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By J. C. Allan

THE purpose of this paper is to outline the pogress made in the development o the gravity separation method o ore dressing known as Sink-and-Float, which, in the opinion of the writers, has now reached a stage justifying the claim that this new method marks a highly important advance in the technique . The essential feature of the newest development of this process is the extremely accurate separation achieved. This is responsible for higher extraction in the sink-and-float unit on coarse sizes and, in consequence, for substantially higher overall recoveries. Whilst this is the most remarkable feature of the new process, it has further important contributions to make to ore dressing under the following heads: (1) The introduction of the sink-and-float separator as a buffer helps to iron-out irregularities in the mill feed. By supplying a concentrate of constant grade to the flotation section, the process also indirectly improves overall recovery. (2) The addition of a sink-and-float unit to an existing mill results in substantial increase in capacity at low per ton capital cost. (3) Where the plant can be installed underground for the pre-treatment of low-grade ores, there is a saving in hoisting costs, and excellent 'fill' is provided near the point of use. ( 4) Low-grade mixed oxide-sulphide copper ores may be pre-concentrated, thereby producing a flotation feed sufficiently rich to render the oxide residue from such flotation economic for subsequent treatment by leaching or other methods. (5) Application of the process eliminates hard gangue constituents which cause high grinding costs, or gangue constituents which adversely affect flotation. The plant is compact and easily controlled. No highly skilled operators or expensive reagents are required, and the labour and power per ton are low. The separator is extremely flexible, working equally efficiently on any throughput up to its maximum capacity.

Jan 1, 1940

By Sakellariou M., Ferentinou M.

Determination of the nonlinear behavior of multivariate dynamic systems often presents a challenging and demanding task. Therefore, complex geomechanics systems are an ideal area for the application of intelligent methods. Often the physical problem is controlled by nonlinear equations, conveying to the use of neural networks. Hudson, (1992a), aiming at efficient engineering modelling and design, introduced and developed Rock Engineering Systems (RES) approach, through the implementation of generic interaction matrix GIM. This approach has been used for a wide variety of rock engineering and other problems, resulting in many case studies available, which illustrate the application of RES to rock engineering problems. Considering the spectrum of the coding methods for the Generic Interaction Matrix, we suggest an objective coding approach. Among the five proposed coding GIM methods, the coding method according to the slope of Pi and Pj relation is relatively accurate, but it might be difficult or impossible to establish or estimate the exact relation of Pi vs Pj for most of the variables involved within a rock engineering system. Self organising maps, are a special type of unsupervised artificial neural networks, which can learn from complex, multidimensional data, and transform them to visually decipherable clusters. The proposed, coding of the GIM, is performed through the combined use of scatter plots produced through, (SOM), and the coding according Pi vs Pj slope. Naghadehi et al., (2013) constructed a database of worldwide experience, in open-pit mine slope stability, the selection of the principal factors was based on Hudson (1992a) who proposed an ‘atlas’ of factors affecting the stability of rock slopes. Slope classification was performed according to status of stability. The rating of the GIM matrix was completed, coupling scatter matrix produced by SOM training, and Pi vs Pj slope plot. Results refer to the identification of clusters, according to status of stability within the original multivariate data set, and the rating of the principal factors according to the dominance and interaction intensity within the open pit mines RES.

Jan 1, 2015

By V. Labiouse, M. Souley, N. Lafrance, C. Auvray

"Chalk rocks have been used for centuries as a strong building material and for the production of lime. The shallow underground room-and-pillar quarries where the chalk was extracted are now abandoned and endure the effects of time and weathering, increasing their risk of collapse. The risk of instability, involving natural or anthropic underground structures, is an issue for more than 10 000 towns in France.In order to investigate the causes of underground quarry collapse and to aid prevention strategies, we carried out a study of the physico-mechanical behaviour and ageing of chalk in abandoned pillars in two underground quarries, in situ-instrumented by INERIS, in the Parisian Basin, France. Core samples were drilled horizontally from pillars in the Saint-Martin-Le-Noeud dolomitic chalk quarry (Oise, France) and the Estreux glauconitic chalk quarry (Nord, France).Uniaxial tests were performed on dry and saturated cylindrical plug samples extracted perpendicular to the cores at regular distances along their lengths. Physical and mechanical properties were obtained. The dolomitic chalk (Saint-Martin-Le-Noeud) exhibits variation in mechanical and physical properties with horizontal pillar depth. Both density and compressive strength seem to increase outwards from the pillar-core to the pillar-wall. This is the opposite of that would be expected from a purely mechanical point of view. SEM analysis revealed progressive degradation and homogeneity of the grains from the edge to the inside of the pillar. At the pillar wall, the crystal faces are well-defined, almost euhedral, and the grains are of variable size. In contrast, at the inside of the pillar, crystals are finer-grained and are anhedral with highly degraded edges. In the physico-mechanical tests and SEM analyses performed on the Estreux samples, no significant variation was observed between the pillar-walls and pillar-core. However, dissolution marks were observed along the entire length of the half-pillar core."

