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By W. H. Stheeman

"The high rock pressures at the deeper levels of the Tsumeb Mine have necessitated the changing of the mining methods and the introduction of cable bolts. After two rockfalls in the cable-bolted stopes, the mechanics of cable bolting were investigated. Practical methods were developed to find the best grout mixtures, the bond stresses between cables and cement, the bond stresses between rock and cement and the load on the cable bolts, which was mathematically calculated from data observed underground. The design of the cable bolt support system as used at the Tsumeb Mine is explained. GeneralThe Tsumeb Mine is in northern South West Africa/Namibia, approximately 400 km by road from the capital city, Windhoek, at an elevation of 1300 m above sea level. Mining operations in Tsumeb were started in 1905 by the Otavi Minen und Eisenbahn Gesellschaft and continued, with an interruption during the Depre ssion, until 1940, when the mine was closed and taken over by the Custodian of Enemy Property. The present owner, Tsumeb Corporation Limited, obtained the mine in 1947 and restarted the underground mining operations in 1949 after the mine was dewatered. Currently, the stoping operations are between 26 Level (840 m below surface) and 38 Level (1260 m below surface). In addition, the mine is an important silver producer and has produced 670,100 kg of silver since 1947."

Jan 1, 1982

By D. Shishin

Continuous integrated experimental and thermodynamic modelling work is conducted in Pyrometallurgy Innovation Center (PYROSEARCH). The target is to develop a tool, which is capable to predict the compositions of phases, phase equilibria and enthalpy balance for pyrometallugical processes with higher accuracy. All these can be achieved by thermodynamic calculations based on Gibbs energy minimization algorithm. Recent developments made it possible to predict important factors, such as the chemical solubility of Cu in slag as a function of matte grade, temperature, P(SO2), Fe/SiO2 ratio in slag, wt% Al2O3, CaO and MgO in slag. In industrial practice, it is not only chemical solubility, but the overall amount of slag, which affects the recovery of copper and other valuable metals. It will also impact the energy balance. The presence of solids in the slag matrix influences the physical entrainment of matte in slag. Oxygen enrichment which is sometimes used to control the temperature of the furnace, may also affect the volatilization of minor elements into the gas phase. Same is true if additional fuel is used to compensate for heat balance. In other words, all key process parameters mentioned above are interrelated. Present study demonstrates the ways thermodynamic modelling can be used to link those process parameters on the example of copper smelting. INTRODUCTION Research methodology involving integrated thermodynamic modelling and experimental approach has been recently reported by the authors (Shishin, Hayes, & Jak, 2018 Shishin, Hidayat, Jak, Decterov, & Belov, 2016). Present paper reviews the updates and discusses challenges in the development of a large self- consistent thermodynamic database. The database covers the Al-Ca-Cu-Fe-Mg-O-Pb-S-Si-Zn-(Ag, As, Au, Bi, Ni, Sb, Sn) chemical system. The database works in the environment of FactSage software Bale et al., 2016 and related family of products: ChemApp (Petersen & Hack, 2007), ChemSheet and SimuSage. A list of phases, and thermodynamic models is shown in Table 1. For the copper industry, most important phases are Gas, Slag, Spinel and Liquid matte/metal/speiss. Black-colored elements represent the base system, elements in green can be considered as slag modifiers (slagging elements), and elements in red are minor. Ni is highlighted because it was included recently, in 2019, and currently under development. In addition to copper, the database covers the conditions typical for lead and zinc pyrometallurgical extraction. It is developed in consistency with the experimental results for high concentrations of Pb and Zn. Examples of applications in the present study include the thermodynamic analysis of fluxing for an agitated copper reactor to show the effect of Fe/SiO2, and calculations of copper slag cleaning using reductant.

