Search Documents

By T. T. Tigert

INTRODUCTION THE PURPOSE of this article on methane is to present the simple basic facts of the characteristics, detection, occurrence, and control of this natural gas in metal mines, so that operators and .safety men may deal with it with some degree of confidence. Fortunately, although methane has been noted at many metal mines throughout Canada, the occurrences usually are small flows of short duration, and the ignitions experienced have, in most cases, had minor consequences. However, methane does present a hazard that can be of serious proportions, and ?any complacent attitude toward this gas that might exist in the minds .of people engaged in underground mining operations should be dispelled when it is ,pointed out that, in Ontario alone, ignitions of this gas have resulted in at least four fatalities, and eight men seriously injured and burned, during the past three years. lt is well-known that liberation of methane can and does .occur suddenly, often in mines where no gas has been found before, and that many ignitions and explosions have occurred in so called non-gassy mines. In view of this, it is apparent that no mine can, with any degree of certainty, be classified as non-gassy. CHARACTERISTICS OF METHANIZ The characteristics of natural gas encountered in most of the Ontario mines indicate that methane is the chief constituent. Methane (CH4), commonly termed 'marsh gas', and by the coal miner 'firedamp', is a colourless, odourless, and non-poisonous gas considerably lighter than air, and is the explosive gas frequently found in coal mines. Its physiological effects, closely resemble those resulting from nitrogen.

Jan 1, 1951

By Owen Matthews

THE property of Cochenour Willans Gold Mines, Limited, is in Dome township, Red Lake area, northwestern Ontario, a hundred miles north of Kenora and close to the Manitoba border. 'Planes of Canadian Pacific Air Lines, Limited, serve the district daily, except Sundays, from Kenora, Ontario, and Lac du Bonnet, Manitoba, passengers being trans-ported from Winnipeg to Lac du Bonnet by car. Freight is handled by the Patricia Transportation Company. It is brought in from Hudson, Ontario, on the Canadian National railway, in summer by water and in winter by tractor trains over the frozen lakes. The first staking on the property was done during the 'rush' of 1926 and 1927 by William M. Cochenour, and on April 30th, 1928, a syndicate was formed, taking in claims of W. M. Cochenour and Dan Willans and H. G. Young. Surface trenching was carried on almost continuously until 1934. In October, 1932, Ventures Limited took an option on the property, and in December of that year they carried out an extensive diamond drilling programme, but considering the results did not justify further work, they relinquished the option. In July, 1934, further surface work was carried out under the supervision of W. P. Mackle and, on the strength of his report, Noah Timmins Incorporated took an option on the property and agreed to make an initial expenditure of $100,000 on development work. This was carried out under the direction of Hollinger Consolidated Gold Mines and included the sinking of a shaft to a depth of 165 feet, establishment of a level at the 150-foot horizon, and some 1,900 lineal feet of drifting and cross-cutting. The results obtained from this work were not considered of sufficient promise and the option was terminated in August, 1935. Mr. Cochenour then approached Alex Mackenzie and associates, and the present Company, Cochenour Willans Gold Mines, Limited, was organized on April 9th, 1936, and sufficient money was raised to continue operations. Under the direction of W. P. Mackle, the shaft was deepened and a second level was established at the 275-foot horizon; but, due to several adverse reports by outside engineers, it was found impossible to raise sufficient money to carry on the work and operations were suspended in December, 1937. In January, 1939, an agreement was entered into with Gold Eagle Gold Mines for the milling in their plant of some Cochenour ore. A total of 4,959 tons was treated, and the results were so encouraging that further finances were raised by C. A. Gentles & Company, of Toronto. Following further development in the mine, it was decided to erect a mill. Construction was started on September 1st and the mill was turned over on November 29th. The first gold brick was poured on December 23rd.