Jan 1, 2015

By A. Mehinrad, F. Rafia

"The Dez River flows from the Zagros Mountains into the great Khuzestan plain, Southwest of Iran. The 203 m high Dez Dam is a double curvature thin concrete arch dam, with a Pulvino, founded on thick bedded to massive conglomerate of Bakhtiary Formation, mainly horizontal and very less jointed at the dam site. Construction work was completed in 1963, creating a reservoir with a total storage volume of 3’315 million m3. The dam body, with some 350’000 m3 volume and (W/H=1.1/1.0), is the dominant feature of the project.Geological and geotechnical studies at the dam site started in 1957 and continued during the whole period of dam construction. Simultaneously, two extensive programs of laboratory testing were carried out in New York and Milan.Most of the results were recorded in the project design documents and in two very well-known papers of Talobre et. al. (1964) and Dodds (1966), however, no detail results are available for most of tests.In order to estimate geomechanical parameters of the rock mass foundation for numerical modeling of the dam body and its foundation, reassessment of the in-situ and laboratory test results using the recent rock mechanics methods was required.This paper explains how the outcomes of those tests, performed more than 50 years ago, were reevaluated and how the definitions and parameters presented in the project reports were re-interpreted using the recent rock mechanics concepts and definitions. Also, the results of classification of Bakhtiary conglomerate using GSI system and application of Hoek & Brown failure criterion for estimating rock mass parameters have been described.INTRODUCTIONThe Dez River flows from its headwaters in the Zagros Mountains into the great Khuzestan plain. Just before entering the plane, the river passes through a narrow, steep gorge that is about 17 km long. The Dez Dam is located at the entrance to this gorge, some25-km upstream of the city of Dezful and 150-km upstream from the provincial capital Ahwaz (Figure 1)."

Jan 1, 2015

By A. F. Banks

"IntroductionThe typical potash processing plant recovers and/ or upgrades from 10% to 20% of its production through the recrystallization process. The feed to the recrystallization section consists of preconcentrated slurries of fine crystal size that cannot be raised to the required product grade by flotation and fine dusts also of low grade and too fine for compaction feed. It produces high grade white potash crystals that can be marketed as a high grade product or blended into the flotation products to raise their grade. In this chapter, a typical leaching (dissolving) and recrystallization operation will be discussed.The crystallization process makes use of the three phase system chemistry of NaCl, KCl, and H,O. The normal potash process operates on a temperature range of 75°F to 95 °F. A three phase slurry feed to the dissolving circuits will be within that range. The potash (KCl) is dissolved by heating the slurry in the presence of sufficient brine to a temperature of 190°F to 200°F. As the slurry is heated, the potash (KCI) dissolves and the salt (NaCI) precipitates. The potash rich brine is then clarified and the salt and other undissolved solids removed by thickening. Clarified brine becomes the crystallizer feed.The hot brine is cooled in evaporative cooling crystallizers. In cooling, the potash precipitates as a high grade white crystal. Crystallization is done in multiple effects to permit energy recovery and control of the process. In a four effect crystallizer the cooling is done by recycling mother liquor counter currently through the first two effect condensors and using brine pond brine for the cooling of the last two.The crystallizer product is removed from the last effect as a slurry at normal process temperature and is fed either to its own centrifuging and drying circuit or blended with the flotation product for centrifuging and drying."

