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By Denni Nikols, Farrukh Z. Khan

"Highvale Mine, located at Wabamun, Alberta, produces thermal coal for TransAIta's Keephills and Sundance generating plants. Overburden stripping in the mine is conducted by draglines which normally operate on top of the overburden bench. In certain areas oj the mine, however, these overburden benches are susceptible to slope failure. In these areas, due to safety considerations, the draglines are utilized in the less productive chop-cutting mode, supplemented by mobile equipment where the bench height exceeds the dragline chop-cutting limit. This paper proposes maintaining the draglines in their normal digging mode through areas of unstable or jailed overburden, by utilizing a dragline carried toe buttress 10 ensure adequate bench stability. The data requirements and methods of slope analysis to design a toe buttress across an unstable highwall area are examined in rigorous detail. INTRODUCTIONLocationHighvale Mine is owned by TransAlta Utilities Corporation and operated under contract. It is Canada's largest coal mine with an annual coal production in excess of II million tonnes. The mine is located approximately 80 km west of Edmonton on the south shore of Wabamum Lake (Fig. I) and production is currently provided by three pits 02. 03 and 04 (Fig. 2). The entire coal production is consumed at the nearby Sundance and Keephills power generating plants.ScopeThe mining operation at Highvale Mine has had occasional slope stability failures of its high wall overburden bench. This paper examines one such failure which occurred in the summer of 1983 at Pit 03. In addition to an investigation of the failure, Terracon was asked to recommend mine operation procedures which would minimize the impact of these slide zones on dragline productivity. A method of removing the failed material by utilizing the dragline in its normal digging mode is designed. The method's advantages and disadvantages as well as the required support procedures are reviewed."

Jan 1, 1989

By R. L. Armstrong, C. I. Godwin, V. M. Kruan

"The Scheelite Dome tungsten-bearing skarn property, central Yukon Territory, is in metasediments in contact with a quartz monzonite intrusion. Contact metamorphism to biotite-muscovite- andalusite hornfels and calc-silicate hornfels is superimposed on low-grade regionally metamorphosed rocks represented by quartz-mica pelitic schist and muscovite-biotite pelitic schist.Metasediments are likely Precambrian in age, even though Rb-Sr whole-rock analyses indicate an early Paleozoic age (511±41 Ma); this apparently too young radiometric age is due to partial loss of radiogenic strontium during regional metamorphism. The minimum age of the quartz monzonite intrusion is Cretaceous at 90.4±2.6 Ma (K-Ar on biotite) and its initial 87Sr/86Sr ratio is 0.7121. The high initial ratio of the intrusion is due to the addition of radiogenic strontium from the metasedimentary country rocks and perhaps an underlying older basement.Scheelite-pyrrhotite-chalcopyrite mineralization occurs in diopside-zoisite skarn pods within the calc-silicate hornfels near the quartz monzonite intrusion. The source for lithophile tungsten is probably the same as for the granitic magma, namely Precambrian sialic basement.IntroductionThe Scheelite Dome tungsten skarn property is within the Mc-Questen map-area (NTS: 115P/16),28 km northwest of Mayo, Y.T., at latitude 63°47' north and longitude 136°15' west. The property can be reached by helicopter or by four-wheel- drive vehicle from Mayo.This paper summarizes the geology of the Scheelite Dome property, which has been described in a general way by Little (1959) and Cathro (1969). Geological mapping and sampling were done by the first author (nee Steffler) during June 1979. Field work was supplemented with detailed work at the Department of Geological Sciences of The University of British Columbia during 1979 and 1980 (Steffler, 1980). Whole-rock chemical analyses were done by Cominco Ltd."

