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By James F. A. Hatley

Northern Saskatchewan comprises an area of about 350 000 km2 and boasts some of the world's largest known high-grade uranium deposits. The successful mining of these deposits, however, can not be accomplished without overcoming technical challenges. Not only is the high grade of the uranium ore challenging, but mining methods need to be developed to deal with groundwater at very high pressures, and ground conditions that vary substantially from excellent to wholly unconsolidated clays and sand. This paper discusses the various mechanisms of inflows at the three of the Cameco Corporation Mines in Northern Saskatchewan. The risk of inflows is quantified in terms of unique challenges from hydrogeologic conditions, rockmass integrity, and uncertainty in geologic conditions. Mitigation strategies in case of an inflow are also briefly described. The paper concludes by addressing the impact of institutional rules (corporate standards) on minimizing, monitoring, and controlling the probability and/or impact of unexpected inflows.

Aug 1, 2010

By A. F. Abd El-Rehim

The behavior of both transient and steady state creep of Al-4.5 wt% Cu and Al-4.5 wt% Cu-0.1 wt% In alloys was investigated using a constant stress where torsional oscillations of different frequencies and shear strain amplitudes were operated at various working temperatures. An augment in both transient and steady state creep rates was observed by increasing both frequency and shear strain amplitude of the applied oscillations. The mean values of the activation energy of both alloys for both creep stages were found to be equal to that quoted for dislocation intersection mechanism.

Jan 1, 2006

By A. Pawelczyk, D. Hreniak, E. Zych, W. Strek, K. Grabas, A. Ostrowski

"Mongolia is a country rich in rare earth resources. In the experiments carried out samples of the Mongolian ore were processed to obtain a lanthanide concentrate. The concentrate underwent hydrometallurgical processing. At the same time radioactive elements were removed. The lanthanides were obtained in the form of hydroxides that were further processed in order to get concentrated products. As a result cerium, lanthanum and neodymium-praseodymium concentrates were separated in the process. Also a concentrate of remaining elements: samarium, europium, gadolinium, terbium, dysprosium, holmium and erbium was obtained. INTRODUCTIONAccording to the assessment of the US Geological Survey, Mongolia possesses 16.77 % of the world total rare earth (RE) resources (Daly, 2011); that is the second largest RE reserve in the world, estimated to be 31 million tonnes in terms of rare earth oxides (REO). In the face of Chinese RE trade restrictions, Mongolia emerged as an important economical partner in mining and exploitation of these critical raw materials. This country is recognized as having more than five large RE deposits in the southern provinces. The most important are Lugeengol (Lugin Gol), Mushgia Khudag, Khotgor and Khalzan Buregtei (Eurasia, 2010). The Mongolian government has taken measures to facilitate western investments, nevertheless the country is economically dependent on China, the basic trading partner, particularly where exports are concerned.In the late 1970s Hydromet Kowary Enterprise, Poland, took up, along with the Wroclaw University of Technology, research and technological study on lanthanides recovery from raw materials of Vietnamese and Mongolian origin and from Indian monazite sands (Kowalczyk, 1990; Kowalczyk & Mazanek, 1987; 1990). For that purpose, a rare earth processing plant had been constructed with the capacity of about 10 t/year of hydroxide concentrate. The facility included several technological operations like grinding, flotation, hydroxide concentrate production in the hydrometallurgical processing, particular RE separation, and production of commercial products in form of CeO2, La2O3, Nd2O3, Pr2O3, and misch metal. Additionally work had been carried out in Kowary on the manufacture of yttrium and europium oxides of high purity, which were intended for color TV picture tube production in Poland."

Jan 1, 2012

By R. Laucournet, E. Billy, M. Joulié

Today Li-ion batteries are considered as one of the most attractive available energy storage solutions. However in order to overcome environmental, legislative and economic issues, the management of used batteries has to be considered and mainly the recycling. The present paper deals with the recycling of used Li-ion batteries and the recovery of high value metals from active materials of electrodes. Lithium nickel cobalt aluminum oxide (NCA) and lithium manganese nickel magnesium oxide (LMNM) cathodes, which contain most of valuable elements typically used in the composition of active materials for Li-ion batteries are investigated by a hydrometallurgy route. The process is divided into two steps: the materials leaching and the elements recovery by a selective chemical separation. Some experimental parameters (acid nature, concentration, time and temperature) are studied to determine the most efficient conditions of dissolution. Then, the valuable elements such as nickel, cobalt and manganese are selectively recovered with a high efficiency close to 100 % and with a high purity by precipitation. By this study, an adaptable hydrometallurgy process can be defined according to the composition of Li-ion batteries to recycle.

