using a magnetite/sand mixture heated up to and beyond the Curie temperature of magnetite. The separation was efficient for both cold and hot processing with recoveries varying from 98.6% to 99.74%. As expected, the recoveries decreased to as low as 83.99% as the temperature was increased beyond the curie temperature of magnetite (585°C). Long term separation was thermally efficient and sustainable.INTRODUCTIONMagnetic drum separators are widely used in the mineral processing industry. These low to high intensity magnetic separators are used to concentrate various magnetic material, such as iron ore, and less magnetic materials, such as ilmenite and garnet from silica and other contaminantsThere are multiple classes of magnetism based on the effect that each material displays while in and out of a magnetic field. In a simplified fashion, which is more appropriate for mineral processing, the magnetic susceptibility of a mineral is used to define it as being ferromagnetic, paramagnetic or nonmagnetic. In this respect, minerals with varying degree of magnetic susceptibility can be selectively separated with different types of magnetic separation equipment in different stages of a process flow sheet. The lower strength magnets are employed early in the process to remove highly magnetic fractions. Later stages utilize greater magnetic fields for capture of less magnetically susceptible material."

Jan 1, 2016

By Ludwig Strah

"The paper outlines the history of rubber mill liners beginning with the awarding of the first North American patent rights in 1923. It includes the early-day applications and installations of the 1920s and 1930s through to their present-day application. Over-all rubber mill liner design is discussed in relation to present day installations employing rubber mill liners in various grinding applications. A brief overview is given of a manufacturer's procedures to produce rubber mill liners.IntroductionIn early 1910, it became apparent that the productive life of the silver mines in Cobalt, Ontario, could be substantially prolonged by mining lower-grade ores.The mining of lower-grade ores required the establishment of treatment facilities which initially employed jaw crushers, stamp mills, jigs and concentrating tables. The recovery of silver in other than a coarse metallic form required the additional step which was introduced at that time: the cyanidation of silver concentrates from the tables. The resultant requirement for increased and improved recovery led to the introduction of grinding mills, i.e. tube mills.Shortly after the introduction of the tube mills to the Nipissing Mine's mill in 1914, the idea of using rubber as a mill liner material was spawned. However, no action was taken due to the advent of World War I. It is interesting to note that the idea of rubber mill liners emanated from having used pieces of solid rubber tire as an emergency repair on a set of liners when a liner had broken and no immediately available spare was on site.The credit for this ingenuity and subsequent application for a patent on rubber mill liners falls to James John Denny, the mill superintendent, and Rolla Barnum Watson, the general manager of Coniagas Mines Limited. The patent application was formally made on December 3, 1921. The U.S. Patent Office issued Patent No. 1,470,597 on October 16, 1923."

Jan 1, 1982

By M. Roy, Y. Roux, R. McIvor

"La Compagnie Minière Québec Cartier at Port Cartier. Quebec, utilizes six parallel wet grinding open circuit ball mills to grind iron ore concentrate for the pellet making process. In an attempt to improve grinding efficiency, a mix of 605 25 mm (1 in.) and 405 38 mm (1 ½ in.) grinding balls was charged to mill D while the opposite mix was used in all the other mills. Because media costs were higher for mill D, it was necessary to find out if the smaller balls provided a net economic benefit. Therefore, a test program was undertaken to determine the relative efficiency of the charge in mill D compared to the other grinding mills. Mill F was chosen for comparative efficiency measurements.The first problem was to choose a suitable definition of grinding efficiency. In management's view, previous attempts to quantify milling performance using population balance modelling were ineffective because they were unable to relate breakage rate functions to grinding circuit economical in any meaningful way. Work index analysis was also not favoured because of a number of complicating effects, including the fineness of the feed and product streams and open circuit grinding.It was decided to adopt a recent innovative approach which had been successfully applied to closed-circuit ball milling. For ""functional performance analysis of ball milling"" (McIvor et al., 1990; McIvor et al., 1991), a size criterion is first chosen to define ""fines·' and ~'coarse particles"". In this case 45 microns (325 mesh) was chosen to distinguish fines from coarse material. The relative grinding efficiency of each ball mill environment would then be ~e cumulative grinding (disappearance) rate of coarse (plus 45 nucrons) material per kilowatt hour of energy applied to the coarse material in the mill, divided by the grindability (measured in the laboratory) of the coarse material in the mill feed.Because the disappearance rate of coarse particles is also the production rate of fines, the following equation can be written for the efficiency of the mill environment, Effe, as defined above:"

Jan 1, 1994

By G. Agar

"E. Joe We have been working for some years on the elimination ofCANMET thiosulphates from solutions as a means of protecting ground water and fish streams. Alot of these thiosulphates result from the mill use of S02. Is there any possibility that thiosulphates will be generated by this process?Reply I don't know the answer to your question. I can speculate that there is no question that when one adds SO2 and air to a sulphide pulp one generates a host of unsaturated poly¬cyanates. I unfortunately cannot remember the reactions which arise in the cyanide destruction scheme. I think that if thiosulphate was generated the process wouldn't work. It is not oxidizing enough to destroy the cyanide.R. Wagner Would you comment on the actual cost/ton at the Scottie goldPamour mine concentrator? Would $2.50/ton be typical or is thisPorcupine just a ball park figure.Reply We wouldn't disclose Scottie's costs even if we knew them. I depends on the amount of cyanide and thiocyanate in the effluent. I have given you a hypothetical case. With the conditions stipulated the figures I Have given are realistic.K. Konigsmann I do not believe the use of S02 adds significantly to the thio Noranda Mines salt problem Certainly we use SO2 in many of our mills and thio-salt levels in a number of them haven't changed whether we use or don't use S02. For instance, Orchan and Matagami mills treating the same ore side by side in which one was using S02 and the other wasn't, the thio-salt level was virtually the same in both pulp streams. In the case of Brunswick where alot of SO2 is used (20 tons/day) there is a thio-salt problem but it is relatively insignificant.Reply Thank you for that comment. It serves to reassure us that we are not polluting the environment."