Jan 1, 1982

By D. E. M. Plaumann

"The wet grinding of asbestos has attracted the attention of engineers for many years as a means of improving fibre liberation from the ore. Today, added impetus for wet grinding has resulted from a need to minimize worker exposure to asbestos dust in conventional dry mills using air, as well as a need to provide a cleaner, technically superior fibre for use by product manufacturers. However, high viscosities of asbestos slurries create transport and grinding difficulties which have been a factor impeding use of a wet grinding process.Laboratory evaluation of reagents to reduce viscosities of asbestos slurries found that one reagent, XFS-4272.00, reduced the viscosity of a 35% (by weight) slurry by 75%. The impact of XFS-4272.00 on actual grinding parameters were evaluated using a pilot plant rod mill. Slurries containing 78% solids and XFS-4272.00 were successfully milled and asbestos fibres from that milling compared favourably in liberation and relative surface area to asbestos fibres obtained from milling a 4% slurry without additives. IntroductionAlthough the use of asbestos has been declining over the last few years, the large deposits of chrysotile in eastern Canada (Quebec in particular) remain very important. Asbestos production in Canada stands ninth in volume for mineral production with over 740 000 tonnes produced in 1985. Of this quantity, 722000 tonnes were exported with a value of $446 million. Canada is second in world production of asbestos behind the U.S.S.R. The milling of asbestos ore has been done primarily by a series of steps involving crushing, dry-grinding and aspiration using large quantities of air to release the fibres. Contrast this process to those practiced in the purification of virtually every other mineral. Ores other than asbestos are crushed and milled using an aqueous slurry in rod, ball or pebble mills. The finely ground mineral, still as a slurry, undergoes beneficiation by froth flotation, heavy media separation, hydrocyclones or any of the other aqueous techniques currently employed in mineral processing.One of the driving forces behind the use of air instead of water as a means of separating asbestos is the nature of the aqueous asbestos slurry. While most slurries of minerals contain as much as 80070 solids and can still be pumped and transported, asbestos slurries become highly viscous with as little as 10070 solids. These high viscosities reduce the pumpability of the slurry and create severe handling problems. Mills become prohibitively large and uneconomical due to having to handle slurries containing only 10% rock and 90% water.Health-related issues have affected the use of asbestos in many countries. In Canada, the asbestos-producing provinces have introduced stricter occupational legislation to improve safety in the workplace. This primarily means strict control minimizing release of air-borne asbestos fibres . Quebec and Ontario have set limitations of airborne asbestos at 2 (2) and 1(3) fibre per cubic centimetre of air. With present milling technology, maintaining these low levels in the workplace requires constant vigilance."

Jan 1, 1988

By P. R. G. Brandão, E. E. Ferreira

The presence of metallic cations in solution will affect the aggregation or dispersion of the fine hematite particles, depending on the type of cation adsorption: either specific or non specific adsorption. In this work, the zeta potential measurements were carried out on a sample of pure hematite as a function of pH and metallic cations concentration (Ca2, Mg2+, Mn2+, K+, Na+ and Al3+). The objective of these studies was to clarify the effect of these metallic cations on the surface chemical properties of pure hematite. The cations K+ and Na+ adsorbed as non-specific species and caused only the reduction on the values of zeta potential (in terms of absolute value) over the entire range of pH values tested. For the Al3+ cation, in the acidic pH range it adsorbed as a specific species and increased the zeta potential values. The cations Ca2+, Mg2+ and Mn2+, adsorbed as a non specific species at pH levels less than 6.0 and caused the reduction of zeta potential due to the compression of the electrical double layer. However, for pH values greater than 6.0, these cations adsorbed as a specific species causing the zeta potential to increase and became more positive as the pH value increased (alkaline). This effect could be due to the adsorption of some hidroxy complex species of these cations, which contributed to the positive charge. The effect of metallic cations (Ca2+, Mg2+ and Al3+ concentration, at pH 4.0, 7.0 and 10, showed the following results: a) at pH 4.0, increasing the concentrations of Ca2+ and Mg2+ resulted in on a small reduction in the value of the zeta potential. At this pH, increasing the concentration of A131 cations caused the zeta potential to increase due to specific adsorption of a negatively charged Al3+ hydroxy complex b) at pH 7 and 10, increasing the concentrations of Ca2+ and Mg2+ caused the zeta potential to change its charge from negative to positive.