Jan 1, 1989

By H. Yasuhara, N. Kinoshita, S. Nakashima, K. Kishida

"In this study, flow-through experiments in fractured rocks have been conducted under different stress, temperature, and pH conditions so as to understand the evolution of permeability induced by these conditions. Throughout the experiments, the confining pressures and temperatures are well-controlled. Effluent concentrations are also measured for the ten elements of Si, Al, K, Fe, Ca, Na, Mg, Ti, S, and Cl that are included mainly in the targeted rocks. In the early stage of the experiments, the fracture permeability measured abruptly decreases with time, and subsequently reaches quasi-steady state within a relatively short period. In contrast, the permeability in the intact rocks monotonically increases with time. The permeability increase and decrease are attributed likely to the relative dominance of mineral dissolution in the pore spaces and to that of mineral dissolution/crushing at the propping asperities within the fracture and/or mineral precipitation onto the free surfaces, respectively. The evolving rates and magnitudes of the permeability are likely to be controlled by the stress exerted over the contacting asperities, temperatures, and pH of the permeant. Thus, this complicated behavior should be attributed to the coupled chemically- and mechanically-induced effects, which should be examined in more detail for longer period to contribute to the safe entombment of radioactive wastes. INTRODUCTIONUnderstanding of the flow and transport characteristics of porous/fractured rocks is required for the safe and long-term geological isolation of energy-related by-products (i.e., high level radioactive wastes and anthropogenic CO2). Notably, under relatively high temperature and stress conditions, mechanically- and chemically-mediated processes may alter the pore structures in rocks, resulting in an irreversible evolution of the flow and transport behavior. Augmentation in permeability may result from mechanical dilation due to shearing and mineral dissolution within the pore spaces, while reduction may result from reversible or irreversible mechanical compaction, mineral dissolution at the contacting areas, and clogging of pore spaces by precipitation of secondary minerals."

Jan 1, 2015

By Bert Taylor

"The Mining concessions of Quemont Mining Corporation, Limited, are located at 48°16'N latitude, 79°00'W longitude, in Rouyn township, northwestern Quebec. Six hundred and forty acres in area, they lie directly north of, and contiguous to, the Horne mine of Noranda Mines Limited. The north arm of Osisko Lake covers a large area of the southern part of the property. The orebodies lie below the Jake with one body extending through to the bedrock surface.The property was staked in 1922, and was originally known as the Murray Claims. The group attracted interest from numerous exploration companies because of the proximity of the Noranda orebodies.The United Verde Extension Company, of Arizona, optioned the claims in 1926. They sank a shaft to a depth of 235 feet, at a short distance from the Noranda boundary. From a level established at a depth of 215 feet they completed 3000 feet of exploratory driving from which 6000 feet of diamond drilling was done. Only a small body of siliceous gold ore was found before they dropped their ·option.In 1928, The Mining Corporation of Canada purchased a 90 per cent interest in the property. An operating company, later named Quemont Mining Corporation, was formed. An electrical survey was carried out on a portion of the property, followed by deepening of the shaft to 922 feet. Levels were established at 500 and 900 feet, and 4,800 feet of driving and more than 3000 feet of diamond drilling were carried out from them.Geological knowledge at that time recognized that a major break, the Horne Creek fault, lay not far inside the south boundary. Striking north 80 degrees east, the fault dipped south at 85 degrees. The rock on the north side of the fault was massive rhyolite; to the south, meta-diabase and a little rhyolite. A left-handed movement was indicated by a 600-foot offset on a north striking diabase dyke. The small body of siliceous gold ore, referred to previously, was confined to the 215 level."

Jan 1, 1949

By George Langford

Introduction The manufacture of ductile beryllium is one of the most interesting problems confronting the metallurgist today, and one which becomes more acutely in need of resolution as the uses for beryllium increase. Beryllium belongs to the hexagonal group of metals along with zinc, magnesium, and cadmium. The melting point of the metal is about l,280°C. ± 5°C. Due to its brittleness, the metal's chief use is as an alloying agent in copper and iron. The very low absorption coefficient of beryllium for x-rays makes the metal ideal for x-ray windows, and, similarly, because of its low capture cross-section for neutrons, the metal holds a preferred position in the field of nuclear physics. Modern x-ray diffraction methods were responsible for explaining the mechanism whereby zinc metal, although brittle when cast, could be made ductile at room temperature. The diffraction patterns of zinc and magnesium indicated the importance of the parts played by recrystallization and preferred orientation in this mechanism. It was with this in mind that an Investigation into the recrystallization and preferred orientation of beryllium was undertaken. The investigation, after a series of experiments were conducted involving cold-working and annealing, yielded the recrystallization temperature curve for beryllium. Both the texture of preferred orientation and the effect of recrystallization on this texture when subjected to compressive stresses were also determined. W. A. Alexander, J. K. Swinton and L. M. Pidgeon, in their investigation on The Recrystallization Temperature of Beryllium (C.I.M., Trans., Vol. L, 1947, pp. 657-664), utilized the Brinell ball as a means of cold straining the metal. Sufficient strain was developed to estimate the recrystallization curve for the metal. However, in order to determine the lowest recrystallization curve, as well as to develop a preferred orientation structure, larger deformation than that possible with a Brinell ball was required. The plastic deformation of beryllium can be carried out at an elevated temperature above 700°C. Above 850°C., the metal is quite ductile and can be readily compressed or extruded.