Jan 1, 2014

By C. W. Eggert

This paper gives an account of the E.V.O.P. tests made at the Copper Rand mill of the Patino Mining Corporation in Chibougamau, Quebec. The runs were made to acquire an understanding of the method. A description of events, along with conclusions, observations and opinions, is out-lined.

Jan 1, 1967

By V. C. Wansbrough

First, permit me to say with what very great pleasure we find ourselves assembled in this most ancient and honourable city of Quebec. It is a city which we deeply respect as the fountain-head of the culture of French-speaking fellow citizens, and the seat of government of a province which is steadily recording new achievements in Canadian mining. Such achievements prove what can be accomplished when enlightened and progressive government policy, technical and scientific competence, and private initiative pull forward together in harmonious rhythm. It is a most happy and auspicious event also that the guiding control of the affairs of the Institute and of the fine traditions of provincial administration have for the past year been united in the person of our honoured President, Dr. A. O. Dufresne. This year, again, we have bad prepared for us by our hosts a splendid programme of technical sessions and social entertainment. Each year we pay an unspoken compliment to the Institute, expecting as a matter of course nothing but the highest standards in both these spheres. In its annua1, semi-annual, and regional meetings, in the constant activity of its local Branches, the Institute continues with marked success to perform its great dual service, social and educational. We are all its beneficiaries to the extent to which we choose to profit by participation in its affairs. But, one imagines, to none are the advantages of Institute membership more valuable than to young men embarking on their professional careers.

Jan 1, 1951

By Q. G. Reynolds, R. T. Jones

"Residence time distributions are well defined for batch and continuously stirred reactors. However, there exist many smelting furnaces that do not fit either of these classifications. A mathematical description is presented for the case of a furnace that is fed continuously, but where the liquid slag and metallic products are tapped intermittently. The effects of slag and metal tapping intervals and the fraction of liquid retained in the furnace, are shown to have a significant effect on the residence time distribution.INTRODUCTIONThere are some processes in which the time spent by the feed material in a smelting furnace plays a significant role in determining the extent of the reactions between metal and slag and influences the degree of settling of metal droplets from a molten slag.Residence time distributions are well defined for batch and continuously stirred reactors. However, there exist many smelting furnaces that do not fit either of these classifications. Here we will focus on producing a mathematical description for the case of a furnace that is fed continuously, but where the liquid slag and metallic products are tapped intermittently. The effects of slag and metal tapping intervals and the fraction of liquid retained in the furnace, are studied.Previous work (Jones, la Grange and Assis, 1997) has studied the effect of retention time on the degree of cobalt recovery achieved in a slag reduction process. In that work, a method was proposed for the calculation of the mean residence time of the contents of a pilot-plant furnace. That approach is to be extended here for the purposes of providing a more general design or scale-up tool.The maximum allowable depth of molten material in a given furnace is affected by such factors as the desired batch size (or the size of the ladles) and the pressure of the molten material that the tap-hole and surrounding system can withstand. Within these constraints, there might be some scope for choice of taphole position. The effects of this choice need to be understood mathematically."

Jan 1, 2012

By The Staff

"SummaryDetermination of uranium in ores may be made by measurement of their radioactivity.The method most commonly used is that of measuring beta radiations with the tray-type beta counter. This will give an accurate uranium determination provided the sample does not contain thurium and that the beta and gamma activities have not been disturbed from their natural state of equilibrium. Such a loss of equilibrium will result from mining the ore or from the weathering of the surface rock.A method is described which eliminates errors due to loss of equilibrium. The beta and gamma activities are measured simultaneously in one simple operation. This method is recommended for assaying samples of products from milling operations and its use is limited only by the presence of thorium in a ratio greater than one part of thorium to twenty parts of uranium.A third method is recommended for the analysis of exploration samples and diamond drill cores, which may be assayed in the uncrushed state.Rapid determinations of a large number of samples is possible with the pressure ionization chamber, which is sensitive to gamma activity."