Jan 1, 2004

By R. Ricketts, D. G. Vanderveer, F. Kirby

"A program of mapping, sampling and testing to determine the potential of the aggregate resources throughout Newfoundland and Labrador was commenced in 1978. A 6-km-wide corridor parallel to all transportation routes was selected as the study area.The study consisted of a preliminary surficial interpretation followed by field mapping and sampling. Granular resources were sampled at a 500-m interval and non-granular materials were sampled at a 1.5- to 2-km interval. Data were gathered on landforms and stratigraphy as well as a field description of the sample and deposit types. Particle-size and lithologic analyses of the samples were also conducted.The locations of samples and deposits were mapped on 1:50,000 topographic base maps and summarized for publication at a 1:250,000 scale. Surficial geology maps were produced at a 1:50,000 scale. A computer program is being developed for the search and retrieval of field data.The first phase of the Inventory of Aggregate Resources program is nearing completion and to date has collected information on approximately 9,500 samples throughout Newfoundland and Labrador. This data will be used to provide f or the effective management of these vital non-renewable aggregate resources.The mineral aggregate industry in Newfoundland and Labrador is very competitive, but undercapitalized when compared to other provinces. The demand for aggregates is forecasted to increase to in excess of23 million tons per annum by the year 2002.In terms of over-all mineral production, the extraction of sand, gravel and stone has a significant impact when compared to other minerals in terms of value and tonnage produced.IntroductionThe province of Newfoundland has approximately 18,000 km of coastline. A large proportion of the population lives in coastal communities where, traditionally, beaches have been a major source of construction materials. Beaches have, therefore, often been subjected to overutilization, resulting in erosion of back beach areas or adjacent headlands."

Jan 1, 1982

By L. Xu, A. Vazirnegad, R. Bhaskar

"Respirable dust samples are collected by the U.S. Mine Safety and Health Administration (MSHA) and mine operators as part of the requirements of the Federal Coal Mine Health and Safety Act of 1969. About 1.85 million respirable dust samples collected by MSHA were analyzed for the period 197/ to 1990 f or the states of Alabama, Colorado, Illinois, Kentucky, Ohio, Pennsylvania, Utah, Virginia, and West Virginia. Trends in coal dust exposure of continuous miner operators, helpers, and roof bolters were analyzed. The reported production in each of the nine states was analyzed. The relationship between average dust exposure and coal production was evaluated. In addition, the differences between mine operator and MSHA inspector collected data were examined for each state. The average exposure levels for the three occupations - miner operator, miner helper, and roof bolter were also compared.The data indicate that the dust exposure levels have significantly dropped over the last twenty years, while at the same time production has increased, attesting to the ability of the U.S. mining industry to adapt successfully to the stringent dust control regulations. In terms of coal production, it appears that some states, such as Utah, Illinois, and Kentucky, were able to increase their production significantly relative to others, such as Pennsylvania and Virginia. The relationship between dust concentration and coal production is, at best, tenuous. 11 implies that existing dust control technology has assisted several mine operators to meet dust standards, while, at the same time, enabling mines to produce more coal than before. Comparison of data collected by MSHA and mine operators indicate that MSHA inspectors recorded higher dust levels and lower shift production relative to mine operators."

Jan 1, 1993

By Gordon E. Agar

"The effectiveness of two this compounds for copper sulphide depression in molybdenite flotation has been examined with batch bench scale tests.Both thioglycollate and trithiocarbonate proved to be good depressants producing results similar to those f or cyanide and hydrosulphide. Ferrocyanide was ineffective on the samplematerial tested.The depressing effect of thioglycollate was readily reversed by an addition of xanthate.IntroductionThe separation of molybdenite from copper sulphide minerals has been practised for a long time(1,2). The predominant scheme is to depress the copper sulphides while capitalizing on the ready floatability of molybdenite to effect the separation. There are many recipes for depressing the copper minerals(3), among them the addition of sodium cyanide, particularly if the copper is present primarily as chalcopyrite; some form of sulphide whether as sodium sulphide, sodium hydrosulphide or phosphorous pentasulphide when there is a mixture of copper sulphides such as bornite, chalcocite and chalcopyrite; potassium ferro cyanide may be used when chalcocite is the principal copper sulphide; peroxide or hypochlorite; as well as heating in the presence of lime. There are different mechanisms by which these reagent s depress the copper sulphides. Sodium cyanide and the sulphide-producing compounds apparently reduce the potential of the copper sulphides to the point where they ar e rendered unfloatable by xanthates . Another study(4) showed that after cyanide addition, chalcopyrite was rendered un floatabl e until the REDOX potential (measured with a gold electrode with a saturated calomel reference) had become more positive than -250 mY. If the chalcopyrite pulp was conditioned after the cyanide addition the REDOX potential increased slowly, whereas if air was introduced during the conditioning the potential r ose much more rapidly. The difference in time to reach the desired level of - 250 mV was as much as 20 minutes' 25 minutes being required when air was absent but less than 5 minutes when air was admitted . These observation s demonstrate some of the problems encountered when using cyanide to depress sulphides such as chalcopyrite . The system adjusts to counteract the depressing effect of the additive. To overcome this tendency it is necessary to either add an overdose at the beginning or to more or less continuously add the regulating reagent. In this study cyanide solution was added under automatic control to maintain a specified REDOX potential."