Jan 1, 1984

By K. K. Mamyrbayeva, G. Z. Guseinova, V. A. Luganov

This paper presents the results of theoretical and technological studies on leaching ore of Bozshakol deposit. The major oxidized copper mineral of Bozshakol ore is atacamite, and the content of malachite, azurite, cuprite, copper sulfide minerals is insignificant. The chemical composition of the test sample, with mass, %: 0.9 Cu; 55.00 SiO2; 22.40 AI2O3; 9.00 Fe2O3; 2.24 CaO; 3.56 MgO. Technological studies on leaching the ore were carried out with sulfuric acid 5, 15, 25 and 50 g/dm3 in one and multistage (up to 4 stages) countercurrents. Size of ore 1 mm, the ratio S : L 1: 4. It was found that with increasing of acid concentration from 5 to 50 at one-mode leaching the copper extraction reaches 55 %. At multistage leaching the copper recovery reaches 73 %, a productive solution containing g/dm3: Cu - 4; Fe – 1.2; Al – 1.1 at a pH of 2.8. When leaching the cinders obtained by ore roasting in an atmosphere with limited value of oxygen, with increasing of the roasting temperature from 650 up to 750 °? the extraction of copper from the Bozshakol ores reduced from 31.5% down to 24.4%. The rate of filtration is markedly increased compared with the crude ore. When leaching the cinders obtained by roasting in a nitrogen atmosphere a marked increase up to 87.2-88 % of copper extraction from Bozshakol ores is observed. As an extractant for the extraction of copper from the pregnant solution used LIX984N extractant in kerosene. On the basis of the calculations it was built extraction isotherm and modeled processes of copper recovery with the circuit 2E + 1P, allowing up to 93 and 96% extraction. The through extraction of copper from the ore in the electrolyte was 67 %. Technology is recommended for semi test.

Jan 1, 2016

By L. B. Smith

Introduction The property of Wright-Hargreaves Mines, Limited is situated at the northeastern end of Kirkland lake, Ontario. It lies immediately east of the Lake Shore mine, and about in the centre of the Kirkland Lake camp. The district is served by the Swastika-Rouyn branch of the Temiskaming & Northern Ontario railway, the town of Kirkland Lake being four miles northeast of Swastika, which is 63 miles north of Cobalt and 392 miles north of Toronto. The mine is developed by three shafts and has more than twenty-eight miles of underground workings. There are two main producing veins, which lie in a porphyry boss 1,350 feet wide. Milling started in May, 1921, and to-date (January, 1933) the mine has produced gold to the value of $21,770,183. Present production is at the rate of 800 tons of ore per day. For a description of the geology, reference may be made to the report by E. W. Todd on the Kirkland Lake Gold Area (1). Structure and Ore Deposits The two major productive veins on the Wright-Hargreaves property are known as the North vein and the South vein. In the vicinity of the shafts these veins are approximately 550 feet apart. A third vein, known as the Inclined vein, has been worked from its junction with the South vein at the 2,000-foot level to its intersection with the North vein at the 800-foot level. This inclined vein has to date been found only in the area east of the eastern major north-south fault. The main North vein and South vein have a general vertical dip, although local rolls and post-ore faults throw certain sections out of line. From the surface to 3,900 feet, the North vein is displaced south less than 200 feet, and the South vein at the same horizon is almost perpendicularly beneath the surface showings.

Jan 1, 1934

By W. A. Jones

"The Hollinger gold mine is in the Porcupine area, Cochrane district, in northern Ontario. The adjacent town of Timmins, which has a population of 25,856, lies 490 miles by railway north of the city of Toronto.The property was staked in the late summer of 1909. The original four claims were staked by Benjamin Hollinger, after whom the mine was named. Development commenced in the autumn of 1909 and mill construction in the winter of 1910. A catastrophic forest fire swept the district in 19ll, destroying the surface plant, so that the original 300-ton mill was not completed until June 1912. From that date up to December 31, 1945, 44,280,750 tons of ore have yielded 13,636,796 oz. gold and 2,613,354 oz. silver. The total value of the production has been $363,638,315.0REBODIES AND MINERALIZATIONThe orebodies consist of quartz veins, stringer lodes, and scockworks, with associated pyritized wall-rocks. The width of individual orebodies, as exposed by mining operations, ranges from 1 to 70 feet and averaged approximately 10 feet in 1942. The vein system forms a complex pattern in a favourable zone, which in places is 6,000 feet long and 4,000 feet wide. The deepest level being developed is 5,150 feet below the surface. Few veins are persistent for long distances, although some are continuous for more than 1,000 feet longitudinally and vertically.The host rocks consist of a closely folded series of Precambrian lava flows of Keewatin age. These are schistose and hydrothermally altered, but are believed to have originally been of intermediate to basic composition. The schistosity strikes N.70°E. and dips to the south at 75° to 80°. The structure of the formations is complex, but in the principal ore-bearing zone the general dip of the flows is to the south at a slightly flatter attitude than the schistosity."

Jan 1, 1954