Jan 1, 1942

By F. L. Thompson

MAN has been using lubricants of one type or another for at least four thousand years, and probably for a much longer period. We have learned from the inner wall of the tomb of Tehuti-Hetep, on which is a decoration illustrating the method of moving a great stone statue, that as early as 1650 B.C. lubricants were used to facilitate the movement of heavy loads. The Egyptians had available nearly every animal oil and fat, as well as at least thirteen vegetable oils, of our time, all used for food products and in the arts. Examination of the deposit on the axle of a chariot from an Egyptian tomb of 1400 B.C. showed it to consist of animal fat, quartz, sand, iron, and compounds of aluminium and lime-indicating a manufactured lime-soap grease made with lime-soap and possibly tallow. This would not be a bad lubricant for the axles of chariots which had to travel over sandy soil in hot weather. As to the method of grease application, it is remarkable that, from the time wheels were first used in the far distant past up to the present day, no one has ever thought it necessary to design a seal to keep the lubricant in and the dirt out. While time certainly marches on, we have today, in waggons throughout the world, exactly the same form of lubrication as was used in the earliest chariots. From the early days of the chariot until shortly before 1800 A.D. (when the steam engine was introduced), the main heavy equipment required for mining purposes, for machines of war, etc., was made of wood. The bearings were rough, but usually they had metal in the lower half. Light machinery was made of bronze, brass, or iron. Large clearances and poor alignment, plus the fact that speeds were much below present-day demands, combined to provide conditions that were quite satisfactorily met by the lubricants available. These were often produced locally by the machine operator. In northern countries, lard oil, prepared from the fat of hogs, was the universal lubricant employed in machine shop, as a belt and leather dressing, and for cutting operations. Sperm oils were used for light machinery, whereas beef and mutton tallow were more generally employed for the largest bearings. For the early steam cylinders, a slug of heated beef tallow was injected wherever the cylinders gruntingly complained of a lack of lubricant. Southern countries, however, resorted to the use of vegetable oils-olive, cocoanut, etc.

Jan 1, 1940

By K. Matsuura, M. Ohno, T. Tokunaga

"Recently, weight saving of automobiles has been strongly required to solve the global environmental problems. Mg alloys are one of the promising candidates as a new material for automobile components. However, the application of the Mg alloys is limited due to their poor corrosion resistance. In the authors’ preliminary study, an Al/Mg/Al clad sheet was fabricated by hot extrusion and subsequent hot forging. The clad sheet exhibited a high corrosion resistance and excellent super plasticity. More specifically, the elongation of the clad sheet was as high as 550%.Importantly, despite that the elongation of Al itself was only 40%without cladding with Mg, the ductility of Al was greatly improved in the form of the clad sheet. In the present study, in order to clarify the mechanism of the superplastic deformation of the Al part in the clad sheet, the microstructure evolution of the Al during the tensile de formation is investigated. The tensile tests were conducted at a temperature of 573 K and at a strain rateof1.0×10-3s-1. The Al exhibited fiber-type texture at the beginning of the tensile deformation, and brass-type texture after the elongation of 50%.INTRODUCTIONMg alloys have been gaining attention because of their excellent properties e.g. light weight, good recyclability and high electromagnetic shielding. However, their practical applications have been limited due to their poor corrosion resistance. In the authors’ previous study, in order to solve this problem, Al coating on Mg alloy by hot extrusion has been proposed, and the Al-coated Mg alloy plate was fabricated(Tokunaga, Matsuura, &Ohno,2012). The plate was hot-rolled and an Al/Mg/Al clad sheet was fabricated. No cracks or debondings were observed at the Al/Mg interface. The sheet exhibited a high corrosion resistance in a comparable level as pure Al. Also, the sheet exhibited a large and uniform elongation of 550%in a tensile test at a temperature of 573 K and at a strain rate of 1.0×10-3s-1without any breakings or debondings of the Al part(Tokunaga, Matsuura, &Ohno,2014).Mg alloys have been known to exhibit a superplastic characteristic when the grain size is small. However, it is known that the pure Al does not generally exhibit the super plasticity. As mentioned above, the Al-coated Mg alloy sheet exhibited 550% of elongation in the authors’ previous study (Tokunagaet al., 2014). However, the pure Al itself showed only 40% of elongation, while the Mg alloy itself exhibited 800% of elongation. Thus, it has been demonstrated that the pure Al was super plastically elongated when it was bonded with the Mg alloy. Generally, deformation mechanism of the super plasticity of the Mg alloy is known as a grain boundary sliding. On the other hand, the mechanism of the superplastic elongation of the pure Al is still unclear. The purpose of the present study is investigation of the deformation behavior of the Al part in the Al/Mg/Al clad sheet during tensile de formation."