Jan 1, 1984

By G. M. Humphrey

It is a great honour and privilege to be here in this distinguished company of the outstanding men in the mining industry and the political life of Canada. I fully recognize that I am so honoured as a symbol of the combination of American and Canadian leaders in industry who are partners in the sponsor-ship of the Quebec-Labrador iron-ore mining development which we are undertaking, and as such I want to express our great appreciation of the co-operation and assistance we have met on every hand in Canada, and the genuine friendliness and never-failing courtesy which we are always accorded. We are quite aware that it springs primarily from the very high regard in which our partner and sponsor, Jules Timmins, and our other Canadian associates are held. Their high standing among you is well known and we quite naturally fall heir to the benefits of the great esteem in which you hold them. Jules Timmins and his long-time associate, John Knox, were the original inspiration for our participation in this venture. It was their enthusiasm and vision which caught our interest and led to our joining with them in the years of exploration and search which have brought us to the threshold of this great development where we stand today. So much has been written of the details of our exploration and plans for future development that it would bore you for me to review them further here, but there is a phase of this development, and an extremely important phase of it, which I think has passed almost unnoticed and which it may be well worth while for me to bring to your attention this evening.

Jan 1, 1951

By The Staff

Introduction In this paper it is proposed to discuss briefly some of the problems confronting the management of a gold mine in choosing a method of milling to recover the gold from various classes of ore. In order to simplify the discussion, the more common types of gold ores are classified from a metallurgical point of view as follows: Class 1?A: Free-milling ores which require only relatively coarse crushing. B: Free-milling ores which require finer grinding to release the gold. Class 2?A: Gold ores which contain the gold partly in the flee state and partly associated with a small amount of iron sulphide, and which do not require fine grinding. B: Gold ores similar to 2A, but which require fine grinding. Class 3?A: Gold ores containing copper in quantities insufficient to interfere to a serious extent with cyanidation. Gold ores containing copper in sufficient quantity to interfere with cyanidation but in which a large proportion of the gold is in the free state, the remainder being chiefly associated with the copper mineral, chalcopyrite. Gold ores containing copper and other sulphides, such as pyrite, etc., in which the gold is principally associated with these sulphides.

Jan 1, 1931

By G. S. Dobby, J. A. Finch

"A significant development in froth flotation technology has been the emerging industrial acceptance of column flotation during the past few years. Flotation columns are being employed now in several Canadian base metal concentrators and they are being considered f or a wide range of flotation applications throughout North and South America. Since the operating principles of a column differ markedly from those of conventional flotation machines, and have been poorly understood, scale-up of columns has been difficult. This paper describes a methodology f or translating laboratory column data to industrial operation.The scale-up model utilizes kin etic data that is obtained from a series of laboratory column experiments; the laboratory column can be as short as 2 m. For modelling purposes the column is considered to consist of two zones: the collection zone, where particle recovery occurs, and the cleaning zone, a packed bubble bed generated by downward flowing wash water. Mixing conditions in the collection zone have been characterized using the results of tracer experiments on plant columns at Mines Gaspe. The same experiments demonstrated the nearly complete suppression of entrained gangue minerals that is attained in the cleaning zone. The model explicitly accounts for both the collection zone and cleaning zone recoveries and allows for the effects of bubble overloading. Results of column scale-up experiments at Gibraltar Mines are presented."

Jan 1, 1986

By E. Yalçin, A. Ünal

In transitional mineralization, the sill of the underlying variogram is relatively easily estimated, but the estimation of the nugget effect may be more difficult. This study aims to determine for massive copper deposits just how sensitive the estimations of ore reserve, economic copper content, total net value, and ultimate pit limits are to the errors of estimation in the nugget effect of the spherical variogram model. A massive copper deposit which has a log-normal distribution and a spherical variogram was simulated on a rectangular grid using the turning bands method. By using the three dimensional random kriging method, the mean block grades were estimated at four different sampling densities. The ultimate pit-limits were determined by using the two-dimensional dynamic programming algorithm. To calculate the estimation errors, the values obtained by using the actual nugget effect and structured component values were deducted from the estimations obtained using the incorrect nugget effect and structured component values. For isotropic massive copper deposits, it was found that the ratio of the nugget effect to the structured component of the spherical variogram has no significant effect on either the block grades estimated by the kriging method or on the ultimate pit limits.