Jan 1, 1948

By L. R. Bottomer, G. M. Leary

"The Copper Canyon property is located in the Galore Creek area of northwestern British Columbia. Porphyry copper-gold mineralization on the property is notable for its high gold content and close spatial relationship with a distinctive suite of high level syenitic and monzonitic intrusions, and pseudoleucite-bearing volcanic country rocks. The style of alteration and mineralization is similar in many respects to the nearby Galore Creek deposits.The current geological resource of 32.4 million tonnes grading 0.75% Cu, 1.17 g/ t Au and 17.1 g/t Ag is insufficient to support a mining operation, but definition of additional reserves combined with improved transportation infrastructure in the region may change this in the future. The climate, topography and proximity to an active glacier indicate that the most likely production scenario is an underground bulk mining operation, with the ore being milled offsite.IntroductionThe property is located at 57°07' North latitude, 131 °21' West longitude, on NTS sheet 104G/ 3W, approximately 160 km northwest of Stewart and 90 km south of Telegraph Creek (Fig. 1). The Galore Creek camp and airstrip are located 6 km to the west. Current access to the property is via helicopter either from Galore Creek, or from the Bronson Creek airstrip 32 km to the south-southeast.The deposit occurs within the Boundary Ranges of the Coast Mountains, on the north side of the east fork of Galore Creek, which flows northerly into the Scud River and thence westerly to the Stikine River.Topography is rugged. In the deposit area, elevations range from 1130 m to 1750 m. Valley glaciers and permanent snow and ice fields are characteristic of the region at these elevations, and at its southern limit, mineralization is truncated by the East Fork Glacier.The property is entirely above the treeline, and except for areas of moraine cover and talus slopes, outcrop is abundant."

Jan 1, 1995

By P .. H. Grimley, N. Champigny

"Fourteen companies (Appendix) representing senior mining organizations in Canada were canvassed in order to assess: (1) the current usage and trends, and (2) the usefulness of computer applications in ore reserve estimation and grade control. The results of the questionnaire suggest that:• computers are unsurprisingly, predominant in the mining industry for calculating reserves and performing grade control;• time and cost savings are realized;• computer-based approaches are applicable to underground situations;• geological information is incorporated to the computerization process; and• computer usage does not hinder the collaboration of geologist and mining engineer.A clear need is identified for case histories comparing computer-calculated ore reserves and actual production data.IntroductionAmong the management and technical staff of some mining companies there is a perception that the application of computer based technology for the estimation of ore reserves and grade control, has as many drawbacks as benefits. Some of the objections raised against the utilization of computers for ore reserve estimation are that:• computer-based reserve models work for large open pit mines but not for underground situations;• computer-generated reserves ignore the geologist's knowledge base; they are strictly a number crunching exercise; and• the use of computers for ore reserves hinders collaboration between the geologist and mining engineer.In order to determine (I) the extent of current computer usage and future trends and (2) whether the objections noted above are borne out by experience, a number of senior Canadian mining companies were canvassed.The respondents were asked to provide:• an assessment of the importance of the computer in the ore reserve estimation and grade control process;• the level of involvement of individuals and staff groups with computer-based ore reserve estimation;• the advantages perceived and realized from the utilization of computer-based ore reserve approaches;• the current and future human resources requirements for the evaluation, implementation, maintenance and development of computer-based ore reserve systems;• the hardware and software resources presently used and planned; and• the developments needed to increase the use and acceptance of computer-based ore reserves."

