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By E. W. Garner

THE greatest problem confronting business managers in modern times is not production, but distribution. It would be easy to expand productive output, but it is difficult to find markets for all that is produced. The fundamental source of economic difficulty, in the view of many observers, is a lack of proper balance between productive capacity and consumptive capacity. The studies conducted by the Brookings Institution on the general subject, "The Distribution of Income in Relation to Economic Progress", have led to conclusions of far-reaching significance, which may be summarized briefly as follows: (1) We have never lived beyond our economic means or capacity. (2) We have not been suffering from general overproduction. (3) We have not as yet reached a stage in our economic evolution at which it is possible to provide adequate standards of living for everybody. ( 4) It would require less than a 25 per cent increase in purchasing power among the masses to absorb our full productive capacity. (5) Production-curtailment programmes-whatever their merit in meeting temporary maladjustments-can only lead to national impoverishment. (6) Our inability to find markets adequate to absorb the full output of our productive establishments is explained by the unequal distribution of income and consequent restricted flow of purchasing power through consumptive channels. (7) The slow rate of growth of consumptive demand serves to retard the rate at which new capital is constructed, and hence the rate of economic progress. (8) The primary need is a rapidly expanding total income, and a broader diffusion of this income among the masses of the people. (9) A broader diffusion of income can best be achieved by reducing prices as rapidly as increasing efficiency will permit. That is to say, the benefits of technological progress, mass production, and low unit costs must be promptly passed on to consumers in the form of lower prices-giving more for each dollar.

Jan 1, 1936

By Jacob Wood, Robert Matusewicz

The Outotec® Ausmelt Top Submerged Lance (TSL) Process has achieved widespread commercial acceptance over the past 40 years for the processing of copper, nickel, lead, tin, zinc and precious metal bearing feeds. The continual development and evolution of the technology during this period has resulted in a more intense smelting process with higher energy efficiency and lower operating costs. More recently, considerable effort has been invested in decarbonisation of the process through improvements such as increased oxygen usage, dry feed injection and use of low-carbon auxiliary fuels. This paper covers a number of these recent advances to the technology and highlights their impacts on energy consumption and CO2 emissions for a range of non-ferrous smelting flowsheets. INTRODUCTION Relentless demand for improvements in the profitability, safety and environmental performance of non-ferrous smelting plants has driven a continual development and evolution of processing technologies in this industry. In response to these demands, a number of bath smelting technologies have emerged over the past 40 years and have been implemented for the treatment of primary and secondary feedstocks. The Outotec® Ausmelt Top Submerged Lance (TSL) Process (henceforth Ausmelt Process) is one such bath smelting technology which has been widely adopted in a large number of greenfield projects and brownfield plant modernisations (Wood & Hughes, 2016). Emissions of CO2 from the production of non-ferrous metals represent approximately 4-5% of global totals (Rankin, 2012), comparable with the almost 10% attributed to iron and steel production. Despite significant improvements in the energy efficiency of modern smelting technologies (Coursol, Mackey, Kapusta & Cardona Valencia, 2015), primarily through increased oxygen usage, further innovation and development is still required to address their reliance on fossil fuels such as coke and coal. Transformation of the non-ferrous smelting industry to low-carbon technologies and process flowsheets is lagging behind the power and transport sectors, which have witnessed significant growth in clean/renewable energy technologies to reduce CO2 emissions and address the issue of anthropogenic climate change (IRENA, 2018; IEA, 2018).

Jan 1, 2019

By Caitlyn J. McKinley, Ahmad Ghahreman, Maziar E. Sauber

Ever-growing global needs for critical metals, such as rare earth elements (REE), and base metals, such as copper and nickel, and their environmental impact motivated the pursuit of sustainable flowsheets. Regeneration of hydrochloric acid in chloride-based leaching flowsheets as a critical unit operation is an example of such endeavours. This study presents an efficient hydrochloric acid regeneration method from a magnesium chloride solution waste by addition of sulphuric acid in relatively low temperatures. The effects of temperature, initial magnesium chloride concentration, amount of sulphuric acid added, and airflow are examined. Experimental work are focused on temperatures between 80 and 95°C, initial magnesium chloride concentrations of 2 to 5 M, 100% to 220% stoichiometric sulphuric acid additions, and airflows between 200 to 600 mL/min. The developed technology recovers hydrochloric acid at a high rate and concentration with a relatively small environmental footprint. INTRODUCTION The mining and metals sector is under increasing environmental and economic pressure to make more efficient use of chemical reagents and energy in the production process. This affects several areas of process and flowsheet design such as the selection of the most appropriate unit operations and the overall design of the process steps paying heed to sustainability. Owing to the inherent inefficiencies of physical and chemical processes used in the mining industry, there will be materials and streams generated that are not a desirable part of the envisaged flowsheet. Thus, methods are implemented to recover and recycle such effluents. If the recovery and recycling are not economical, then procedures are followed to neutralize and dispose these effluents. In many cases, these effluents are potential contaminants or hazardous materials, and their disposal should be looked at in a much larger context. The safe management of the contaminants and hazardous effluents is a daunting problem in the mining industry and is fraught with technological, economic, social, and political challenges.