Jan 1, 1992

By A. McLean, Q. R. Yang, Y. D. Yang, G. F. Zhang, P. Wu

A considerable amount of waste slag from the pyrite roasting process is discharged every year. Recovery of iron from the waste slag is of great interest for comprehensive use of mineral resource, reduction of environment impact and improvement of economical benefit to related companies. The objectives of the present work are focused on: (1) Compare various procedures for recovery of iron from the waste slag and select a best procedure for industrial trial which is roasting of the waste slag under reducing condition followed by magnetic separation; (2) Investigate the optimum conditions for maximum recovery of iron from the waste slag. The experimental result shows that under the treatment condition found in this work about 36 to 40% of the iron oxide was recovered from the waste slag, where the iron content in the concentrate is 57%. The total recovery ratio of iron, up to 70%, from the waste slag was achieved. After the treatment, sulphur content in the waste slag was reduced significantly and gangue materials, such as SiO2, Al2O3 and TiO2, were reduced effectively.

Jan 1, 2011

By P. Chu, J. A. Finch, R. Li

"Bubble size plays a key role in flotation. In this paper, we examine the possibility of using passive acoustic sending to determine bubble size. Software was developed based on Short Time Fourier Transformation (STFT) to analyze the frequency of the acoustic signal emitted from the bubble formation event. Bubble size was determined from the Minnaert frequency. The acoustic-determined bubble size responded to reagent dosage, but was less sensitive than that given by imaging. The constraints of the current acoustic approach are discussed.INTRODUCTIONBubble size is an important factor affecting flotation performance. It determines the maximum available surface area for collecting and transporting mineral particles. Of interest for process control (Gorain, Franzidis, & Manlapig, 1995), it can be determined by several techniques, such as image analysis (Grau & Heiskanen, 2002; Hernandez-Aguilar, Gomez, & Finch, 2002); conductivity (Barigou & Greaves, 1992); drift flux analysis using gas rate and gas holdup measurements (Banisi & Finch, 1994); and optical sensors (Randall, Goodall, Fairlamb, Dold, & Connor, 1989). However, all these techniques have drawbacks.Acoustic analysis of the oscillatory behaviour of bubbles to extract bubble size is another possibility (Leighton & Walton, 1987; Pandit, Varley, Thorpe, & Davidson, 1992; Al-Masry, Ali, & Aqeel, 2005; Manasseh & Ooi, 2009; Leighton & White, 2012; Spencer et al., 2012). The fundamental basis is the work of Minnaert (1933) who first correlated the bubble size with its natural frequency, Equation 1:"

Jan 1, 2016

By Sadan Kelebek, Gerhard W. Heinrich, Peter F. Wells, Peter J. Whittaker

"Based on plant mineralogical data & circuit modelling, a pilot plant campaign was conducted at Strathcona Mill to identify options for improving the performance and operating flexibility of the copper-nickel separation circuits. Emphasis was given to options for combined treatment of the two feed streams (primary nickel ore rougher concentrate and copper ore circuit concentrate) to obtain more flexibility in the use of existing column flotation capacities. The pilot test results demonstrated that circuits which incorporated preconcentration prior to column flotation gave the best metallurgical results. Maximum column carrying rates were also related to the preconcentration. Predictions that increased circulating loads from the column scavengers would improve copper-nickel separation performance were confirmed.IntroductionThe copper-nickel separation circuits at Falconbridge's Strathcona Mill produce a high grade (>30% Cu) copper concentrate most of which is treated at the Falconbridge Kidd Creek smelter near Timmins. The remaining nickel concentrate which still contains approximately 20% of the feed copper is combined with other lower grade nickel/copper concentrates and shipped to the Falconbridge nickel smelter. The separation is required because of the limited capacity for copper processing in and downstream from the nickel smelter. Since no nickel is recovered at present in the Kidd smelter, the challenge is to minimize the nickel content of the copper concentrate.At present there are two independent separation circuits treating the rougher concentrates from the copper and nickel circuits as shown in Figure I. The minerals treated are mainly chalcopyrite and pentlandite but there is up to 25% of the copper present as cubanite in the copper ore. The circuits were kept separate because the flotation rates and froth characteristics were very different, probably because of the presence of cubanite and the large difference in the concentrate grades feeding the circuits."