Jan 1, 2019

By M. K. Senicie

IN July, 1937, Asbestos Corporation, Limited, installed an automatic conveyor-belt scale at their King mine, Thetford Mines. The equipment was purchased 'on trial' and its performance has proved entirely satisfactory to the staff. The scale was manufactured by the Toledo Scale Company, of Toledo (Ohio) and Windsor (Ont.) and was installed by one of the Scale Company's engineers. The electrical recording equipment was installed by Keith Mercer, Reg'd, of Montreal, Canadian agents for the manufacturers, the Builders Iron Foundry, of Providence, Rhode Island. The recording equipment consists of a duplex weight-transmitter attached to the scale-head, a duplex totalizer with differential relays, a belt-speed compensator, a rate-of-delivery recorder, and a rate-of-delivery indicator. As it was desired to run the mill at maximum capacity, it was necessary to know the rate of feed at all times, so that any lessening of rate of feed or delays would be recorded and could be corrected. After having determined the rate of feed to the mill (100 per cent capacity), it was required to have weighing and recording instruments which would show: (1) The rate of feed per hour. (2) Any variations .in the load. (3) Any delays and the duration and time of the delays. Rate of tonnage recorders in the mill superintendent's office and at the point of feedi.ng were necessary, as was automatic control between the mill superintendent and the feed man. It was desired to have a permanent record of total tonnage fed per shift or day; weighing and recording mechanism had to have sensitivity to record all previously mentioned conditions. Mechanism had to be of rugged construction, and be so built that it would stay in good operating condition, with the minimum amount of adjustment and supervision.

Jan 1, 1938

By Leonardo F. Trivino

Despite the significant economical and safety implications of poor blasting practice in mining and quarrying operations, the default practice in many large operations, including both design and control, is considerably below standard when compared to other engineering disciplines. Blast dynamics includes a number of complex phenomena, such as the generation, propagation and interaction of stress waves, the expansion of high pressure, high temperature gases from the explosive components, and their combined effects upon the rock mass, including fracturing, fragmentation and displacement. In order to tackle this non-trivial problem, a combined finite and discrete element software (Y-code) capable of accounting for the dynamic behaviour of stress waves, simulating fracturing processes and modeling the expansion of gases is under development. The constitutive models used in this particular software are fundamentally simple, however, non-linearity and chaotic behaviour result as natural consequences of the interaction of relatively simple processes and mechanisms, such as propagation of stress waves in the immediate vicinity of the exploding borehole, and the creation and propagation of fractures. Advances on its development and comparisons with experimental data show realistic propagation of stress waves and fracture networks.

May 1, 2013

By F. Dubuc

"The Adams Mine of Jones & Laughlin Steel Corporation is situated on the Boston Township iron range near Kirkland Lake, Ontario. Eight different orebodies have been outlined, with a total open-pit crude tonnage of. about 100,000,000 tons or the equivalent of 30,000,000 tons of concentrate. The ore consists of fine-grained, well-bedded Early Precambrian iron-formation composed essentially of alternating layers of chert and magnetite. Grade averages about 22 per cent magnetic Fe.The iron range varies in width from 3,000 to 4,000 feet and is made up of several horizons of iron-formation, cherty ""quartzite"" and volcanic rocks. The latter consist mostly of basic lava flows and minor amounts of tuff. Dips are rarely Jess than 50 degrees, and the range is believed to represent an overturned limb of a major synclinal structure. Dykes of syenite and lam prophyre, along with local younger diabase, intrude the rocks of the iron range."

Jan 1, 1966

Particles for heap leaching in conventional circuits are prepared using Cone crushers. Cone crushers are used to crush the ore to the size range between 12-25 mm for heap leaching operations. Leaching from these relatively coarse particles occurs only at the surface and in subsurface regions, which are accessible from the surface by cracks and pores. Since leaching is active in accessible areas, the method of comminution that forms multiple cracks and pores would offer an advantage for heap leach. An investigation was performed in this study to assess if the HPGR forms more cracks and pores than a Cone crusher. X-ray computed tomography (CT) was used to characterize the degree of cracks and pores formed by these two comminution devices that can be used for heap leaching feed preparation. In order to compare the effect of the comminution method on crack network generation, a bulk sample of the zinc ore was crushed by the HPGR at three different pressure settings (45, 95 and 120 bar), the sample ore was also crushed with a Cone crusher for comparison purposes. Products from the Cone crusher and HPGR experiment were then screened into five size fractions ((-25 +19), (-19+16), (-16+9.5), (-9.5+5), (-5+4.75) mm), Subsamples from the ((-25 +19), (-16+9.5), (-5+4.75) mm) size fractions were analysed. The spatial distribution of the crack network of sphalerite particles was characterized using a highresolution industrial X-ray CT system. This paper describes the use of image analysis techniques including image segmentation, which uses a combination of thresholding and other methods to characterize and quantify crack and mineral dissemination in the sphalerite particles after comminution with the HPGR and Cone crusher. The results are validated with those obtained using traditional techniques such as physical gas (with N2) adsorption, mercury intrusion porosimetry, SEM and QEMSCAN.