Jan 1, 2018

By V. L. Eardley-Wilmot

Introduction We are all familiar with the bright and attractive colours of the roofs of our suburban houses and country cottages, but few of us fully appreciate the complex and highly technical nature of the business which produces these roofings. Roofing granules consist of small broken particles of certain types of rocks, burned clay products, or slag. They are used either in their natural state or artificially coloured and are affixed onto one surface of asphalt sheeting. Their main functions are to protect the roof against external fire hazard, prolong its life, and enable it to withstand extremes of weather over a period of years without deteriorating. Pleasing designs and colours are added attractions. The reverse, or inner side, of the sheeting is coated with a film of talc or fine mica to prevent it sticking when rolled up for convenience of handling for transport. The finished sheets or strips are cut into a variety of shapes for roofing shingles, or for sidings which resemble rows of bricks separated by mortar. In Canada, the consumption of roofing granules has increased fourfold in the past decade and indications are that the industry will continue to expand. At present, fourteen plants are engaged in manufacturing granule-coated roofings, but, apart from a relatively small output of natural granules by two producers in British Columbia, only one eastern producer makes both natural and coloured granules in Canada, and these only for its own use. The annual consumption is now over 77,000 tons of roofing granules valued at more than $1,600,000, but about 62 per cent of the tonnage is imported. It is thus apparent that the production of Canadian-made granules, both natural and coloured, could be considerably increased. This paper is presented in order to give an outline of the industry and at the same time to point out its difficulties and pitfalls.

Jan 1, 1946

By T. Patching

"COAL MINING at Canmore, Alberta, began in 1886 and continued there until July 1979. A comprehensive review of the Canmore operations is not available, but the following notes have been prepared to accompany references where more complete information about the operations and problems encountered may be found. The coal measures at Canmore occur in the Kootenay formation of Lower Cretaceous age, and at least twelve seams of varying thickness and horizontal extent have been recognized. The coal, of high quality, ranges from semi-anthracite in the lower seams to low-volatile bituminous in the upper seams. The strata of the Canmore area form an asymmetric syncline, the western limb being overturned to vertical by the Palaeozoic strata of the Mount Rundle Range. The eastern limb is more gently sloping, its dip ranging from IO to about 40 degrees. Tectonic movements have created a troubled and complex coal field. The seams were initially not regularly or uniformly deposited, and mountain-building activity folded and faulted the strata and in many places sheared and crushed the original structures of the coal(l,2,3,4,5). The lack of continuity of the seams, and changes in thickness and inclination, made longrange mine planning difficult and necessitated flexible methods of work, and the local disturbances sometimes created problems in roof support and contributed to the occurrence of sudden outbursts of coal and gas. Almost all of the coal was produced from underground operations but a few relatively small V-cut mines were worked along surface outcrops. These have all been filled in and revegetated. The basic method of underground work was by room-and-pillar, but many variations in layout and sequences were employed. Gangways were driven along the strike and pillars were formed by raises or slopes and crosscuts. Pillars were recovered to the extent that ground conditions permitted by splitting or taking 'skips' from the sides of the raises or splits or crosscuts. Because of the variety of conditions that existed in the field a considerable range of cutting, loading and support methods were employed, including hand-loading, slushers and scrapers, duckbills and shaker conveyors, mobile loaders and continuous miners. Except for the coal obtained by continuous miners most of the coal was broken out by blasting when driving narrow working places or by air picks when cutting from a face. In the early years hand loading onto cars or conveyors predominated. A continuous miner was introduced into Canmore in 1959 on a trial basis, and another was subsequently acquired, in order to minimize blasting and hand loading, but problems in operating these machines in inclining seams and difficulties in providing safe roof support for the machines and workers restricted their efficient use. Another major difficulty with the use of continuous miners was the problem of ventilating development headings in the very gassy conditions. In the early 1970s they were replaced by electric and compressed air slushers and scrapers, resulting in a general increase in productivity, roof control, safety, and flexibility of operations(1.6,7,8.9)."