Jan 1, 1995

By D. L. Mckay

"Texasgulf's 25-foot-diameter Kidd Creek No. 2 Shaft is situated next to the company's No .1 Shaft, so that both shafts can share common surface and underground facilities. Shaft construction, covering a span of five years from collaring to commissioning , started by freezing the 50 feet of grey varved clay covering the bedrock. After first completing the concrete headframe, sinking to the 5/00-foot horizon utilized two different shaft construction methods. The top 2800 feet was excavated by slashing a 6-foot-diameter borehole and hoisting the broken material up the 3OO0-foot-deep No.1 Shaft; the remaining 2300 feet was excavated by conventional benching methods. Prior to installing the 5100 tons of galvanized shaft steel on the completion of the sinking operation, the sinker was able to install all thirteen armoured electrical cables in one pass with his modified galloway stage. All loading and lip pockets were installed along with the steel guides and bun tons during the equipping operation.IntroductionOn November 21,1978, Texasgulf completed the No.2 Shaft at Kidd Creek Mine. This shaft is the largest shaft (volumewise) in North America and is unique in many respects . The purpose of this paper is to outline some of the unique features of this shaft as well as to record how this facility was constructed.Soon after sinking No.1 Shaft, Texasgulf started a program of exploratory diamond drilling from the 2800 level in No. 1 Mine. The drilling outlined sufficient ore below that level to justify a second mine. Feasibility studies demonstrated that production could be increased by one-third exclusively from this deep ore . The study called for a 25-foot-diameter shaft to the 5100-foot horizon to service a mine producing 5,000 tons of ore per working day.The site chosen for No.2 Shaft and head frame is 350 feet southwest of No.1 Shaft. This, along with No.1 Shaft, will be away from the influence of ground subsidence when mining at depth. The collar and head frame foundations are located on a flat rock ledge some 70 feet below surface. This flat ledge was found by test holing through the overburden by diamond drilling. It is fortunate that this location best satisfied the surface layout and logistics for both surface and underground ore and waste handling."

Jan 1, 1981

By T. L. Carpenter

"The presence of excessive water in the Gibraltar East Stage 1 pit initiated studies into the ground-water conditions that might be expected in mining the Granite Lake and Pollyanna pits. Hydrologic information obtained from a pump test program in 1974 indicated that deep-well dewatering could be successfully achieved. In 1976, a deep-well dewatering system consisting of eight wells was installed on the periphery of the Pollyanna pit . it has been responsible for a significant decrease in mining costs.This paper presents a brief summary of the engineering principles involved in the design of the Gibraltar dewatering system and the problems encountered in the installation and operation of deep-well pumps. IntroductionGibraltar Mines is an open-pit copper-molybdenite mine located approximately 230 air miles north of Vancouver in the central interior of British Columbia (Fig. I). The property consists of three separate major orebodies; Gibraltar East, Granite Lake and Pollyanna (Fig. 2). Mining commenced during the summer of 1971, in the Gibraltar East Stage 1 pit. The Granite Lake Stage 1 pit was mined from mid 1974 to mid 1976. The Pollyanna Stage 1 pit is currently being mined and the Gibraltar East Stage 2 is being developed for production. This report is intended as an outline of the studies and tests leading up to and including the installation of the Pollyanna pit dewatering system, the problems encountered and the results obtained.GeneralIn the summer of 1973, mining in the Gibraltar East pit progressed below ""daylight"" on the down-hill side and ground water very quickly became a costly nuisance in the day-to-day mining operations. Mining, maintenance and blasting costs reflected these problems. Concern was also voiced about adverse effects of ground water on pit wall stability. This concern later became a reality in the Gibraltar East and Granite Lake pits. A study was initiated to investigate various methods and costs of dewatering the pit areas. The study included in-pit sumps, deep peripheral wells and a drainage gallery system located below the pits."