Jan 1, 2012

By F. D. Ford, Z. Dai, A. Byers

Manitoba Mining and Milling operations consist of three underground mines (Thompson T1, T3 and Birchtree), feeding the Thompson Mill. Three years ago, a collaborative effort involving Mines Exploration, Mines Project Development Group, Mill Technical Services and Vale Base Metals Technical Excellence Center was initiated to establish and verify the cause of high variability in concentrate quality and Ni recovery at the Thompson Mill. One outcome of this work was a program to systematically characterize the mineralogy and metallurgical performance of current and future orebodies feeding the mill. Mineralogy results show that high value enabling ores typically have a high pentlandite to pyrrhotite ratio and benign silicate mineralogy (mostly quartz and feldspar from sedimentary host rocks), while metallurgically challenging ores contain some combination of pyrite and/or hexagonal (non-magnetic) pyrrhotite and/or elevated quantities of talc (with associated carbonates) derived from altered ultramafic host rocks. The metallurgical performance of these two ore types has been simulated by an Excel particle flotation model, which calculates the relative floatability of particles based on the particle’s mineral composition and mineral locking determined from MLA images. The relationship between these mineral properties measured by MLA and metallurgical performance was established using 13 monthly composite samples from Thompson Mill. The model was then validated with results of laboratory flotation experiments on individual orebodies submitted for evaluation. The validated model can be used to predict plant metallurgical performance based on individual ore bodies feeding the mill. Individual orebody’s grade-recovery curves are assumed to be additive, and can be weighted in proportion to the mill feed to predict mill recovery over the same time period. The ore mixing can be done with a simple Excel spreadsheet and comes with a graphical output for easy visualization of the resulting grade-recovery curve.

Jan 1, 2015

By Jun Xiao, Xiaodong Li, Chunming He, Daixiong Chen

"Benzohydroxamic acid with double dentate can react with copper to form cyclic hydroxyl oxime acid copper chelate, indicating a good selectivity for copper oxide minerals flotation, especially for malachite. The flotation results of the pure malachite mineral show a strong synergistic collecting effect between hydroximic acid and xanthate at a ratio of 3.5:1, achieving a malachite recovery of 95.5% when certain sodium sulphide is used as a sulphidizing agent in the optimal pH range. Thanks to the synergistic collecting effect, the flotation recovery of malachite of a real refractory ore flotation ,can reach over 92% under the optimum condition, a 12% higher than the recovery achieved with the conventional collectors. In industrial application, the flotation of the copper oxide minerals is improved , leading to a high efficiency utilization of copper oxide resources.INTRODUCTIONBeneficiation of copper oxide ore is one of the recognized challenges in mineral processing industry. Copper oxide ore is a vital component of copper resource in the world, which has abundant reserves. Majority of sulphide deposits has oxidized zone on the top, and some deposits are oxidized to become the large and medium-sized oxide deposits (Willam & Chavez, 2001). Most copper oxide ores have problems in processing such as high slime consumption, poor floatability, complex ore property and low flotation recovery, so it is of great significance in beneficiation of copper oxide ores to develop an efficient collector, especially for malachite (Fu & Gu, 1997; Wen, 2010; Riet, Kier & Elford, 1980). At present, sulphidization-xanthate flotation is the most important method for treating copper oxide ore. In recent years, due to the decreasing ore grade, the utilization of refractory copper oxide ore draws more and more attention (Zhou, 2003), except for sulphidization-xanthate method, chelating flotation method (Lu, 2009), chemical beneficiation method (Lenormed, Salman & Yoon, 1979), microwave radiation flotation method, hydrothermal sulphidization-flotation method (Mao, Wen & Fang, 2014) and ammonia leaching of sulphide precipitation-flotation method (Lekang, Stevik & Bomo, 2003) have occurred."