Jan 1, 1985

By Norman A. Jull

"IntroductionThe scope of this paper is intended to deal mainly with cyclones in grinding circuits. However, the section on Cyclone Equipment, contains more general descriptions, to cover most of the cyclones used in mineral separation plants.Cyclones as heavy media separation vessels are discussed by Mark Freberg (this volume). Their use in tailings deslirning is discussed by P.C. Lighthall, eta/. (this volume).Cyclone EquipmentCyclone TerminologyApexObviously, the apex of a cone is the smallest end. However, with cyclones the term is usually used for the fitting which is attached there to control discharge of the underflow stream.Note that the cyclone need not be so mounted that the apex is the lowest point, although that is the most common method. Keeping the apex below the rest of the cyclone helps avoid plugging when the underflow contains very large particles and the tonnage of underflow solids can fluctuate.Various devices for adjusting apex diameter or otherwise controlling the underflow are discussed later under ""Cyclone Construction"".BypassingThe transport of a portion of feed slurry directly to underflow so that classification is bypassed by that portion of feed.This action is fundamental in any hydraulic classification, whether gravity or centrifugal, because the large particles going to underflow have some void volume between them, which remains filled with feed slurry.D50By definition, D50 is that particle size which has an equal probability of going to overflow or to underflow.It is calculated by drawing a curve of recovery vs particle size and noting the size at which 500Jo recovery occurs.This may be done equally with corrected or with gross recovery (see below) and the resulting value should therefore be identified as d50c (corrected d50) or as d500 (gross d50).Note also that it is common in industry practice to plot size analysis data on the retained size rather than on the average of passing and retained (the middle of the size band). To portray sizing alone it makes no difference, but to read a D50 value off a curve of corrected or gross recoveries, it is essential to plot those values on the average size of each band."

Jan 1, 1989

By Karl J. C. Harries

"11.1. INTRODUCTIONA royalty is the most common form of interest retained by an owner who is disposing of his property or by a joint venturer who ceases, voluntarily or otherwise, to participate in the joint venture.1 Rarely does a non-participating party have a direct interest in the property and/or assets of a project, except during the period when a non-participating joint venturer may be having his interest diluted.2 In fact, it is most unwise to permit a non-participating party to have such an interest. It may give rise to significant problems that could prevent or substantially delay the development of an orebody and require a considerable amount of time, effort, and money to correct.3In negotiating a royalty, it is important to be aware that there is no law or convention that says, “You just don’t do that in a royalty agreement.” As with any agreement, inventiveness and flexibility on the part of the negotiator are not just assets, but virtual necessities. It is also for this reason that it is essential to read the royalty provisions, particularly the definitions, fully and carefully, especially if you are the recipient. On the other hand, if your inventiveness and flexibility cause your accountant to question the advisability of one of your ideas, pay close attention to him.4 He will, no doubt, have had experience with many royalties, and no competent professional adviser disparages a client’s ideas without good reason. Besides, he will probably be the person who will be determining the royalty or certifying the determination of it. At the very least, he will be auditing your general financial statements, which must, of course, reflect the royalty determination. Heeding his advice not only may make the determination easier but also will probably save you money in professional fees later on. Nevertheless, do not simply accept his comments or recommendations without understanding what they mean, why they have been made with respect to your royalty agreement (sometimes professionals have comments that they make as a matter of course and, even though they may be valid, you should require them to be justified to you), and how they will affect your royalty determination.It is surprising how many corporate payors insist that their standard royalty provisions5 be used as a matter of policy, thereby effectively tying the negotiator’s hands. It is even more surprising, and to professional advisers frightening, how many of the senior employees who insist that such a policy be adhered to, virtually without variation, do not know what is in the standard provisions and would be hard-pressed to explain them to a property owner. If standard provisions are appropriate, by all means use them, but do allow experienced and sensible negotiators some flexibility and make sure that everyone within the corporation who may have to use such provisions – negotiator and supervisor alike – knows what they contain and why they are as they are. For example, there are often good and convincing reasons to describe a deduction in a certain way or to include a certain specific type of deduction. On the other hand, quite often the answer to the question “why?” is that “it has always been there”; this is, or should be, an unacceptable answer. In any event, if you, as the negotiator, appear to know what you are talking about, are familiar and comfortable with your royalty provisions, and are backed up by your boss with the same words or reasoning, a property owner cannot help but be satisfied that you are knowledgeable, even if he still does not agree with your provisions. Finally, if you do use standard provisions, make sure that they are reviewed and updated regularly."