Jan 1, 1990

By Franc R. Joubin, Donald H. James

"The Property of Rix Athabasca Uranium Mines Limited is in the Beaverlodge area of northwestern Saskatchewan. The mine is 7 miles by road north from the Lake Athabasca port of Bushell which is 250 miles north by aircraft or boat from railhead at Waterways, Alberta.The Rix property is a portion of the former DD 1 concession, obtained in 1949 from the government of Saskatchewan when private prospecting for uranium became permissible. The concession was prospected in detail during the 1950 and 1951 field seasons and a great many radioactive show-ings were discovered. Development and drilling were concentrated in two and later three of these areas, and by late 1952 enough ore had been indicated in the Leonard series and ""Smitty"" zone to justify underground development by adit-drift and shaft respectively.Four hundred feet of drift development on the Leonard No. 1 vein and the sinking of a two-level vertical shaft with related lateral development on the Smitty zone were completed in 1953. In April 1954 the Smitty section was placed in production at a rate averaging 100 tons per day, thereupon becoming the first privately owned post-war uranium producer in Canada. The ore is shipped for processing 12 miles by truck to the mill of Eldorado Mining and Refining Limited.Until recently all of the Company's production was drawn from the Smitty orebody. In April, 1955, ore production commenced from the faulted west extension of the Smitty structure, known as the Rix 62 zone, and from the Leonard No. 1 vein.The Smitty orebody (see Figure 2) is an intensely brecciated lenticular zone bordering both sides of the Smitty fault. The breccia areas are irregularly mineralized with pitchblende, very sparse galena, and pyrite in a gangue of feldspar, chlorite, calcite, and silica, all thoroughly soaked with fine-grained hematite to produce the locally-called ""red alteration"".Tenor of the ore currently being mined is classified information. The average grade estimated, prior to production, was 0.20 per cent (4.0 lbs.) U308 per ton."

Jan 1, 1949

By Halet. R. A.

"The Quesabe Mine, owned by Brunston Mines Limited, is a gold deposit in the south central part of Duprat Township, northwestern Quebec, about 10 miles northwest of Noranda.It is one of the few gold mines in Canada developed since World War II. The occurrence of gold in the Quesabe fault structure has been known for many years, but serious exploration did not begin until 1945. By 1949, 15,000 feet of diamond drilling had been completed and a shaft sunk to a depth of 635 feet, with lateral development on the 450 and 600 foot levels. In 1949 and 1950 a new shaft was sunk to 1,050 feet, new levels were established at 800 and 1,000 feet, and ore was mined and shipped to the Noranda smelter. After November 1950 the ore mined was treated in the company's own mill. The grade of ore was marginal under existing economic conditions, and production was suspended in April 1952. The mine produced about 100,000 tons of ore with an average grade of approximately 0.25 ounces of gold per ton.GeologyThe bedrock formations are all Precambrian, and the general geology of the area has been described by W. G. Robinson (1). The rocks in the vicinity of the mine are acidic and basic volcanics of the Keewatin series (andesite, rhyolite, agglomerates, tuffs) invaded by the Flavrian Lake granite stock, and by dykes and irregular masses of intrusive felsite, quartz porphyry, syenite-porphyry, diorite, and lamprophyre.The mine workings are in andesite and felsite, about 2,500 feet west of the western margin of the Flavrian Lake stock.The principal structural feature of the area, from an economic point of view, is the Quesabe fault. This steeply-dipping fault cuts all the rock formations of the area, and has been traced for about 11 miles along a strike of north 65 degrees east from Lake Arnoux in the northeast corner of Dasserat Township to Duprat Lake near the east boundary of Duprat Township."

Jan 1, 1949

By C. C. Allen, H. H. North

"The Young-Davidson mine is in the southeastern part of Powell township, in the district of Matachewan. It is 32 miles southwest of Kirkland Lake, 48 miles southeast of Timmins, and 25 miles by road Northwest of Elk Lake. The Young-Davidson and its neighbour, the Matachewan Consolidated mine, are the two present producing properties in the district. The Matachewan area is the southwestern extension of the Kirkland Lake belt or 'break'. The Young-Davidson mine since coming into production in 1934 has been operated by Hollinger Consolidated Gold Mines, Limited. ""Desultory prospecting(1) has been carried on in the Matachewan area from time to time since the discovery in 1906 of silver in James township at Elk Lake, and some gold has been found in the southeastern part of Alma township and in the central part of Cairo township. The first real interest in the area, however, was started by Jake Davidson's discovery in 1916 of gold on Davidson creek, on what is now part of the Young-Davidson property, and by Sam Otisse's discovery shortly afterwards on claims a short distance to the east, now part of the Matachewan Consolidated property. In subsequent years both of these properties were investigated by various companies; the Young-Davidson chiefly by the Porcupine Gold fields Development and Finance Company, Limited, and the Matachewan Canadian by the Colorado-Ontario Development Company, Limited, which in 1919 was reorganized under the name of Matachewan Gold Mines, Limited. Work was discontinued as the ore was considered too low in grade to be mined"". However, in 1933 Hollinger Consolidated Gold Mines, Limited, undertook to bring the property into production and this became a reality in 1934.The total production to the end of year 1945 was 319,185.59 oz. gold and 73,541.94 oz. silver from 3,228,845 tons of ore. This represents an average grade of $3.67, including exchange and premiums."