Jan 1, 2019

By C. R. Edwards

"Equity Silver Mines Limited owns and operates an open pit silver 111ine located 37 km southeast of Houston in the Omineca District of British Columbia, Figure l. The property is approximately 300 km west of Prince George and 417 km east of the Port of Prince Rupert. This is the largest silver mine in Canada and one of the five largest silver producers in the western world.Equity is a subsidiary of Placer Development Limited, which owns 70 of Equity's issued and outstanding Common Shares. Placer was responsible for financing and developing the property and now provides general management and administrative services.All mining is by open pit methods, using three 1600 P&H shovels, eight 45 t Caterpillar and three 77 t Wabco trucks, three 40R rotary drills, two D8 dozers, one 9888 front-end loader and other ancillary equipment. We are now phasing out the Southern Tail pit and bringing the Main Zone into productior; see Figure 2.Comminution in the concentrator is achieved via primary, secondary, and tertiary crushers followed by rod and ball mills. The flotation circuit, with Southern Tail ore, produces a concentrate containing between 3 000 and 6 500 g of silver per tonne, 201 copper, and 25 g of gold per tonne. Running Main Zone ore, concentrate grades will decrease to about 4 500 g of silver per tonne, 17% copper, and 18 g of gold per tonne. Design milling capacity is 5 300 tpd, which is often exceeded.The flotation concentrate contains relatively high levels of antimony and arsenic. Because of these impurities Equity was unable to obtain economically acceptable long-term sales contracts for flotation concentrate. For this reason a leaching process was developed, and a plant designed, built and placed in operation. Due to the complex and unique technology involved, many start-up problems were encountered but by March 1983 the leach plant was considered ""fully operational."" This paper describes some of our progress and problems in bringing the Equity leach plant through its start-up period."

Jan 1, 1984

By Brian Lewis

"IntroductionThe leaching method selected for the dissolving of uranium ore depends in part on the physical characteristics of the ore, as well as the uranium mineralization. There are two main methods of leaching employed in uranium circuits in North America, one being ""Carbonate Leaching"" on sandstone, and pitchblende type ores, and ""Acid Leaching"", on quartzite type ores.This chapter deals with the acid leaching of uranium ores found in the Elliot Lake area, using sulphuric acid, for the leaching reagent, and ""pachucas"" for the contactor.The basic leaching process involves the pregrinding of the ore and the mixing of the ore with hot sulphuric acid, the mixture is air agitated, the air being used to oxidize the ore as well as produce agitation. The chemical reaction is as follows: Uranium Oxide + Sulphuric Acid = Uranyl SulphateThe following equations show the actual dissolution of the ""Hexavalent Uranium"", which show that more than one ion or complex of uranium is possible.Uranium DissolutionsI. UO, + 2H-(UOJ + H,O2. (UO,) + SO, = -UO, SO, (Uranyl Sulphate)3. UO, + SO,-(UO, (SO,),)=4. (UO, (SO,),)= -(UO, (SO,),)'Test work carried out in the laboratory as well as in the plants has determined the most efficient methods of extracting the uranium from the ore. A very important factor that has to be determined during initial test work are the amount of acid consumers in the ore. Within the Elliot Lake orebodies are minerals such as chlorite schist, diabase, and basement material, all very high acid consumers, consuming up to 1000 lbs. of acid per ton of ore. Of these minerals only the chlorite schist contain any uranium at all.The following are the operating parameters that have to be controlled in the plant:I. sulphuric acid concentration2. temperature of the mixture3. agitation4. liquid solid ratio5. reaction time6. particle size of ground ore"