Jan 1, 2003

By E. A. Sobering

Hydraulic filling or sand filling is a system used to fill underground excavations with a dense pulp. Milltailings are commonly used. The resultant fill, after being placed and de-watered, is known as hydraulic fill. These fills have been effectively used to (a) provide immediate support of the workings place during stoping; (b) .facilitate the recovery of pillars; (c) limit surface damage by subsidence; (d) permit the mining of parallel hanging-wall orebodies; (e) stabilize the ground in mines, thereby reducing the rock burst hazard; (f) control and prevent mine fires; and (g) solve tailings disposal .problems. The various uses of hydraulic filling have been developed during the last ninety years. In 1864 it was employed only to control surface subsidence. Since then, this system has been successfully applied to meet specific ground support problems. Methods of handling mine-fill derived from a source outside the stope include transportation ?Of the material by hand, mechanical, pneumatic, and hydraulic means, or a combination of these methods. The hydraulic method has many advantages not found in the others; a significant advantage is the low compressibility of the material when compared to other mine-fill material. The hydraulic filling operation includes the .preparation and transportation of the pulp. The underground practices in .placing the fill vary from mine to mine. The chief reason, formerly, for using hydraulic fill was the exceptional support it provided, but recently it has proven an important means of reducing mining costs. This is especially true where labour is a large part of an alternative method, or where timber consumption is greatly reduced. The indirect benefits of hydraulic filling are probably as important as the direct benefits, and few important disadvantages exist.

Jan 1, 1956

By J-F. Turgeon, Y. Pépin, G. Hudon

The high temperature hydrochloric acid leaching of titanium slag in the QIT upgraded slag process (UGS) creates a harsh environment for materials of construction. Because of their excellent chemical and heat resistance, furan mortars are a key component in the lining system of the UGS pressure leach reactors. Similar products coming from different manufacturers have however shown unequal performances in service. In some cases, cracking of the brick lining has occurred during curing of the mortar. This study was initiated to investigate the impact of the curing conditions on the properties of carbon filled furan mortars. Laboratory tests were conducted to assess the effects of temperature and exposure to water and HCl solutions on mortar shrinkage and mechanical properties (tensile, compressive and bond strength). Three commercial furan mortars that have been used in the UGS leach reactors were thus characterized. Mortar properties were found to vary significantly among the products tested. Moreover, the curing conditions were shown to have a significant impact on mortar shrinkage and mechanical strength. The experimental findings provide explanations for the unequal performances of the mortars in service and for the problems of brick lining cracking that occurred in the UGS leach reactors. This study demonstrates that excessive mortar shrinkage compromises the integrity of a brick lining system.

Jan 1, 2004

By E. L. Martheleur

A general expression for the electric power generated or consumed in a direct-current circuit is obtained by multiplying the reading of a suitable voltmeter connected across the terminals by the reading of an ammeter connected in the circuit. by the reading of an ammeter connected in the circuit. When alternating currents are considered, the product of the volts and amperes as obtained for direct currents does not generally represent the true energy involved; it is greater than this true energy and should be multiplied by a certain factor which may have any value between zero and unity, according to the conditions of the circuit. This factor is called the power-factor of the system, and expresses the ratio of the true energy over the apparent energy. For a three-phase system the expression of the apparent energy is [\/ 3 x E x I; ] while the expression of the true energy, that is, energy suitable for mechanical work, is [v? 3 x E x I x cos A.] The total current of a circuit corresponds to the reading of an ammeter connected in the circuit; it is the resultant of two currents. The first of these is in phase with the voltage, and is that portion which really produces mechanical work; the second is always in quadrature with the voltage, that is it may lead or lag the voltage by 90° according to the condition of the circuit. This second current does not produce any sensible work, and for this reason is sometimes called wattless current or reactive current.