Jan 1, 1980

By Balmer Neilly

FOR some reason or other, Mine Taxation has never been a popular subject for discussion in meetings of this kind, and it has never received the study and consideration its importance demanded. By way of explanation, two reasons may be suggested. In the first place, the computation, adjustment, and final payment of taxes, have been looked upon in most cases as a Head Office chore. Secondly, since taxes represent an uncontrollable cost item, they have been treated as an added rather than an inclusive factor in calculating unit costs. Perhaps because I am half Irish, and therefore resent taxation of all kinds, perhaps because I am half Scotch, and am therefore concerned with the results, I have always taken a keen interest in this subject. Moreover, by some peculiar set of circumstances, I have been a participating party in many, perhaps I can say most, of the discussions on this subject that have taken place between the Governments and the representatives of this industry during the past thirty years. Looking backward, certain very definite impressions remain. Political considerations have never had any part in these discussions. Views have been exchanged upon the frankest possible basis, and, when we disagreed on some particular detail, we agreed to disagree, and followed through in harmony and? good faith. We, as an industry, have refrained from importuning and fault finding in ways that might be considered as petty or selfish, and it is only upon occasions such as this, where conditions necessitate action, that we lay our troubles before the powers that be. The Ontario Mining Tax Act was drafted by men who were earnest and practical as to purpose, and that Act has stood the test of time. It is, in. truth, a mode! Act, for it has been copied in its general clauses by both Quebec and the Dominion of Canada. We all know that, when a machine is running smoothly, we should leave it alone. The same is true of a well designed, easily applicable, taxation measure, and our present difficulty with respect to this Act arises out of an amendment made in 1916. In that year, you will remember, the Dominion imposed a wartime Income Tax Act. It was described as a wartime necessity and it was intimated t:1 at with the end of hostilities its application would be restricted and finally cancelled. The rate of taxation was modest, and when the Ontario Government was asked that the Dominion Income Tax be allowed as a deductible item, they readily acquiesced. This was looked upon as a temporary arrangement.

Jan 1, 1943

By Gareth Barnes

Traditional approach to improvement is driven by good ideas from good people, but whilst the business case may sound good, when the project is complete, we often have not delivered on expectations.  Why not? The best way to successfully predict the outcomes of an improvement effort in a complex process is through using statistical modelling, which considers the impact of uncertainty within interacting processes.  Applying Theory of Constraints in a model helps leaders to quickly determine: Where the constraining process(es) is/are Where improvements will add the most value How much improvement should be targeted - before the next constraint takes over The confidence level of achieving a given target, which gives us a clear pathway to improvement. This paper presents an analysis of statistical modelling applications at Vale’s Manitoba Operations to achieve the 2025 targets. Using statistical modelling of the Business Structure, the team identified a series of process improvements through a Theory of Constraints Analysis. Value Stream Maps (VSMs) were used to define the required data, apply statistical modelling, and conduct training sessions enabling the Manitoba team to identify the constraining influences. Scenario testing in the Statistical models helped identify the Hierarchy of Constraints and specify the pathway to improvement. How to Improve a Business Video link https://vimeo.com/344418155

May 4, 2025

By I. H. Warren

"An analysis of form-coking processes is presented in terms-of their fundamental mode of operation and their relationship to other coking processes. Their possible future role in cokemaking is assessed. IntroductionAny consideration of prospects and possibilities for form coking must commence with the iron and steel industries-their likely future growth pattern on a world and regional basis and the potential direction of future developments in iron and steelmaking technology.From the numerous estimates which have been made of the future consumption of steel, it would appear that a growth rate of 3lllo between now and the year 2000 is likely. Radical changes in steel making technology are unlikely to be commercially adopted on any significant scale during the next twenty years . However, widespread commercialization inter alia of continuous casting and direct reduction, growth of electric furnace steelmaking and a decrease in coke rates will all contribute to a growth rate for metallurgical coke that certainly will be less than that of steel."

Jan 1, 1980

By Joseph F. Daniels

In recent considerations of the utilization and beneficiation of coal, the chemise and the chemical engineer have been occupying the centre of the stage with their presentation of the picture of new fuel forms. The stage hands and supernumeraries of the cast, the miners and the consumers who carry the brunt of the present-day burdens, are apt to be forgotten in the glorious settings and the climaxes. With no thought of disparaging or of minimizing the efforts of our fellow scientists, let us remember that economic considerations as well as technical difficulties will limit these chemical applications, which must develop and move slowly before reaching the stage of industrial application at some later day. In the meantime, the coal-mining industry must live on hard faces rather than rosy dreams. In these brief notes it is my purpose to call attention to some of the physical advances in the utilization and beneficiation of coal which have been of interest to the industry in the State of Washington.

Jan 1, 1928