Jan 1, 1989

By D. Houghton

"I am very pleased to be a member on this panel today because abrasion and wear in mills has been part of my life over the past forty years. My presentation will deal with the role Ni-Hard has played in combating wear and abrasion in mills. Before Ni-Hard came along, unalloyed white iron was used in ball mills and one of the most popular types of white iron liners was the pocket liner. Slide #1 illustrates the design of the pocket liner used at Dome Mines in their 5' x 22' pebble mills. The pockets filled with grinding media and it was on this that the wear took place. The concept is not unlike the new magnetic liner which utilizes the ore and grinding media to form the wearing surface. In this slide please note the tapered pocket type of cored bolt holes. If cored bolt holes do not fill with grinding media there could be excessive grooving or wear at that area. Therefore all cored bolt holes should be tapered in this manner.Before Inco patented the name Ni-Hard several foundries were making the alloy iron under different brand names such as Diamond Alloy, Mitchelloy and Specialloy. In the early days some of the Ni-Hard was poor quality. Inco provided foundries with technical knowhow and provided assistance to licence them to sell under the name of Ni-Hard. This was in the late 1920's. The material we are discussing is Ni-Hard type I with the following analysis (slide #2). It is a martensitic white iron with BHN of 550 - 600 and is a relatively easy alloy to produce in either a cupola or electric furnace. The annual average production of Ni-Hard in North America is 90,000 tons with mill liners making up 25,000 tons and grinding media 30,000 tons, according to Warren Spear of Inco."

Jan 1, 1983

By D. J. Bowen

"The Great Dyke extends for 540 krn along strike, to unknown depth, and varies in width from 3 krn to 11 krn. Rb/Sr age determinations gives values of approximately 2470 Ma.The Dyke consists of a variety of mafic and ultramafic rocks, many altered to serpentinites. These include, from top to bottom, successions of gabbro/norites, pyroxenites, peridotiteslharzburgites, and dunites/picrites.The layers dip in from either margin, steepest at the edges (20°) and flattening toward the central sections. Four separate complexes are present. From north to south these are termed the Musengezi, Hartley, Selukwe and Wedza Complexes, and are described as 'boat-shaped' or 'pig-troughed'. All are similar in petrological sequences. Physiographically, the Dyke is spectacular in the north, where as a mountain range it meets the Zarnbezi Escarpment. Elsewhere in the central parts there is little surface expression and wide open treeless vleis. Dyke rocks cut through granite and schist belt terrain and are termed 'The Magnesia Belt'.Vast mineral resources exist. These include chrome, asbestos, platinoids, nickel, copper and gold. Up to eleven chrome seams numbered from top to bottom occur in the lower ultramafic sequences. Seams are narrow at 10 em to 30 em but of high grade. The platinum occurs 10 m to 40 m below the norite/pyroxenite contact.A 60 m wide zone of Light Sulfide Mineralization (LSM) of less than 1070 sulfides extends from the lower gabbronorites into the upper pyroxenites. In this LSM, four enriched zones carry up to 5% sulfides. The most widespread is termed the Main Sulfide Zone (MSZ), below which is No. 2, No.3 and No.4 Sulfide Zones. The MSZ carrying pyrite, pyrrhotite and chalcopyrite is a multielement ore with components of PGE (pt-Pd-Rh), nickel, copper and cobalt, and gold and silver.Extensive exploration programs have been carried out and the intention is to begin large-scale mining. However, with Russia supplying metals to the western markets, South Africa's cutback on production, and an end to the Cold War, delays have arisen."

Jan 1, 1994

By W. G. Worcester

THE chief object in presenting this paper on bentonite is to attract general attention to one of the least known and most neglected industrial minerals of Canada, a mineral of extensive commercial possibilities, as evidenced by its development and exploitation in other countries, principally in the United States and Germany; countries upon which some of our most important Canadian industries are dependent for their supply of special clays, such as Fuller's earth and activated bentonite. It was in view of these conditions and the knowledge of the presence of large quantities of bentonite in the Canadian West that the work here reported on was under-taken. While the investigational studies have not been completed, the results indicate that among the many samples tested there are some of real merit and industrial importance. While it has been the author's privilege to examine deposits and samples of bentonite from all of the Western provinces, his chief activities in the field and in research studies have naturally been confined largely to Saskatchewan materials. Notwithstanding this, however, this paper is presented in the hope that it may stimulate a more general interest in, and study of, Canadian bentonites as a whole, a matter of such importance that provincial boundary lines should be overlooked. GENERAL CHARACTERISTICS OF BENTONITE According to Schurecht and Donda (1), bentonite is commonly composed of the mineral montmorillonite, (Mg,Ca)O.Al203 5SiO2 .nH2O, though some are known to consist of beidellite, Al2O3 3SiO2.nH2O: where n is approximately 4, and in certain cases the Al2O3 may in part be replaced by Fe2O3. For present purposes, however, it may be sufficient to state that, in general, bentonite is a "soapy' clay-like mineral or group of minerals composed largely of hydrous aluminium silicate, but frequently carrying small percentages of the alkalies, alkaline earths, and free silica, as well as silicates, gypsum, and other admixed detritus such that purification is usually necessary.