Jan 1, 1954

By H. S. Bostock

"The study of the geology of Mayo district began in 1900 with R. G. McConnell's exploration of Stewart river; and has been carried on intermittently for the Geological Survey of Canada by J. Keele; D. D. Cairnes; W. E. Cockfield; C. H. Stockwell; and the writer. It has revealed a thick group of schists; quartzites; and crystalline limestones; almost certainly Precambrian in age; placed in the Yukon group; and a group of younger intrusive igneous rocks varying in composition from granite to gabbro. These include bodies of many sizes and forms; from stocks 8 miles across to small dykes and sills.Prospecting in the district began about 1897 and has disclosed the presence of gold, silver, lead, tungsten, tin, zinc, antimony, copper, arsenic, manganese, and mercury, both in placer and lode deposits . Of these metals gold, silver, and tungsten have been produced from placer deposits, and silver, lead, and gold from lodes. This paper deals only with some structural relationships of the lode deposits, generally referred to as the silver-lead veins, which have yielded the greater part of the mineral production of the area. From 1938 to 1941 the geological mapping of the district for publication on a scale of 4 miles to 1 inch was carried out by the writer, and the ideas expressed here are based on this work. No comprehensive detailed study of the silver-lead veins has been made, but the mapping has brought to light several important facts. The rocks of the Yukon group, although of a striking sameness throughout their section, can be divided into a number of distinguishable members by means of which the main structures can be traced . The most important and readily distinguishable rock in the group is referred to here as the quartzite member. This is composed of quartzite and schist in about equal proportions. The quartzite is resistant to erosion and forms most of the outcrops of the'member and the peaks of the Gustavus mountains and mount Haldane (Lookout)."

Jan 1, 1954

By G. H. Hope, D. Green

"Recent process advances at Lac Minerals Est Malartic Division are discussed. The Est Malartic mill treats massive sulphide, copper/gold ore from the Bousquet 2 mine, to produce a gravity gold product, a copper flotation concentrate and a gold precipitate. The mill operation has been improved by the application of Knelson concentrators to replace jigs, the application of Outukumpu high grade tank cells and optimizing the cyanide leach to minimize cyanide consumption.IntroductionThe Est Malartic mill is located near the town of Malartic in the Abitibi Region of north western Quebec. The mill has been in production since the mid- J 930's processing mainly clean, free milling gold ore. Active mining from the Est Malartic mine ceased in 1980 and the mill was switched to custom processing of ore from La Mine Doyon located some 45 km to the west. This custom processing arrangement continued until 1987 when the expanded mill at Doyon became available to process all of the Doyon production. In 1987, Lac Minerals commenced development of a second mine at the Bousquet site located 40 km west of Malartic. This mine, Bousquet 2, is the down dip extension of the La Ronde mine developed by Agnico Eagle. During pre-feasibility evaluations, it was determined that modification of the Est Malartic mill to handle this new ore source was economically more attractive than building a new facility on site. Modifications to the mill to process the high copper/gold ore from the Bousquet 2 mine began in late 1989 and the new circuits were commissioned in mid-1990. Modifications included the addition of both gravity and flotation concentration circuits to recover coarse, free gold and the copper. The existing leach circuit continued operating on the copper flotation tailings with some minor modifications to the solution handling systems. An SO,/air cyanide destruction circuit was also added at this time."

Jan 1, 1993

By Robert B. Parsons

"Starting January 1, 1987 employers, employees, and other taxpayers are confronted with a new set of rules dealing with the taxation of travel and housing benefits relating to isolated posts in Canada, and tax deductions relating thereto. The new rules may not be the simplest in the world, but they are reasonably fair and at least offer a hitherto unknown degree of certainty as to how such allowances are to be treated. IntroductionThe first shoe fell on February 26, 1986 when Finance Minister Michael Wilson' s first budget presented the framework for the future tax treatment of travel and housing benefits (so-called ""northern allowances"") provided to residents in remote regions of Canada. The second shoe fell some ten months later on December 22, 1986 when the Minister of Finance filled in the earlier announced framework with detailed rules. We now have a complete picture of the tax treatment of northern allowances. Here is what the picture looks like with effect from January 1, 1987.According to Revenue Canada, the net housing benefit to employees will be equal to the value of the accommodation provided, less any rent they have paid. In major centres such as Yellowknife and Whitehorse, having a developed rental housing market, the benefit amount will be based on the fair market value of the rental accommodation in that location. In other areas, employers will be required to value the housing benefits and, to assist employers, Revenue Canada has provided the following guidelines: Common shelter- $100 per month including the value of subsidized utilities Apartment! duplex-$270 per month excluding the value of subsidized utilities, or $400 per month including the value of subsidized utilities House-$450 per month excluding the value of subsidized utilities, or $650 per month in- c1uding the value of subsidized utilities.The amounts established as the values of accommodation will be adjusted annually after 1987 based on the housing component of the Consumer Price Index."

