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By P. R. Jones

"The paper describes the CDN$400 million expansion at Hudson Bay Mining and Smelting Co., Limited’s operations. The company has continuously produced copper and zinc at the Flin Flon metallurgical plant since 1930. The paper details the company’s operational turnaround since 1995 and “The 777 Project” which includes two new mines, expansion and modification of the concentrators, expansion of the zinc pressure leach plant, a new zinc electrolytic cellhouse and modifications to the copper smelter gas handling system to reduce ground level gas emissions. IntroductionHudson Bay Mining and Smelting Co., Limited (HBMS) was incorporated in 1927 to mine and process the world-class Flin Flon copper and zinc orebody. Ore concentration and metallurgical facilities were constructed in Flin Flon. The metallurgical complex capability is currently 89 000 tonnes of copper metal and 98 000 tonnes of zinc metal from both HBMS and purchased concentrates.Flin Flon, with a population of 7000, is in Manitoba and partly in the province of Saskatchewan. The location is isolated with Winnipeg 800 km to the south and Saskatoon 500 km southwest.Today, three underground mines feed the Flin Flon concentrator, one feeds the Ruttan concentrator at Leaf Rapids, 200 km to the north of Flin Flon, and Chisel North mine feeds the Snow Lake concentrator (Fig. 1).The Flin Flon volcanic massive sulphide proximal and distal orebodies are within the Flin Flon greenstone belt which stretches from the west of Flin Flon to the east of Snow Lake, a distance of 250 km. The belt is approximately 100 km wide before disappearing below the paleozoic cover to the south."

Jan 1, 2001

By Y. Prevost, D. Verhelst

"A computer-assisted process control system for the Noranda Process Reactor Wa5 installed at the Home smelter in 1991. Operating as an advisory tool for the furnace operators, and using a flexible oxygen balance as the basic building block of the computer program, the new control system has helped to reduce the variability in the grade of the matte produced by 50% . The smelter is currently operating at a 72% copper matte, with a standard deviation of between 1.0 and 1.3, while treating a wide range of concentrates, secondary materials and scrap. Enhanced process control has allowed for longer reactor campaigns, and improved stability of the converting operation.This paper reviews the feed classification and blending practice used at the smelter, outlines the strategy of the new computer control system, and summarizes the benefits realized since the implementation of the system in early 1990.IntroductionUntil 1989 the Noranda Home smelter operated as a multi furnace smelter, at one time having three hot-charge reverberatory furnaces, and finally operating with one wet-charge reverberatory furnace and one Noranda Process Reactor. In 1989 the final reverberatory furnace was closed and the Reactor became the sole smelting vessel, treating around 750 000 tonnes of concentrate, secondary materials, and scrap per year. At the time of the start-up of theNoranda Reactor in 1973, the vessel was capable of treating 750tonnes per day of a heterogeneous mixture of copper concentrates, fluxes and secondary materials to produce copper in one operation. Through process improvements, and the introduction of additional oxygen in 1982(pannell and Mackey, 1988), the daily tonnage increased to 1800 tonnes by 1988. The operation had also been changed to produce a high-grade matte instead of copper. Figure1 illustrates the main features of the reactor, a cylindrical vessel which is charged at one end and tapped at the other. The matteis subsequently processed through Peirce-Smith converters and anode furnaces before being cast into copper anodes. Figure 2 outlines the progressive increase in the furnace smelting capacity. Over this period the percentage of secondary materials, and scrap feeds also increased to approximately 110 000 tonnes per year, or 15% of the total new feed input (Bedard et al., 1991). Now operating wholly as a custom feed operation, the Home smelter can treat over 2200tonnes per day of new metal-bearing material."