Jan 1, 1937

By D. F. Sangster

"Eight volcanogenic massive sulphide districts in Canada and Japan, ranging in age from Archean to Miocene, were selected for analysis. The study revealed that: {I) each district has an average diameter of 32 km; (2) the districts contain 4 to 20 deposits each, with an average of /2 deposits; (3) the average total base metal content per district is 4.6 million tonnes, with a coefficient of variation of32%; (4) ranked in order of size, the largest deposit in each district contains, on average, 67% of the total metal and the second largest about 13 %. The remaining deposits range downward in decreasing proportions. Empirical observations such as these may be used f or evaluating the undiscovered potential of new districts as well as contributing to a better understanding of the distribution and genesis of this important class of base metal deposits. Previous WorkThe characteristic geological features o f volcanogenic massive sulphide deposits have been described by numerous authors over the past fifteen years and it is now widely accepted that these deposits comprise a major ""deposit-type."" In addition to these geological parameters, however, it is now becoming apparent that volcanogenic massive sulphide deposits show several distinctive grade-tonnage relationships. For example, Singer et al. (1975) demonstrated a slight negative correlation between copper grade and ore tonnage in their sampling of 146 volcanogenic deposits. In a study of 110 Canadian Precambrian deposits of this type, Boldy (1977) showed that the size of the modal deposit is 0.39 million tons, that of the median deposit is 1.40 million tons and the arithmetic average is 6.85 million tons. Moreover, Boldy's data show that the average grade of combined base metals (Cu + Zn + Pb) for each of the se three deposit-sizes is, respectively, 6.25%, 5.35% and 6.47%. Because the lead content of Canadian Precambrian volcanogeni c deposits is most frequently less than 1%, the total metal grade is effectively Cu + Zn . From Figure 2 in Bold y (1977), it can be seen that 83% of the deposits fall in the range of 0.l to 10 million tons ."

Jan 1, 1980

By R. M. Giegerich

"This paper begins with a brief history of Madawaska Mines, a brief description of the geology, mining methods and equipment, and the historical development of 7/8-in. tapered drilling equipment. It then outlines the performance characteristics and inherent problems of the traditional 12-degree, tapered connection, blade-type insert bits. Service-free (Unipass) button-type insert bits are presented as a possible solution to these problems. The paper concludes with an analysis of button bit performance to date.Brief History of Madawaska MinesMadawaska Mines Limited, near Bancroft, Ontario, is in Faraday Township, Hastings County, approximately 100 miles northeast of Toronto and 100 miles west of Ottawa.The mine re-opened in 1976committed to 1,500 tons per day to provide approximately 600,000 pounds of yellowcake (U3O8) per year to A.G.I.P. (an Italian government agency) under the joint ownership of Federal Resources Corporation, Salt Lake City (51%), and Consolidated Canadian Faraday Limited (49%). At present, the mine is producing approximately 2 pounds per ton (.1% U3O8) at the rate of 60,000 pounds per month. The mine employs about 400 employees.Brief Description of Geology, Mining Methods and EquipmentGeologyThe Grenville metasediments of the mine area are chiefly paragneiss, amphobolite and some carbonate (marble) rocks, which were intruded by a basic rock, altered to a metagabbrometadiorite. These formations are generally concordant, striking east to northeast and dipping 45 to 60 degrees to the south. Locally, the metagabbro is often discordant. The above group of rocks (mixed hybrid gneisses) forms the hanging wall of the south-dipping Faraday granite sheet and was intruded by nepheline syenite, syenite and granite. The ore occurs in the granite-syenite-pegmatite dykes that are probably genetically related to the Faraday granite. The ore shoots are erratically distributed within the pegmatite bodies, mainly on the hanging or footwalls but also medially. Their size varies from 50 to 350 feet in length, 5 to 40 feet in width and 50 to 200 feet in height. The ore occurs most frequently in a medium-grained granite or syenite, well hematized with 10 to 30% mafic crystals of pyroxene, amphibole and/or relics of the host rock."