Jan 1, 1950

By M. Maeda, H. Sasaki

"Precious metals such as platinum, palladium, and rhodium are commonly used for automotive catalyst converters. To recover these metals with lower environmental load and cost, exposure to zinc vapor was proposed as a pretreatment prior to acid leaching. Alloying with zinc is expected to enhance dissolution of precious metals and reduce the amount of toxic chemical agents required for the leaching. This study examined the effect of the alloying pretreatment using model catalysts. Catalysts were immersed in leaching solutions for a predefined time with or without zinc vapor pretreatment, and the amount of dissolved metals was then evaluated. INTRODUCTIONAutomotive catalysts account for large portions of the world consumption of platinum (Pt), palladium (Pd), and rhodium (Rh) (Butler, J., 2011). Because of their limited production and high prices, recovery of the precious metals from spent catalysts is obviously an important issue. A representative structure of an automotive catalytic converter can be described as a monolithic ceramic support with a honeycomb structure which is coated with porous oxides such as those of aluminum (Al2O3), cerium (CeO2), and zirconium (ZrO2) (Heck & Farrauto, 2001; Matsumoto & Shinjoh, 2007). Small particles of precious metals, originally several nanometers or less, adhere to the oxides.Typical recovery processes for precious metals from scraps are shown in Figure 1 (a)-(c) (Habashi, 1997; Yoo, 1998; Shibata & Okuda, 2002). Precious metals can be recovered by using non-ferrous smelters as Figure 1 (a) shows. Smelted along with metals such as copper (Cu), precious metals and ceramics in the scraps are absorbed into the metal phase and slag phase, respectively. Because the metals separate from Cu as anode slime in successive electro-refining, such a process can recover precious metals as co-products of Cu. However, in order to prevent negative effects on the principal process, a specialized procedure is needed. Figure 1 (b) is a specialized approach in which scraps are melted at elevated temperature with Cu called the Rose process. Here Cu works as a collector metal, and the absorbed precious metals are then concentrated by oxidation of Cu. Such a process can separate precious metals from ceramics constituting catalyst converters. Smelting at high temperature, however, requires extensive equipment as well as a high amount of energy. On the other hand, hydrometallurgical processes may be effective to separate precious metals from ceramics into an aqueous solution, because chemical agents can act on these metals selectively (Figure 1 (c)). Leaching of precious metals, however, requires strong acid and oxidizing agents because of their corrosion resistance; this makes liquid waste disposal and the large environmental load controversial. To solve the problem, a metal vapor pretreatment was proposed (Okabe, Yamamoto, Kayanuma, & Maeda, 2003a, 2003b; Kayanuma, Okabe, Mitsuda, & Maeda., 2004; Kayanuma, Okabe, & Maeda, 2004; Itoh, Kayanuma, Miyake, Kimura, & Maeda, 2005). A flow diagram of the process using zinc (Zn) vapor is shown by Fig. 1(d). Exposed to Zn vapor, precious metals are expected to form alloys which dissolve more easily than pure precious metals, thus reducing the use of chemical agents and environmental load involved with successive leaching. In previous research (Sasaki, Miyaki, & Maeda, 2010; Sasaki & Maeda, 2010, 2012), dissolutions of precious metal alloys formed by Zn vapor treatment were examined in hydrochloric acid (HCl) using an electrochemical method, and it was confirmed that these metals dissolved at higher rates from the alloys than from their pure states (Figure 2). In this study, the availability of the alloying pretreatment was demonstrated by testing on model catalysts in leaching solutions."

Jan 1, 2012

By J. R. Parga T., P. J. Valenzuela G., L. S. Quiroz C., P. Guerrero G., J. L. Valenzuela G.