Jan 1, 1993

By P. E. Auger, J. E. Gill

"The zinc deposits to be described are in Lemieux township a few miles south of the Tabletop mountains, in the central part of Gaspe peninsula, Quebec. Although they have not been worked, they constitute a potential source of zinc and lead and structurally are of considerable interest.GENERAL CHARACTER OF THE OCCURRENCEThe deposits are in sharp-walled quartz-carbonate veins and breccia zones carrying locally, and in varying amounts, sphalerite, galena, pyrite, marcasite, chalcopyrite, and specularite. Those so far investigated strike generally north-south, have steep dips, and cut through sediments of Lower Devonian age, mostly calcareous argillites or siltstones. Wall-rock alteration is negligible.BROADER STRUCTURAL FEATURESThe principal known occurrences are in the south-central part of a dome structure measuring 6 by 5 miles. The argillaceous sediments are overlain, on the flanks of the dome, by sandstones and lavas. These have been stripped from the central part of the dome, revealing the veins in the argillites beneath. The veins formed after the volcanics, but the fractures and breccia zones in which the deposits formed are confined mainly to the argillite series, presumably because of the greater strength of the lavas.The argillite series is cut by dykes, sills, and small stocks of diorite and acidic 'porphyries' that are older than the veins.THE FEDERAL HILL AREAThe distribution of known veins is shown in Fig. 1. The characteristic features of the veins are well illustrated by the Federal Hill area shown in Figs. 2, 3, and 4, and described below."

Jan 1, 1954

By R. Mulligan

"The Cordilleran tin belt extends from the Seward Peninsula in westernmost Alaska to Mexico. Throughout the Canadian section, tin occurrences are mainly confined to a belt along the northeastern fringe of the Western Cordilleran sub-region; the boundary of this sub-region with the Eastern Cordilleran sub-region is most simply defined as the practical eastern limit of granitic and metamorphic rocks. The stratified rocks underlying this restricted belt are mainly late Precambrian and early Paleozoic miogeosynclinal sedimentary rocks, which are more or less metamorphosed but otherwise are generally equivalent to the less severely deformed rocks underlying the Eastern Cordilleran sub-region. They contrast with Late Paleozoic - Mesozoic eugeosynclinal volcanic-sedimentary assemblages farther west, but the granitic intrusions throughout are mainly of Late Mesozoic age. Thus, narrowly defined, the tin belt largely coincides with a lead-zinc-silver metallogenic province, whereas copper deposits characterize the eugeosynclinal regions to the west. In addition to the tin occurrences, which include the Sullivan lead-zinc-silver mine, Canada's only significant producer, the belt contains all the beryllium, most of the tungsten and many Iithophile-type molybdenum occurrences. Several types of occurrences of each metal are represented. Lithium occurrences in the Canadian Cordillera are notably rare, and these few are insignificant."

Jan 1, 1971

By Gary Hawthorn

"Laboratory testing, by Lakefield Research determined that, in spite of the relatively high pyrite content of the Golden Rose (GR) orebody, it would respond well to gravity concentration followed by either direct cyanidation or cyanidation of a flotation concentrate.Economics favoured the former choice which became the adopted process flowsheet.During the latter stage of the feasibility study, the Dekalb 500 tpd copper flotation mill became available in the Highland Valley of British Columbia. This plant was purchased by Emerald Lake Resources, and was moved to the Golden Rose minesite.The crushing and grinding circuit were retained and the original flotation circuit was replaced with a 400 mtd gravity plus cyanidation circuit. The plant commissioning commenced in August 1987, and by early September was processing 300 tpd.HistoryThe Golden Rose Mine was operated by Cominco during the period 1935 to 1941 processing 146000 tons grading DAD oz/t Au at a rate of 110 tpd .The process flowsheet consisted of gravity plus amalgamation followed by cyanidation.Metallurgical TestingThe underground mining areas were below the level of Emerald Lake, and had flooded on abandonment in 1942.Samples for metallurgical testing were collected from exposed faces underground once the mine had been dewateredin 1984.In spite of the anticipated high pyrite content of the ore samples, it was apparent from the success of the previous operation that the pyrite did not interfere with the gold recovery.Laboratory testing, at Lakefield, determined that, from a technical perspective, two metallurgical options existed, as shown in Tables I and 2.Note that the very high pyrite content of the ore precluded the possible direct marketing of concentrate. The process flowsheet, therefore, had to contain a cyanidation circuit for the operation to succeed financially.Although the anticipated metallurgy was excellent with either option, the 2.1 % increase in gold recovery with the gravity plus cyanidation option made this the clear choice."