Jan 1, 1982

By Gerhard Heinrich

"The composition and the possible utilization of a slag sample from a nickel smelter operation in the Sudbury district were investigated. The sample containing 0.136% Ni, 0.10% Co and 0.29% Cu could be characterized as a fayalit e slag with a sulphide bearing glass matrix. While the fayalite contains most of the Co, Ni is enriched within both sulphides and fayalite. Cu is only present in higher amounts in sulphides or metallic droplets. This, as well as the small grain sizes of the mineral phases, inhibits significant enrichment by mechanical methods. However preconcentration is possible by means of magnetic separation. An effective hydro metallurgical approach proved to be oxygen pressure leaching with sulphuric acid. Accordingly, the method proposed for processing the investigated slags is to combine the normal utilization as coarse construction material with pressure leaching of the fines and the ground upgraded magnetic fraction. Besides the recovery of base metals, the leaching process yields fine-grained residues, which, after addition of CaD, will set to a cement-like material, or can be used as a pozzolanic cement colour. INTRODUCTIONDuring the pyro metallurgical processing of nickel ores, the valuable metals nickel, cobalt and copper are enriched within the sulphide matte, similar to the smelting of copper ores. The resulting waste materials are slags, which may still contain valuable metal contents of up to 0.5% or above . These contents are close to the cut-off grades of some nickel-cop per-cobalt deposits. The methods that are used to recover copper from slags comprise in general further pyro metallurgical treatment, often after enrichment by flotation. Another area of application for such slags is their use as construction material. However, this might cause environmental problems if nickel containing dusts are produced.Based on one example, the characerization and a possible way of complete utilization of nickel smelter slag s as construction materials in combination with partial base metal recovery are described in this paper. The investigated slag sample was obtained from a dump site in the Sudbury district, Ontario. Detailed test series for process optimization were not possible during this study due to the limited available sample size."

Jan 1, 1989

By D. A. Whitley, R. P. Carone

"Shaft sinking by blind shaft drilling offers the mining industry a different approach to shaft construction. Benefits of blind drilling include less construction time and fewer safety problems while being economically competitive with conventional sinking methods under certain conditions in soft water-bearing formations.Technology and equipment are available today to drill blind shafts up to 20 feet in diameter and over 3,000 feet deep. This is an extension of proven techniques developed by private industry. To drill a 20-foot-diameter hole requires a specially manufactured hydraulic drill unit with a rated hoist of 2 million pounds and 500,000 ft/lb of torque.This brief paper will discuss the following aspects of blind shaft drilling:• benefits;• equipment f or drilling;• circulation methods;• circulation rates;• shaft alignment; and• progress rates to be expected.IntroductionShaft sinking by blind shaft drilling offers the mining industry an improved alternative method for shaft construction. The drilling of 20-foot -diameter shafts to depth s in excess of 3,000 feet can be done by utilizing equipment specifically designed and manufactured for this purpose. This equipment is a distant cousin, at best, to standard oil well drilling rigs. The actual configuration of these drilling machines will be discussed in a following section.To most people, the drilling of mine shafts may seem to be a new or at least recent technology. However, as early as the I920s, German and Dutch miners were drilling 24-foot diameter shafts for coal mines. What led ""miners to resort to drilling were ground conditions that still plague conventional shaft sinkers today: soft , water-bearing formations. The major drawback to these early attempts was a low penetration rate due to poor bit cleaning. This, however, was not the end of early shaft drilling activity. From 1936 to 1950, a number of mine shafts were drilled by the calyx or coring method. This method employs a bit with a large bore down the center. As it drills, it cuts a core. When 5 to 10 feet have been drilled, the core is removed. The disadvantages of this method were the time and work involved in breaking the core free, and it was inadequate in soft water-bearing ground."

Jan 1, 1981

By J. H. Ambrose

Introduction The copper smelter of the Hudson Bay Mining and Smelting Company, Limited was constructed to handle the concentrates produced by the concentrator, the residues from the zinc plant, concentrates from Sherritt-Gordon Mines, Limited, and other custom material that might be available. Construction work was started in 1929 and the first concentrates were smelted in November, 1930. From that time until June, 1931, the operations were more or less of an experimental or tuning-up nature. In designing the smelter, it was believed the following tonnages would be available for smelting per day: Copper Tons Content H. B. M. & S. Co. concentrates 331 14.0 per cent H. B. M. & S. Co. zinc plant residues 77 2.5 per cent Sherritt-Gordon concentrates 200 20.0+ per cent Consideration was given to wet-smelting the charge, but, in order to give greater flexibility to the operation, it was finally decided to roast the concentrates before smelting. Briefly, the smelter equipment consists of three Nichols-Herreshoff roasters, one reverberatory furnace, two Peirce-Smith converters, a Walker casting wheel, and a four-unit Cottrell plant, two units' for the roaster gases and two for the converter gases. A general plan of the smelter is shown in Figure 1. Prior to the commencement of smelter operations, a stockpile of about 9,000 tons of concentrates had been built up by the concentrator and the old pilot-mill. The zinc plant was in operation, but the equipment for drying the residues at the smelter had not been completed. Sherritt-Gordon did not commence their shipments of concentrates until April, 1931. The only flux available was barren sand from a pit ten miles south of Flin Flon, adjacent to the railway. Several small deposits of metal-bearing siliceous ore had been discovered in the district, but none of these were sufficiently large to be considered as a source of smelter flux. At the beginning of March, 1931, the entire stockpile of concentrates had been consumed, and available tonnage for smelting was limited to the daily production of the concentrator.