Jan 1, 1987

By H. W. Smith, D. Guerin

"This paper describes the development of a dynamic model for a full- scale (64 tonnes/hr) grinding circuit consisting of an open-circuit rod mill feeding a ball mill operating in closed circuit. The object of constructing such a model is to investigate the static and dynamic behaviour of potential control schemes. Model selection, data processing and parameter estimation are discussed, and static and dynamic results of simulations are presented. Conclusions are drawn with regard to the adequacy of the model for such studies, and likely limitations on control of such circuits.IntroductionThe model developed here is intended to represent the grinding behaviour of the grinding circuit at the Lake Dufault division of Falconbridge Copper Mines Ltd. This circuit has been fully described elsewhere. At the time of the work described, the ball mill discharge was fed to a four-cell flotation bank , the tailings of which are combined with the rod mill discharge in the cyclone feed pump sump. As the emphasis in this study is on grinding behaviour, and as only I per cent of cyclone feed material is removed in these cells, no attempt has been made to include flotation behaviour in the simulation. One effect of the presence of flotation cells in the grinding circuit is, however, to necessitate somewhat finer grinding than might otherwise be needed: this will have effects on mill parameters.The paper is organized as follows. First the selection of models is discussed, paying particular attention to the problem of selecting simple, yet adequate, flow models for the mills. Then data gathering is described. The method of deriving mill model parameters is explained, and the results presented. Mass balancing of cyclone data and the determination of cyclone model parameters are described. The linking of the models into a full simulation is outlined, and sta tic and dynamic results presented. Finally, the results are discussed and conclusions drawn ."

Jan 1, 1980

By D. A. Stokes

"In October of 1980, Fording's 15 000-ton capacity clean coal silo was put into operation. The silo was unique in that it was designed to allow the mass flow of material, coupled with a sophisticated computer-controlled dynamic loadout/weighing system.On May 1, 1982, an explosion, attributed to a methane buildup, devastated the top of the silo. .As a contingent plan to minimize the effect of the production loss the conveyance to the silo top was reconstructed and the silo' was back in operation by mid-June 1982, although the silo top was left open to prevent further methane buildup. As dust emissions occurred during loading ofthe silo, an extensive study was undertaken to establish dust control for the silo, while at the same time providing methane detection and control.The study, conducted for Fording by the Cambrian Engineering Group of Calgary, involved an extensive search and investigation of existing coal silos. The study concluded that there IS no universally accepted method of controlling methane emissions in coal silos.The study also produced a unique conceptual design for handling methane, while at the same time, protecting the environment . Fording's Clean Coal SiloThe silo, located at Fording Coal's Fording River metallurgical coal operation, is utilized for the storage and load out of cleaned thermally-dried metallurgical coal.The silo was constructed in 1979 and 1980 to reduce unit train loadout times and to facilitate the handling of “special shipments"" of coal. The silo went into operation in October 1980, under manual control and utilized computerized, automated loading by February 198I.The silo is 21 m in diameter and 76 m high, with a nominal capacity of IS 000 live tonnes. The coal storage portion of the silo is 21 m in diameter and 48 m high, utilizing 2 steep-angled hopper outlets allowing mass flow of the stored coal during train loading. Each hopper has a control gate and a discharge chute, one chute for bulk loading, the other for trim loading of coal into the rail cars."