Jan 1, 2016

By D. R. Murray

"A promising long-term solution to uranium tailings reclamation has been the establishment of vegetation directly on the waste material. The soil profiles of areas vegetated for 0, 3, 8 and 10years are characterized as to the extent of soil development . The characteristics are compared with the visually interpreted soil profile. Organic matter content is presently the best analytical judge of soil development. Vegetation procedures have produced a build-up of organic matter in the top 10cm of tailings where natural microbial populations flourish. The CN ratio is high and the profile depth shallow, suggesting the need for additional fertilizer applications.IntroductionThe reduction of surface erosion and improved aesthetics of uranium tailings has been accomplished by the establishment of a vegetative cover. In 1971 research concentrated on developing methods of plant establishment at minimal cost and soil amendments on uranium tailings in Elliot Lake , Ontario. One process of stabilization used on a large scale was direct seeding of grasses and legumes into the tailings which had received only limestone and fertilizer treatments. Fertilizer additions were made for varying lengths of time in subsequent years to maintain the established cover.Maintenance-free vegetation would be possible once a soil profile developed with sufficient nutrient recycle and plant regeneration capabilities(1). Tailings vegetated in 1971, 1973 and 1978 that had received either 3 or 5 years of maintenance fertilizers were examined for soil development profiles. The extent of soil development and the relative effect of time and fertilizer would provide direction on reaching the goal of a self-sustaining cover.This paper is an abridgment of Division Report MRP /MRL 81-126 of the same title(2)."

Jan 1, 1984

By G. J. Arseneau

ABSTRACT Mineral resource statements are reported as tonnage (volume × density) and grades. Effort is expended to accurately estimate the volume and grade, but accurately determining bulk density is often overlooked. Oversimplification of density modelling can result in potentially serious errors in tonnage estimates and over- or underestimation of metal content. A reliable bulk density model requires an adequate database and a good understanding of density variability, along with grade and geology. In this paper, the author summarizes density measuring techniques, draws on examples of simple and complex grade-variable density domains, and discusses improved approaches for minimizing density estimation errors.

Jan 1, 2014

By Valérie Tremblay

Plan de la presentation ?Laboratoire d?analyse des fluids ?Gestion des contaminants ?Condition Monitoring Avantages reliés aux analyses ?En savoir plus!! ?pour minimiser les coûts d?opération; ?pour prévenir les défaillances catastrophiques; ?pour maximiser le rendement des équipements; ?pour maximiser la longévitédes équipements; ?pour permettre des périodes d?entretien planifié.

Feb 1, 2008

By J. -F. Blais, M. -O. Simonnot, J. Mocellin, G. Mercier, J. -L. Morel

Large quantities of ferromanganese sludge are generated as waste material by blast furnace during the manufacturing of ferromanganese. These slags are very rich in manganese, zinc and lead (5 to 40 %). These residues have been deposited in a pond in periphery of the steel industry. Considering the market value of the metals, these brownfield sites can now be considered as secondary resources. We have developed a hydrometallurgical process for selectively leaching Zn, Mn and Pb in order to produce compounds of Mn and Zn pure enough to be economically recoverable and a residue rich in Pb. Batch laboratory experiments were carried out to determine appropriate leaching conditions to maximize zinc extraction (100 %) and manganese extraction (90.0 %) with sulfuric acid and to generate a residue containing 15 to 30% of lead. This Pb residue is recyclable by another process.

Jan 1, 2015

By Barrick Pierina, John Frostiak, Igor Gonzales, Kenneth G. Thomas, Ken Manning

"1.0 SummaryThe Pierina deposit was acquired in August 1996 when Barrick Gold Corporation purchased Arequipa Resources, a junior mining company with extensive land positions in Peru. At the time of acquisition, Arequipa had developed two exploration tunnels and drilled 9 exploration holes on the property. Metallurgical tests, conducted on samples taken from the tunnels and the exploration holes indicated that the ore was amenable to heap leaching. Barrick’s exploration team recognised from the outset, the potential of the mineralisation and convinced Barrick’s senior management that this was indeed a world class deposit that should be pursued. On November 25, 1998, the first dore was poured at Pierina, 27 months from the time of acquisition. During that period Barrick completed its exploration programme, prepared a development plan including capital and operating costs, production profiles and cash flow forecasts, and a schedule to develop a mine and construct a leach pad and gold recovery plant, including all supporting infrastructure. All permits and environmental approvals were obtained and all necessary land was acquired. The project was completed on schedule and under budget. This was the first greenfield mine development for Barrick outside of North America.This paper details, mainly, the research and development, engineering, construction and commissioning of the metallurgical process facilities.2.0 The Develonment ProgrammeImmediately after the acquisition, Barrick began the process of gathering information required to prepare a project development plan. Kilborn Engineering was retained to assemble all of the information collected and collated into a development plan document. This document was completed by June 16,1997. The objective was to bring the mine into production as quickly as possible with critical activities identified as follows:"

Jan 1, 2000

By Eric Bain Wasmund, Lance Christodoulou, Julio Jairo Carmona Franco, María Fernanda Castillo, Jose Concha