Jan 1, 1935

By Jean-Yve Tremblay, Michel Prud'homme

"IntroductionIndustrial minerals include a group of more than fifty minerals which are marketed into almost all the sectors of industry; they are the backbone of many manufacturing industries. Table I shows major end-uses for industrial minerals in manufacturing sectors.At Energy, Mines and Resources, industrial minerals are divided in three major categories: (I) agro chemicals and chemicals, (2) construction mineral s, and (3) ceramics and specialty minerals. A listing of the major groups of minerals is shown in Table 2.Agro-chemicals and chemicals are sold on the basis of their chemical content and purity. Construction minerals or structural minerals are distinguished by their performance as low cost and bulky commodities. Ceramics and specialty minerals are performance-specific and are sold for their property on imparting functions on end-products. Industrial minerals have unique characteristics that are useful in many stages of processing; from crushed size to micronized fillers a single mineral ca n serve a broad range of industries because of its intrinsic properties.The Industrial Minerals Industry in CanadaCanada has a large industrial minerals base and is a world leader in the production of some of them. The total value of product ion of these minerals has grown by 115% over the last 10 years despite the misfortune of asbestos.Industrial minerals have shown remarkable growth in the past 20 years at an average annual rate of 9.2%, compared to 7.8% for metallic minerals. In 1986, the value of industrial minerals rose to $4.9 billion, or 14% share of the Canadian mineral production. Figure I shows the value of production of the mineral industry for the period 1966-1986.The leading segment in industrial minerals segment since 1976 is the chemicals and agro-chemicals sector which has an average annual growth rate of 5.75 % in terms of value of production. During the same period, the growth rate for the whole Canadian mining industry was 1.37% per annum. The chemicals and agro-chemical s sector increased its contribution to the total value of production of 2% to 6070 despite the growing share of fuel minerals with 59%."

Jan 1, 1988

By T. Robson

"A multi-seam underground mining operation employing mechanized room-and-pillar methods provides 35 per cent of Smoky River's annual ROM coal production (2 million tonnes) from the Smoky River coal field. Surface mines provide the balance. Tectonic activity, which has resulted in deformed coal seams and a rugged relief to the area, has provided natural boundaries to discrete mining areas. Coal transportation to the central preparation plant incorporates a 3 km long underground conveying system. The company is committed to efficient resource recovery and a safe and healthy working environment. Introduction The Smoky River Coal operations (SRC) are centered in the Smoky River coal field located in west central Alberta, some 360 km west of Edmonton and 20 km from the Town of Grande Cache (Fig. 1). Combined mining operations produce 2 million tonnes per year run-of-mine coal from which 1.65 million tonnes of high quality, low volatile, coking coal are obtained. The underground mining operation, employing mechanized room-and-pillar methods, provides 35 per cent of SRC's annual ROM coal production (2 million tonnes). Surface mines provide the balance. Tectonic activity has resulted in deformed coal seams and a rugged relief in the area. This has provided natural boundaries to discrete mining areas. Background Exploration of the Smoky River coal field took place over four decades commencing in 1926. McIntyre Mines Limited (parent to Smoky River Coal Ltd.}, commenced development of the property in 1968 having secured a long-term contract for the delivery of coal to Japan. Production of underground coal at Smoky River commenced toward the end of 1969 and has been continuous on the property since that time. Originally, two underground mines were opened. Original planning envisaged the use of room-and-pillar techniques to develop blocks of coal for subsequent extraction by longwall retreat mining. Following initial development, two longwall units were established, the first in November 1970 in No. 5 Mine and the second in February 1971 in No. 2 Mine. Difficulties were experienced on both units, and production commitments could not be achieved. The longwalls were discontinued, and mechanized room-and-pillar was subsequently used for all underground mining."