"Pressure oxidation is a hydrometallurgical pre-treatment process used for refractory auriferous sulfide ores. In the work reported here, a refractory pyrite ore sample from the Mulatos mine, Sonora, Mexico was characterized. The sample was subjected to atmospheric cyanidation, and simultaneous pressure oxidation and cyanidation under different conditions of temperature and pressure. The mineral characterization showed that pyrite is the main species in the sample. The gold extraction tests were carried out in an 1-litre Parr pressure vessel at a temperature range of 90-120 °C, a pressure range of 87.5–150 psi (using oxygen), and a cyanide concentration range of 0.8-1.2 %wt. The gold extraction results showed that diffusion through product layer is the rate controlling step.INTRODUCTIONIn hydrometallurgical precious metals extraction, the most commonly used process is sodium cyanide leaching, as shown in the next reactions involving gold and silver (Fleming, 1992; Marsden & House, 2006):4Au(s) + 8CN- (aq) + 2H2O(l) + O2(g) ? 4Au(CN)2 -(aq) + 4OH- (aq) (1)4Ag(s) + 8CN- (aq) + 2H2O(l) + O2(g) ? 4Ag(CN)2 (aq) + 4OH- (aq) (2)“Refractory” auriferous ores or concentrates often contain gold and silver locked inside other sulfide phases such as pyrite. This renders cyanidation ineffective, so oxidative pre-treatments are used in order to convert the sulfide minerals into porous oxide minerals and make cyanidation work more efficiently. Reaction (3) shows the oxidation of pyrite in alkaline media and the formation of the porous oxide (Koslides & Ciminelli, 1992):2FeS2 (s) + 15/2 O2 (g) + 8H- (aq) ? Fe2O3 (s) + 4SO4 2- (aq) +4H2O(l) (3)A process of simultaneous chemical pressure oxidation and cyanide leaching for a flotation concentrate produced in the mine of Bacís, Durango, in Mexico, gave good results (Parga et al., 2007). In the study presented here, it was intended to use this treatment for a refractory pyrite mineral from the Mulatos mine in Sahuaripa, Sonora, Mexico, modifying variables such as pressure, temperature and cyanide concentration to find the best conditions for oxidation and leaching."

Jan 1, 2012

By A. G. Roach

MILLING of Nova Scotia gold ores began in 1862 with arrastras and wooden stamps and has progressed to heavier stamps followed by tables and other forms of concentration for the recovery of the sulphides. Chlorination and cyanide leaching of the concentrates has been practised, and in 1936 the combined amalgamation and all-sliming cyanidation plant of Seal Harbor Gold Mines was built. It has since been found that it is not necessary to cyanide the full tonnage and now only the concentrates are being so treated. From his experience in this enterprise, the writer presents the accompanying flow-sheet and supporting comments. While he has not the temerity to maintain that it is the best or only flow-sheet for Nova Scotia gold ores, he feels that it warrants thoughtful consideration by members of the milling fraternity of the Province. THE ORE The gold in Nova Scotia ores occurs both free and intimately associated with sulphides-chiefly arsenopyrite. Much of the free gold is quite coarse and crystalline. This is especially true when it occurs in white quartz. When in the slate, it is usually in thin plates and scales, frequently coated and discoloured. The values bound up in the sulphides are not amenable to amalgamation and must be recovered by other methods-concentration, followed by smelting or fine grinding and cyanidation. While high freight and treatment charges apply on concentrates shipped from the Province to smelters, a considerable quantity has been so disposed of in the past. However, it is hoped that a customs cyanidation plant will soon become available to producers of these auriferous concentrates in Nova Scotia.

Jan 1, 1938

By C. Dupuis, P. Waite, S. Leboeuf

Casthouses throughout the aluminium industry commonly use chlorine gas as a fluxing agent for furnace and in-line treatment of molten aluminium alloys. Its positive impact on the removal of alkali elements and non-metallic inclusions is well known. Nevertheless, the elimination of its use is highly desirable due to the safety and environmental risks associated with storage and handling. Over the past decade, the aluminium industry has significantly reduced its chlorine consumption. This has been achieved by maximizing the chlorine gas utilization efficiency during metal processing, and by the chlorine replacement for crucible and furnace metal treatment applications. Operation of sealed in-line treatment units without chlorine is now possible, but with lower or inconsistent removal efficiency penalties for alkali and inclusions respectively. This paper reviews the progress made at Arvida Research and Development Centre regarding the reduction of chlorine and the development of an alternative to chlorine for in-line metal treatment without penalizing the metallurgical performance.