Jan 1, 1924

By H. Kalsi

"IntroductionRoyal Oak Mines Inc. is currently operating the Parnour No.1 mine located 10.5 miles east of Tinunins, adjacent to Highway No. 101. The operation produces gold ore at a rate of 3300 tpd. A variety of mining methods including bulk open stopes with ITH drilling, narrow vein panel, longhole and open pits contribute to produce 1.2 rnillion ore tons annually. The milling process requires water at an average rate of 1 rnillion gallons per day (700 U.S. gpm). The Pamour No.1 mine has historically had two major problems with regard to water control. The flfSt is the underground pumping system's inability to handle excess inflows of water generated from spring thaws. When the thaw takes place over a short period of time, large inflows of water are created which quickly migrate underground and sometimes result in the loss of the lower loading pocket. During the spring runoff the pumps at Parnour No.1 work continuously for 24 hours at a rate of 1200 U.S. gpm. The system was further taxed in 1991 when the Hoyle mine was connected to the Parnour No. 1 mine and whatever excess capacity the pumping system had for the spring runoff was taken up handling the inflow of mine water generated by the Hoyle.The second problem facing the mine was the lack of water for the mill during the severe cold winter months and dry summer months. At those periods inflows of water into the mine decline and the underground water normally pumped to the mill is severely decreased. In times of dry spells mill personnel increase the water diversion from the Three Nations Lake in order to meet water demands and maintain tonnage throughout. In essence, the mine was going through an annual ""feast or famine"" situation and a solution was needed.A number of options were considered but were narrowed down to the two most practical. The first was to modify the existing pumping system and increase its capacity in order to handle the excessive runoff during spring thaw. This involved a heavy capital expenditure and would have caused additional operating costs (i.e. increased hydro). The option chosen by the management was to divert the spring runoff into an unused section of the mine called The West End. A bulkhead placed at the 2400 level would provide an ultimate underground water storage potential of 150 million gallons. This would also cut down the load on the existing pumping system, save electric consumption in pumping and provide water storage which could be used for milling during the two dry spells. Pumping stored water from the mine could be scheduled for off peak hours to take advantage of lower power rates. To meet these objectives an underground bulkhead capable of withstanding high water pressure up to 2000 ft head (866 psi) was constructed in the spring of 1992. The bulkhead site was at the 2400 level at the Pamour No.1 mine. A major consideration was the requirement to construct a safe bulkhead and to minimize water leakage from the bulkhead.Figure 1 shows the bulkhead location."

Jan 1, 1994

Why A Friction Hoist? B EFORE discussing the equipment, it would be instructive to review the factors that influenced the management of Opemiska Cop-per Mines to purchase a friction hoist rather than a conventional double drum hoist. The mine required that 250 tons per hour be hoisted from a depth of 2,100 feet. They also specified that the top speed should not exceed 1800 feet per minute. Eight ton capacity skips in balance and a maxi-mum hoisting speed of 1725 feet per minute would meet the required <Output on a duty cycle as follows: Loading time Creep away from loading pocket Accelerate (at 2.5 ft/sec2) 10 seconds 5 seconds 11.5 seconds Full speed run Decelerate (at 2.5 ft/sec2) Creep into dump Cycle time Trips per hour Tons per hour 72 seconds 11.5 seconds 5 seconds 115 seconds 31 31x8 = 248

Jan 1, 1962

By H. R. Banks

"AbstractThe rod mill introduced into the Sullivan Concentrator flow in September, 1948, replaced two sets of 74 in. by 20 in. rolls and their accompanying screens. In addition to the work formerly done by the rolls, the rod mill is producing a finer product and one which provides more suitable feed for the primary ball mills. The object of the finer discharge from the •primary mills is to make possible a feed directly from the mills to a series of flotation machines where a large portion of the galena can be withdrawn from the circuit while the galena is still coarsely granular. The interposing of classifiers between grinding and flotation means the selective grinding of the galena.The author summarizes the work of •the rod mill through the first 5,000,000 tons and shows that the wear of lifters, liners, and rods is definitely below normal for the type and size of rock being handled. The rod mill, driven through a fluid coupling, has been subject to a range of speeds from 13.5 to 19. A pinion in the speed reducer was changed to allow the extending of the normal speed range of the fluid drive of approximately 4 r.p.m.Since the critical speed inside the liners •is 23.4, the range •in per cent of criticial is 57.7 to 81.2 per cent.The capacity of the mill increases with increasing speed but the character of the product remains approximately the same except when the feed rate is too high.Consumption of steel, power, and other operating data are presented.The author introduces certain fundamental observations on the principles involved •in comminution in order to support the special design and operation of the rod mill."

Jan 1, 1952

By A. Tabereaux

The reduction of perfluorocarbon (PFC) emissions from aluminum electrolytic cells is a key issue in reducing greenhouse gases at all aluminum smelters, The emission of tetrafluoromethane (CF4) and hexafluoroethane (C2F6) gases during anode effects at smelters can be calculated using two different methods; anode effect-minutes slope and anode effect overvoltage, This work demonstrates that the optimum method depends to a large extent on the type of anode effect kill algorithm utilized at each smelter, CF4 emissions vary widely from aluminum cells depending on the type of anode effect kill method utilized; fast, aggressive anode bridge movements vs. slower, moderate anode bridge movements.