Jan 1, 1985

By Gérald Riverin, Bernard Boily, Daniel Bernard

"The Donalda gold deposit is located in the southern part of the Noranda camp, approximately 2 /em east of the Horne deposit. Combined past production and current reserves total over 1.5 Mt grading 6.9 grams of gold per tonne.The deposit consists of two main veins which are both narrow but persistent structures. The No 1 Vein, hosting about 80% of the gold ore, consists of an EW-striking banded quartz vein dipping on average 25° to the south. Although variable in thickness, it averages lessthan1 metre. TheNo2 Vein, located 300 m below the No 1 Vein, is similar to it except that it dips 50° south. Gold occurs mostly as fine native gold disseminated in quartz, with 2-3% pyrite and trace amounts of various sulphides and tellurides. The wall rocks are only weakly altered and mineralized with pyrite, hematite and sericite.Structurally, the Donalda deposit is centered on the Horne Creek fault (HCF) where the HCF is intersected by the ENE-trending DonaldaandCentralfaults.NorthoftheHCF,theNo1Veinishosted by the Quemont Rhyolite, whereas south of the HCF the vein is hosted by magnetic basalts and diorites of the Wilco Andesite Formation. The No 1 Vein extends across the HCF with only minor displacement, whereas the No 2 Vein has not yet been traced south of the HCF.A preliminary structural analysis of the mine shows that the HCF, which strikes 275° and dips 70° north in the mine area, has a reverse movement a lid that the flat dipping No 1 Vein is a tension vein. Both probably developed in response to the same stress conditions. Although the absolute age of the Donalda veins is not known, structural relationships indicate that they formed during the latest movements of the HCF, but prior to the intrusion of the Proterozoic D' Eldona Diabase. The occurrence along the HCF of gold-rich synvolcanic massive sulphides and much younger vein-type gold deposits (such as Donalda) indicates that the HCF was a very long-lived, deep-seated (?)structure which has controlled the circulation of gold-rich fluids over a long period of time."

Jan 1, 1990

By G. M. Proudfoot

IN June, 1940, a .Longyear 3420 blast-hole drill rig was purchased by the Hudson Bay Mining and Smelting Company for the purpose of experimental blast-hole drilling. The early work showed that the diamond drill could compete successfully with the Leyners then in use and by January 1st, 1941, the diamond drills were placed on contract. They may be considered to have gone into production at that date. Prior to the introduction of the diamond drill, our bench drilling was clone with Ingersoll-Rand S-70 and S-83 Leyner drills. Holes up to 40 feet were drilled, using coupled 1 1/4-inch round steel. Bits with a gauge of 3 inches were used in starting these holes, with a drop of 1/8-inch gauge for each 2-foot run. The holes normally ended with 1 1/2 or 1 5/8-inch gauge, though bits down to 1 1/8 inches were available for holes longer than normal or for holes drilled in abnormally abrasive ground. This method of drilling was very satisfactory and was a distinct improvement on the three-bench method of drilling used previously (1). During 1937 and 1938, when experienced miners were available, Leyner long-hole drilling was carried on with great efficiency. However, in 1939 and 1940, labour turnover in-creased and experienced long-hole drillers became scarcer. This resulted in a marked decline in the efficiency of the drilling operations, with increased costs and less tonnage drilled per shift. EXPERIMENTAL WORK The initial diamond drilling test work was carried on with a Longyear 3420 air-driven blast-hole drill rig, drilling a 1 3/16-inch hole. Tests indicated that diamond drilling could be successfully adapted to local conditions and that overall costs compared favourably with costs obtained with the Leyner long-hole method. Leyner long holes on stope benches usually bottomed at a larger diameter than the 1 3/16 -inch diamond-drill holes. More powder could therefore be introduced into the Leyner holes than could be placed in the diamond-drill holes. To achieve the same fragmentation, additional footage was required when diamond drilling. Diamond-drill holes of larger diameter were indicated.

Jan 1, 1943

By R. H. McCorkell

"Sample drops were evaporated on Lexan disks on which fission tracks were etched and counted after neutron irradiation. [U] = 0.01 ppb could be measured. This method is rapid, cheap, simple and accurate when sufficient control and checking is possible, but several types of error were encountered in its use, the natures and sources of which were investigated but not always determined. The most troublesome were ""annealing"" processes which sometimes altered the fission tracks so that they were not fully revealed, or acquired varying appearances, after etching. Several different annealing processes seemed to exist and seemed to be produced by alteration of the Lexan surface by dissolved components in the solutions and, perhaps, by radiation and local heating during neutron irradiation. The appearances of the etched tracks varied with the natures of the droplet solutions even when annealing appeared to be absent, and for this reason, standards are best prepared by additions of known amounts of U to sample solutions. Contamination of sample droplets by U-containing dust from the air is best avoided by preparing the disks away from areas where other U-analysis is being carried out.IntroductionFission tracks are regions of radiation dam age produced in solids by the fission of atoms such as 235U. They can be enlarged to sizes visible through optical microscopes by etching surfaces through which they pass with solvents in which the solids are slightly soluble. A sample to be analyzed for U may be placed against a film of a solid of negligible uranium concentration (the track detector) and irradiated with thermal neutrons in a nuclear reactor to fission 235U. The track detector is then etched and the number of tracks per unit area, ?, which is proportional to the uranium concentration of the sample, c, and the neutron do se, n, is found by counting through a microscope. Uranium in solutions ha s been measured by evaporating drops on track detectors, irradiating, etching and determining the number of tracks per drop, T(2,3,4,5). Extensive test s of this method are reported here."