Jan 1, 1986

By Jorgen Elbrond, Yve Lizotte

"A mine evaluation method based on the use of a computer program that models mining with an open-ended dynamic program is presented in this paper. It is demonstrated with an example how the method is applied, how some of the decision variables are related in the particular industrial environment and how these decision variables affect the economic benefits of mining.IntroductionTwo distinguishing characteristics make mining a particular industrial environment: exhaustible resources of a given ore deposit and heterogeneousness of the ore grade spatial distribution in the ore deposit. These characteristics, inflation, uncertainty and cyclical price variations which greatly affect estimated cash flows also contribute in rendering difficult the evaluation of the economic outcome of mining at the preliminary planning stage.Decision variables and their ranges are associated with the mine plan; basically, they include choices of the mine's nominal capacity (mining and milling), the sequencing of mining through the deposit, the ranges of possible production rates and cutoff grades, and the stockpiling policies. The mine's nominal capacity must be decided at the planning stage, but this capacity choice must take into account flexibility to modify other decision variables as mining commences and uncertainty resolves.Although it is generally accepted in discounted cash flow analysis that the sequence of mining must proceed from the richest to the lowest-grade stopes (with correspondingly ajusted cutoff grades), this may not produce optimum benefits when cyclical price variations are considered.Mining rates and cutoff grades can also be adjusted to changing economic conditions, taking into account the limited resources of the ore deposit. Thus, when prices are thought to be in the trough of a cycle, production can decrease to preserve ore for expected greater future profits. When prices rise, mining rates can increase possibly even over the mine's nominal capacity. In both cases, the machinery's effectiveness shall not be-optimally utilized; this will be reflected particularly by the variations in the concentrator's metal recovery and the concentrate grade output."

Jan 1, 1982

By J. Weyer

"Threshold limit values (TLVs) for NO and NO2 have been reduced drastically in 2016. Emissions of diesel engines have been reduced in the last decades. In many cases this is not enough to stay under required TLVs. Remaining gas in the blasted rock is set free during loading and transport. At the end, only a combination of measures with a further reduction of diesel motor emission, replacement of diesel driven vehicles by electric equipment, new explosives with reduced NO and NO2 production, changes in the organization and effective ventilation measures can ensure that the TLVs are not exceeded. KEYWORDS Blasting fumes, Gas measurement, Gases, Threshold Limit Value (TLV) SHORT HISTORY OF EXTRACTION OF RAW MATERIALS The extraction of raw materials—besides haulage and transport—is one of the main processes in mining and tunneling. The principal objective is the creation of a material which can be transported from the place of extraction (face) to another place for further processing or direct use. Furthermore, recently a desired grain size distribution has become more important. A revolutionary development took place in the course of centuries. With the use of primitive tools like shovels, chisel and iron in the 12th/13th century in Freiberg, ore mines performances of 12 16 cm per man and week could be reached in cross section of 1 m2. Recalculated with an assumed average density, this is a performance of 0.006 t/h. With the introduction of black powder into the subsurface in 1627 by Caspar Weindl in the city of Schemnitz (nowadays Banska Stiavniza), performances increased. A part of the hard extraction work was now done by the explosives. But first in 1767, it was recognized in a mine in Freiberg that more than 1 hole could be drilled (hand-drilling) and blasted. This discovery was so important that a plate with an inscription describes this event. The performance was increased more than 10 times compared to the extraction with chisel and iron. Since this time, drilling and blasting has been the most important innovation force for the increase of the performance. Problems with noxious blasting fumes led to the construction of the first tunneling machine by Henry Joseph Mouse in 1846 (beside shield machines from 1825) (Stack, 1982)). The triumph of mining machines started only interrupted by the two world wars. The development of hard rock boring machines (including roadheaders) was enforced after 1945. An almost complete extraction by hard rock boring machines was predicted for the turn of the millennium. And, indeed, one could find powerful machines in Canadian potash mines which were able to mine 1000 t of potash in 1 hour. That was an increase in the performance by several powers of ten. But already in the 1980s, it was recognized that the ambiguous goal could not be reached. Problems with wear of the tools, radius of curves, minimal length of openings, size and costs of the machines led to a more or less parity between drilling and blasting on one side and machines on the other side, of course depending on the field of use. For this reason, drilling and blasting is still the dominant extraction technology for raw materials. In the development of drilling and blasting technologies, the focus was mainly on the improvement of the safety but also on the increase of the blasting power and performance. Examples are tests for the increase of the power of ANFO explosives which were carried out in the 1970s and 1980s, in the former teaching mine “Reiche Zeche”. In the 1980s vehicles for loading and mixing of ANFO explosives have been developed, the blasting cab sensitivity of ANFO has been increased, emulsion explosives have been developed and up to now further developments like PMMA LOX EX explosives are on the way. At the turn of the millennium, 6 million t/a explosives have been used worldwide. Annual rates of increase of 2% have been predicted (Weyer, 1999). Meanwhile the focus changed from an increase of the performance"

Jan 1, 2019