Jan 1, 2005

By J. G. A. Stevenson

"IntroductionDURING the past quarter of a century diamond drilling, in all its phases, has increased in Canada to a very impressive total of more than 10,000,000 feet annually. Since the art of diamond drilling became recognized, some seventy-five years ago, there has been considerable improvement in both equipment and technique. The steam-powered drill rig and the hand-set diamond bit are things of the past.The purpose of this paper is to review some drilling trends •and practices over the past few years. In the preparation of the paper, the author wishes to acknowledge kind assistance of Mr. J. D. Cumming, Dr. C. H. Hitchcock, Mr. Robert H. Quintal, and many others, including staff members of the Longyear organization.Figures provided by the Dominion Bureau of Statistics for the year 1954 show that Canadian production of metallic and non-metallic minerals and structural materials had a value of $1,454.2 million, excluding radioactive products for which figures are not available. Approximately 61 per cent of this valuation comprised production of nickel, copper, gold, coal, asbestos, zinc, lead, and iron. It is certain that diamond drilling played a major role in the discovery and development of the orebodies from which these minerals were obtained.After the surface examination of an ore deposit, diamond drilling will automatically follow, in order to determine structure, grade, and tonnage.Footage drilled by mining companies and contractors on mineral exploration and blast-hole work since 1945, as reported by the Do-minion Bureau of Statistics, is shown in Figure 1."

Jan 1, 1955

By A. K. Ratte, C. M. Brown

"IntroductionIn autogenous mills, grinding of the ore feed is accomplished by the direct action of the variously sized rocks making up the feed on each other and the mill interior. Size reduction is achieved without the addition of grinding media foreign to the mill feed, differing in this respect from rod, ball, and pebble milling. The term ""semi-autogenous"" has entered common usage in the industry to describe autogenous mills to which supplementary grinding ball additions are made to increase the rate at which size reduction of the mill feed occurs, beyond that achieved solely by the action of rock on rock. In a semi-autogenous grinding (SAG) mill, 50Jo to 10% of the mill volume may be occupied by grinding balls, whereas a range of 25% to 45% is more usual in ball milling.The comminution of mineral ores using ""conventional"" techniques has been described in previous papers. In conventional operations, size reduction of mined ore, and liberation of the valuable mineral content, is typically accomplished using flowsheets employing two to four stages of crushing. Crushing is followed by one to three stages of grinding and classification using various combinations of rod, ball, and pebble mills. This paper describes operations employing autogenous or semiautogenous grinding as the principal stage in the comminution process. Primary crusher discharge, or in some instances run of mine ore, is fed directly to an autogenous or semi-autogenous grinding circuit. Final product sizing may be accomplished in a single grinding stage, the mills operating in closed circuit with classification equipment, or in several stages using autogenous and semi-autogenous mills for primary grinding, and ball or pebble mills for secondary grinding. In several operations a portion of the primary mill discharge is crushed in an intermediate comminution step, these circuits being called autogenousball- crusher circuits (ABC). Many different flowsheets exist, but common and important to most are crushed ore stockpiles, reclaim feeder and mill feed conveyor systems, pumps, screens, and classifiers. Autogenous grinding is employed in both wet and dry processes, although today wet predominate."