Jan 1, 1983

By F. Goodarzi

"The nature and distribution of the inorganic components and the concentration of certain environmentally sensitive elements, such as, As, Sb, B, Hr, Cd, CI, F, Mo, Pb, S, and Se, or elements detrimental to the utilization by industry such as B, Na, P and S are very important in determining coal quality. The concentration of elements in coal is controlled by the following factors: (1) frequency of occurrence of partings; (2) nature and habit of occurrence of minerals in coal; (3) environment of coal deposition; (4)rank; (5) geological setting; and (6) weathering.Selective mining of coal can reduce the input of undesirable elements such as As and S by eliminating the part of the seam which is enriched in these elements. IntroductionCoal is a sedimentary rock of plant origin and consists of two components: an inorganic component (minerals) which is formed by influx of clastic sediments into the coal-forming environment and also by activities of groundwater, and an organic component (macerals) which is formed from plant debris deposited in the original coal-forming environment. However, coal is a generic term with a broad range of utilization characteristics. A coal that is attractive for use in one market - such as generation of electricity (combustion) - may not necessarily be acceptable for use in another, such as for production of coke (Davis, 1978).The elements in coal are associated with both macerals and minerals. The elements associated with coal macerals, such as C, H and N are termed elements with organic affInities. The concentration of elements with organic affmities (including B, Br, CI, F, I, S, and Mg) often decrease with increasing ash content of coal (Nicholls, 1968; and Finkehnan,1982). FIgure 1 shows such a trend for chlorine. Some of these elements may also be incorporated in coal maceral as very fine mineral grains (Finkelman, 1982). The elements with organic affmities are often volatilized during combustion and carbonization of coal and emitted in the vapour phase. Emission of some of these elements (C, 0, Br, N and S) is of major environmental concern (Lim, 1979)."

Jan 1, 1994

By David Deglon, Martin Harris, Mehdi Safari

"This study investigates the effect of energy/power input on the flotation of a platinum ore in a pilot-scale oscillating grid flotation cell. Oscillating grids generate near ideal hydrodynamic environments, characterised by turbulence which is relatively homogeneous and isotropic. Secondary rougher feed and primary cleaner tail streams were floated in a pilot-scale oscillating grid flotation cell at power inputs from 0 to 2.5 W/kg, using 0.50 and 1.50 mm bubbles. From this study one may conclude that the effect of energy/power input on the flotation rate is strongly dependent on the particle and bubble size. Flotation rates generally increase with increasing particle size and decreasing bubble size. Increasing energy input generally leads to an increase in the flotation rate for finer particles and an optimum flotation rate for intermediate & coarse particles. Optimum conditions for intermediate/coarse platinum ore particles are using small bubbles at low energy inputs, or large bubbles at higher energy inputs. These results suggest that higher energy inputs are generally beneficial for the flotation of platinum ores as these consist predominantly of finer particles (-75 µm) and most industrial flotation applications operate with larger bubbles (1.0-2.0 mm).INTRODUCTIONThere is a considerable body of experimental and theoretical evidence to suggest that energy/power input plays an important role in flotation kinetics, particularly in the finer particle sizes where flotation efficiency is poor (Deglon, 2005). A number of excellent studies into the effect of energy on flotation kinetics have been carried out in impeller stirred cells (Ahmed and Jameson, 1985; Deglon, 2002; Pyke et al., 2003; Newell and Grano, 2006). Investigations using this type of cell have a number of limitations. The impeller influences particle suspension, bubble break-up and agitation in the cell. In addition, turbulence is highly inhomogeneous and anisotropic, with power intensities near the impeller orders of magnitude higher than those found elsewhere in the cell (Deglon, 1998; Koh and Schwarz, 2003; Schubert, 1999, 2008). These limitations have resulted in the development of novel cells for the investigation of energy input."

Jan 1, 2016