Jan 1, 2006

By D. R. Reeves

This paper describes the geophysical Jogging of coal that is being carried out for the National Coal Board in the United Kingdom. An explanation of the technique is given and results are presented showing estimations of seam thickness to within ±0.5 in. and ash content to accuracies of ±2.5 per cent. The basic method of interpretation is outlined and some conclusions regarding the conditions of roof and floor material which may be deduced from the logs are explained. An indication of future developments in this field, including the identification of coal type and moisture content, is given.

Jan 1, 1971

By H. A. Steane

THIS PAPER describes the re-grind section in the 2,000-ton per day, lead-zinc flotation mill of Canadian Exploration Limited, near Salmo, B.C. There are two re-grind units in operation, one in the lead circuit, and one in the zinc circuit. In each unit, the rougher concentrate is re-ground ahead of the cleaning circuit. These re-grind units have made valuable improvements in the concentrate grade, and with no loss in recovery by over-grinding galena or sphalerite. Since the initial installation of this re-grind section, the plant throughput has been increased gradually from 1,600 tons to 2,000 tons per twenty-four hours, and, as shown by the screen analysis of mill tailing, a coarser overall grind has been accepted, from 55 per cent to 45 per ?cent minus 200 mesh. There has been no increase in grinding capacity, but two mills have been taken from the main grinding circuit and used for the re-grind circuit.

Jan 1, 1958

By Perry L. Christian

"Mana/ta Coal Ltd. 's Costello Mine, located seven kilometres east of Estevan, Saskatchewan, has been operating at the present location since 1960. During the past twenty-three years, annual production of lignite coal has averaged 590 000 tonnes. The coal is used for power generation and for industrial and domestic use. Markets, operating techniques, facilities in place and reclamation are discussed. Introduction The Costello Mine (formerly the Klimax Mine) is located in southeastern Saskatchewan, approximately 210 kilometres southeast of Regina and 5 kilometres from the United States' border, as shown in Figure 1. The present mine, 7 kilometres east of the city of Estevan, has been operating since 1960. The history, however, of the Costello Mine and of the Estevan Coalfield dates from the initial settling of the area in the 1880s. Mining initially consisted of burrowing into the side of the Souris River Valley or other hillsides where the coal seam was exposed. Eventually, slopes were put down in the Estevan/Bienfait area and vigorous underground activity existed. In the late 1920s surface strip mining was introduced. The use of strip mining progressed to the point where it completely replaced underground mining in the 1950s and it is the only method in use today. During the past 23 years annual production has averaged 590 000 tonnes. The lowest production was in 1960 when the mine was starting up at 167 000 tonnes and the highest production was in 1977 when production was 1 509 000 tonnes. The Costello Mine controls approximately 11 400 hectares of mineral rights in the Estevan area. This is made up of 4 900 hectares of freehold coal and 6 500 hectares of leases. The area forms a large, irregular shaped configuration. Most holdings are contiguous and adjacent. The mine also owns some 8 400 hectares of surface rights. The lands under the Costello Mine's surface rights control, which are not being used in the mining operation, are leased to local farmers and ranchers. This ensures that the land is used to its full potential, both before and after mining."