Jan 1, 1989

By Y. J. Yu, C. D. Zhang, Y. S. Zhao, W. G. Liang, Y. D. Geng, J. L. Bi

"In this study, the coefficient of thermal expansion (CTE) and thermal conductivity (TC) of shale in two perpendicular directions, vertical and parallel to the bedding of the rock, are firstly measured by experiment. The experimental results show that the CTE of the shale in the vertical direction is 1.69 times as much as that in the parallel direction, while the TC in the parallel direction is 1.83 times larger than that of the vertical direction. Based on the no initial strain assumptions, numerical simulations for the thermal cracking in shale are carried out with the obtained coefficients. Using finite element method and thermal-mechanical coupling theory, two fracture mechanical parameters, KI and J-integral of the shale are also analyzed for the shale thermal fracturing. The simulation results show that both KI and J-integral reach the maximum when the fracturing angle equals to 0° in the shale, namely parallel to the bedding direciton. It is also found that the two fracture parameters decrease with the fracturing angle increasing. Additionally, the thermal fracturing effect of the shale is evaluated under a given scenario. The study is significant for shale oil or gas recovery by in situ pyrolyzing method.INTRODUCTIONOil shale is regarded as a kind of unconventional resource that can contribute to production of shale oil and hydrocarbon gases (Rahm & Dianne, 2011). As is known that there is a large deposit of oil shale in China, and present technology for exploiting oil shale still concentrates on open-pit mining, obtaining shale oil by retorting shale on the ground (Li, Tang, Xue, Zheng, & Du, 2014; Qian, Yin, 2008). However, this technology has a bad effect on the environment and leads to severe resource waste for utilizing oil shale. Exploiting oil shale by in-situ method for ‘green mining’ is widely necessary in China (Kang, Wang, Cao, Liang, Chang, & Liu, 2013; Waters, Dean, Downie, Kerrihard, Austbo, & McPherson, 2009; Zoback, Kohli, Das, & McClure, 2012). One of effective method to exploit shale oil is heating oil shale in-situ underground. For instance, the in-situ conversion process (ICP technique) from Shell Company etc. used a direct heating method, and convectional heating oil shale underground by injecting superheated steam from Mining Technology Institute, Taiyuan University of Technology and constructing horizontal fracture in vertical wellbore from Chevron Company use indirect heating method (Chaudhary, Ehlig, & Wattenbarger, 2011; Hazra, Lee, Economides, & Moridis, 2013; Zhao et al., 2010). The convectional heating method contains complicated procedures with phase transition in double-media of fracture-matrix of oil shale subjected to multi-physical fields coupling including stress, fluid and temperature (Zhao et al., 2010). The characteristics of thermodynamics, solid mechanics and seepage of oil shale have been investigated separately (Hopkins, Rosen, & Hill,1998; Kang, Zhao, & Yang, 2008; Yang, Zhao, Kang, 2010; Zhao, Yang, Feng, Liang, & Kang, 2008; Zhao, Cao, Zhao, Lin, & Wang, 2008). However, it is quite difficult to solve constrained problem of each physical parameter in multi-physical fields coupling for oil shale pyrolysis."

Jan 1, 2015

By Malcolm Black

Introduction The circumstances under which the gold mining industry was forced to curtail operations a few years ago are well known to all. The effect of those circumstances is particularly noticeable when one enters some of the larger mills that were designed, in better times, to treat relatively large tonnages. On their grinding floors one will observe that a number of grinding units are lying idle, and the ravages of rust are everywhere in evidence. The same applies to other equipment throughout these mills. In 1941, we found it necessary to reduce tonnage considerably in the Wright-Hargreaves mill. With the assumption that a lower tonnage would be of long duration, we began to focus our attention on some of the idle equipment with a view to re-arrangement of the flow-sheet for greater economy of operation. In an article by the writer published in the Engineering and Mining Journal (March, April, and May issues, 1939) , it was pointed out that in designing this mill there was a very limited floor space available. The choice of thickening equipment was imposed solely by the necessity for space: conservation. Thickening by gravity would have required 5,000 square feet of floor space, whereas Genter thickeners required only 1,200 square feet. As a consequence, the Genters were chosen. The decision to add flotation to the circuit was made primarily to save space. On a cost-plus-tail basis, 80 to 100 hours agitation on all pulp would produce similar overall results, but would require much more space.

Jan 1, 1948

By L. H. Van Loon, E. G. Crayston, A. F. Heather

"IntroductionTHE PROPERTY of Madsen Red Lake Gold Mines. Limited. comprises fifty-eight claims located between Russet and Faulkenham lakes in Baird and Heyson townships of the Patricia portion of the District of Kenora. The mine is seven miles •by road from the hamlet of Red Lake. This latter centre is .connected to the Trans-Canada Highway at Vermillion Bay, Ontario, by provincial highway No. 105 which has a total length of one hundred and thirteen miles.The Madsen property went into production on August 11th, 1938, with a milling plant rated at 300 tons per twenty-four hours. Changes made in the crushing plant during October, 1938, allowed a •finer feed to be delivered to the mill and increased the milling •capacity to 400 tons per twenty-four hours. During 1948, steps were taken to add an additional 400-ton capacity unit to the milling plant. This latter work was completed by May, 1949, giving the present total milling .capacity of 800 tons per twenty-four hours.It was realized in the early 1940's that, eventually, considerable material suitable for backfilling in the mine would be required. As there were no large deposits of gravel or sand anywhere near the mine, officials of the Company considered it advisable to investigate the feasibility of using mill tailings for mine •backfill. Since Madsen mill tailings contain approximately 80 per •cent of minus-200-mesh material, it was not considered practical to use them for underground backfill without first removing the finer material. as the slimes present in a product of this fineness would without doubt preclude any drainage whatsoever. Table I gives a sizing analysis of Madsen mill tailings and Table II a quantitative chemical analysis of the ore."

Jan 1, 1954