"A number of technologies are in practice for selectively removing coarse gold from conventional grinding circuits. These circuits are often designed so that the grinding mill output is size classified, with large (or high density) particles being removed before the tails are returned to the mill. Because of the high relative SG of gold, and the fact that some forms of gold such as free gold do not break apart during conventional grinding, coarse gold can accumulate in the circulating load, and eventually be lost.Eriez has developed a flotation cell that uses fluidized water to enhance flotation of coarse particles. More than 40 units have been sold in industrial minerals, and we are now exploring applications for the HydroFloat® in the gold industry.Mineros S.A currently produces approximately 120,000 oz of gold per year by gravity concentrating alluvial gold ore with a mean particle size of 330 microns, followed by further concentration using a Falcon separator. A 400 mm diameter pilot-scale HydroFloat® was added to the tails of the Falcon to test whether additional gold could be captured. Enrichments of 50-600 times and recoveries normally in excess of 95% were achieved. This allowed direct smelting of the gold concentrate. Metallurgical results and operating parameters such as water addition and power will be discussed during the presentation.This result opens the possibility that the HydroFloat® could be used directly in the grinding circuit, for removing coarse gold in the circulating load. This work also shows the possibility that the HydroFloat® can produce a throwaway tail, which will reduce the size of the circulating load stream and the grinding requirements of the circuit."

Jan 1, 2015

By C. C. Li

Traditionally, strong and stiff devices, such as fully encapsulated rebar bolts and concrete lining and arches, are preferred for underground rock support. The attention is paid to the strength of the devices in such support designs. It has been observed in engineering practice that support systems composed of stiff devices are often subjected to premature failure in high stress rock conditions. In the past decade, yield support devices have been employed to improve the performance of rock support systems. A rock support system usually consists of both internal and external devices that are installed within the rock mass and on the free surface of the rock, respectively. Great efforts have been made to make support devices yieldable in the past years, but the efforts are focused on different aspects in the tunneling and mining branches. In the tunneling branch, the focus has been put on the development of external yield devices, such as Lining Stress Controller (LSC), Honeycomb element and compressible concrete, which are embedded in shotcrete liners, while in the mining industry the efforts has been on the development of yield rockbolts, such as the cone bolt and the D-Bolt. The internal and external support devices are not compatible in their deformation capacities in the current rock support systems even though yield devices have been employed. All the devices in a support system should be deformable and compatible in deformation capacity. With such a support system, premature failure of individual support devices could be avoided and the service life of the support system is prolonged. It is a trend at present to take into account the energy absorption capacity, instead of simply the strength, of support devices in rock support design.

Jan 1, 2015

By N. Dufresne

"The evolution of mining methods at Brunswick Mining and Smelting shows that, in the case of a large deposit, technology must adapt itself to changing conditions. As problems are encountered, management must modify their approach if they hope to meet production quotas. This search for new and better ways started with the opening of the No. 12 orebody and is still going on today.From large-scale open stoping, the Brunswick Mining and Smelting operation went to transversal cut-and-fill, then to longitudinal cut-and-fill, which is the main mining method now employed. However, to solve certain problems experimenting with Avoca methods was initiated. Rebar grouting and prepinning have also made significant contributions toward securing relatively poor ground in the hanging wall and holding badly fractured backs in wide stoping areas. Introduction In 1952, M.J. Boylen had incorporated Brunswick Mining and Smelting Corporation Limited to develop his recently found No.6 and No. 12 orebodies near Bathurst, New Brunswick. Between 1953 and 1957, St. Joseph Lead Company was brought in to try and initiate mining . During those years, St. Joseph Lead sunk two shafts at No. 12-No. 1 to 425 feet below surface and No.2 to 982 feet. They also developed the 200 and 350 levels to sample the ore at depth and operated a 100-tpd pilot plant at the No.6 mine. However, due to low recoveries and depressed metal prices. the project was shut down in 1957.A group including M.J. Boylen. K.C. Irving. Maritime Mining Corporation and Patino Corporation reacquired the property from St. Joseph Lead Company in 1961. New financing was arranged involving La Societe Generale des Minerais, S.A . of Belgium and Engineering Consultants Limited of Saint John. New Brunswick. During the years 1962 and 1963, construction of surface and underground facilities progressed rapidly."