Jan 1, 1985

By E. B. Wright

THE property of Falconbridge Nickel Mines, Limited, is in Falconbridge township, approximately thirteen miles northwest of Sudbury. From the mine and townsite, a gravel road leads to Sudbury and a branch line of the Canadian National railway to Sudbury Junction. The ore deposit is at the southeastern corner of the 'Sudbury Basin'. It lies in a sheet-like mass, striking in an east-west direction, between norite to the north and greenstone to the south. The general attitude of the ore-body is vertical, but with frequent variations in dip of as much as 10° to the north or south. The deposit is very irregular in outline, the width varying from a few feet to eighty or ninety feet. In the vertical plane, rolls are frequently encountered. Gouge is very often present at one or both ore contacts. The norite to the north shows much jointing. The chief ore minerals are pyrrhotite, pentlandite, and chalcopyrite. The nickel content of the ore is usually double the copper content, average assays ranging from 2 to 2.5 per cent nickel and 1 to 1.5 per cent copper. The ore is of three main types: massive, breccia, and disseminated. The breccia ore, the most common type, consists of sulphides containing rounded inclusions of highly-altered country rock, which, in places, are so large that sorting is necessary in the stopes. In the norite, the breccia ore may grade into disseminated sulphides, which frequently are sufficiently concentrated for profitable mining. In places in the breccia ore, concentrations of solid sulphide masses occur. Until early in 1935, all mining at Falconbridge was by shrinkage scoping, but this method was found to be less and less suitable as greater depth was attained. When mining had reached the 1,000-foot level, ore drawn from large shrinkage stopes was found to have been diluted as much as fifty per cent by sloughing of the wall-rocks, caused by increased rock pressure and insufficient support. It was apparent that this method of stoping was un-suitable and -Mr. E. J. Martin was engaged to superintend the introduction of a method of cut-and-fill stoping.

Jan 1, 1938

By E. W. O'Brien

Personnel protection involves effective grounding and positive clearance of faults. This paper deals, in a general sense, with various schemes to achieve fool-proof personnel protection and reduce, to a minimum, the effects of human error and equipment failure.

Jan 1, 1966

By A. R. Byers

"Porcupine Peninsular mine is at the extreme southwestern end of North peninsula in Night Hawk lake, Cody township, Ontario. It lies approximately 10 miles east of the gold-producing district of Porcupine.Gold in quartz veins was discovered on Gold -island in Night Hawk lake in 1906, the same year in which the now famous Porcupine gold camp was discovered farther west. Most of the claims now held by the Company were originally staked by Charles Auer in 1907, In 1910 when the level of the lake was exceptionally low, gold was discovered in a quartz vein near the shore and close to the present mine shaft.The mine is developed by a main shaft to a depth of 440 feet, with levels at 80, 180, 300, and 425 feet vertical depth and by a winze to 630 feet. A mill of 200 tons daily capacity was built during 1923, Production commenced on January 4, 1924, and continued without interruption to May 31; 1926, when operations were suspended due to a decrease in the grade of ore. Total production amounted to 27,416 ounces of gold from 99,708 tons of ore, an average grade of 0.275 oz. a ton. Diamond drilling and re-sampling of the mine during 1934 and 1935 indicated approximately 1,480,000 tons, having an average grade of 0.123 oz. gold a ton, above the 625-foot level."

Jan 1, 1954

By Garnet T. Page

"ABSTRACTThese societies •can help in such aspects of a vocation.al guidance program in secondary schools as: provide vocational and technical information, and assist with vocational exploration, such .as plant visits, films, etc. -as .an aid to the student in making a self-inventory. Evidence is increasing that employers' needs for engineers and scientists will not be adequately met in the future. Therefore, the E.I.C., C.I.M., .and C.I:C. recently revised and extended the activities of the Canadian Counselling Committee on Engineering and Science. This co-•operative .action is to be integrated with the educational system, nationally .and locally, and with provincial associations •of engineers and other groups interested in the provision of vocational information in engineering and science.DEFINITIONTHE scientific treatment of any subject requires consistent use of carefully defined terms. The National Vocational Guidance Association of the United States studied the problem of the definition of vocational guidance for .about 16 years, and in 1937 finally arrived at:""Vocational guidance is the pro-cess of assisting the individual to choose an occupation, prepare for it, enter upon and progress in it. It is concerned primarily with helping individuals make decisions and choices involved in planning a future and building a career -decisions and choices necessary in effecting satisfactory vocational adjustment.""RANGEAs a school enterprise, vocational guidance begins at that point in the individual's life when he finds it necessary to make choices that have a definite significance for his future vocation."

Jan 1, 1952

By R. L. Brown

The burning sulphide ores in the Sullivan Mine generate sulphur dioxide gases. This paper is concerned with a description of the environmental impact of SO2 on air quality in the mine and on the surface in the vicinity of the mine. As a result of surface caving, SO2 has been able to vent to surface and winds have carried it into one residential area of Kimberley. Monitoring practices are described.

Jan 1, 1977