Jan 1, 1981

By R. R. Bassermann

THE property of Howey Gold Mines, Limited is situated at Red Lake, Ontario, one hundred and seventy-five miles east of Winnipeg and eighty miles north of the Canadian National Railway line. The mine has been a continuous producer since April, 1930, and at present is hoisting 1,350 tons of ore per day. The shaft consists of three compartments (see Figure 1, sketch B). Hoisting is done by two skips in balance, one of which is interchangeable with a cage. Sinking had been carried out at three different periods. From the surface to the 550-foot level, the sets had been spaced at seven-foot intervals, from the 550-to the 1,060-foot level at six feet, and from the 1,060-to the 1,500-foot level at five feet. During the five years of operation, the shaft had become so badly worn, due to spill from the tipples and, to a certain extent, from the loading pockets, that re-timbering was considered necessary. It should be explained that no crushing is done underground because of the ad verse effect this would have on the efficiency of the sorting. The grizzly rail clearance above the ore pockets is fourteen inches and consequently the ore hoisted is extremely coarse. More than average spill from the loading pockets and tipples under such conditions is inevitable, and it is debatable whether underground crushing with consequent increased sorting costs would not be justified. The cost of complete re-timbering according to standard practice would have been prohibitive both directly and indirectly: A lengthy shut-down of the shaft, with the consequent disorganization of the staff and workmen, could not be considered. Accordingly, it became necessary to devise some method of repair that could be carried to completion in a very limited period of time. The shaft timber for the first 250 feet was not in need of repair. From the 250-foot to the 1,060-foot level the wall plates, posts, and blocking were in good condition. These had been protected by a one-inch plank lining put in at various periods during the past three years without interfering with hoisting operations. The end plates and dividers, however, could not be protected in this manner because of the lack of skip clearance. It was decided that, if the dividers in this section could be replaced, the shaft would be as satisfactory as if completely re-timbered. Several factors had to be taken into consideration in deciding on a suit-able method of repair. In the first place, the tipple plates had to be re-designed. To complete the job in a reasonable time, not only had work to be done on the tipple, but simultaneously from several points in the shaft itself. Since all power cables, pipe lines, etc., were in the manway compartment of the shaft, this compartment had to be left undisturbed.

Jan 1, 1935

By A. L. Mular

This paper provides a worthwhile contribution to the knowledge needed for calculating and using mathematical models of grinding circuits. The computer search algorithm proposed should remove much of the guesswork as well as reduce computer processing costs and improve the accuracy of estimated parameter values for grinding mill models.

Jan 1, 1970

By C. E. Cartwright

That the construction of the Pacific Great Eastern railway has not been accompanied by an immediate and great development of traffic is not due to lack of natural resources in the district traversed, but to conditions that have delayed their development. Prior to the commencement of construction of the railway there was no inducement to prospect for any other mineral than gold, on account of lack of transportation. The war began soon after the commencement of construction; diverting practically all capital that might have been available for opening up the country; and since the war, capital has continued shy of new investments. If the war had not occurred it is probable that prospecting and development would have been carried out concurrently with the railway construction. Moreover, much of the district is covered deeply with drift, marine and alluvial deposits, rendering prospecting for lode mines a more expensive and slower proceeding than can generally be undertaken by the ordinary prospector, without capital. On the other hand, the presence of large porphyry dykes, granitic intrusions, and the finding of rich float in places are indications that rich miIY.ral deposits probably exist in several localities in the Lillooet district, in the Cariboo proper, and also nearer the coast. Two or three mines producing the tonnage of the? Britannia would materially alter the earnings of the railway and pay interest on the bonds. There are, however, other possibilities, in addition to large metaliferous mines, in the non-metallic minerals of the district.

Jan 1, 1925

By R. D. Brown

The first step in achieving a rationalization of Canada's crazy patch-work quilt of mining taxes and policies was achieved with the release last November of a major federal-provincial report on the taxation of the Canadian mining industry. The Report, prepared by a group of federal and provincial civil servants, sets out a concise and informative discussion of the major taxation and incentive issues facing the mining industry, and puts forward major recommendations for more rational and moderate taxation policies. It will be considered by federal and provincial Finance and Resource Ministers for possible action by the respective governments. The Report was prepared in response to a directive from the federal and provincial First Ministers in February 1978 to review the federal and provincial taxation of the resource areas. Although the Report is more successful in identifying problems rather than concrete solutions, and reflects some continuing differences of opinion between the governments involved, it does represent an important first step in a comprehensive review of our resource tax policies. This article summarizes the Report, which is some sixty pages in length. Although a detailed discussion of it is beyond the scope of this article, the report itself is recommended reading for those concerned about the current state of the Canadian mining industry.

Jan 1, 1979