Jan 1, 1989

By F. J. Tozer

Canal communication between Lake Erie and Lake Ontario has been one of almost constant enlargement and reconstruction to meet the rapid growth of trade and commerce, and the consequent steady increase in the size of vessels plying the Great Lakes. The original canal was built by the Welland Canal Company. The first sod was turned in 1824 and the first vessel passed through in. 1829. There were 36 lift-locks, built of oak timber, 110 feet long and 22 feet wide, with a depth of water over the sills of 8 feet. The second canal, now known as the 'Old Welland Canal' was built by the Dominion of Canada (1842 to 1850). There were 26 masonry locks, 150 feet long, 26.5 feet wide and with a depth of water over the sills of 9 feet. The walls and banks were raised in 1853 to give a 10-foot draught. The third canal, known as the 'New Welland Canal, built by the Dominion of Canada, was started in June 1874 and the first boat passed through in the spring of 1882. There are 25 masonry lift-locks, 270 feet long, 45 feet wide and with a depth of water over the sills of 12 feet. The walls and banks were raised in 1891 to give a 14-foot draught. This is the canal in use at present.

Jan 1, 1926

By Harry V. Warren

THE Dolly Varden Mine is situated in the Kitsault River valley, eighteen miles from the head of Alice arm, in the Nass River mining division of British Columbia. During the years 1919-21 it was operated by the Taylor Engineering Company, Limited, and produced 1,300,000 oz. silver from about 36,000 tons of ore. The ore was transported by narrow-gauge railway to Alice arm, and shipped to the smelter at Anyox for treatment. During recent years, a small quantity of high-grade ore has been extracted by leasers and shipped to the smelter at Tacoma, Wash. Unfortunately, the bonanza ores which occurred near the surface did nor persist to depth, and two hundred feet below the exposures only rare and unimportant patches of ridi ore have been found. Two opinions exist as to the origin of this rich silver ore. One, advanced by V. Delmage (1, 2), is that the ores are hypogene, and the other, by G. Hanson(3, 4, S), is that the ores are the result of supergene enrichment. It is true that Dolmage first advanced his view that the ores are primary because they appeared to be analagous to Premier ore, which he had studied in detail and which he considered to be primary. The present authors have had access to excellent suites of material from both mines and consider the comparison well taken, the only essential mineralogical difference being the presence at the Premier mine, in addition to the silver minerals, of gold and electrum. Although the arguments put forward by Delmage have, in the opinion of the authors, never been refuted, nevertheless R. P. D. Graham(6), in a reference to the mine published in 1924, describes the ore as of supergene origin, and since then nobody appears to have challenged this statement. Indeed, in 1935, Hanson re-affirmed his earlier conclusion(5).

Jan 1, 1942

By N. B. Clarke

LIMESTONE, as a name, covers a great variety of stone consisting of varying amounts of calcium carbonate, magnesium carbonate, and impurities such as silica, alumina, and sulphate of lime. In the iron and steel industry, limestone is roughly divided into two classes, namely Calcite, when the magnesium carbonate content is below 10 per cent, and Dolomite, when the magnesium carbonate is over 10 per cent. In practically all metallurgical melting, the quality of the finished product is related to the character of the slag, and limestone is used as a flux to achieve the proper analysis and consistency. The function of the slag is to absorb the undesired elements of the bath and also to act as a protective cover over the metal. The carbonates of calcium and magnesium are the effective agents in fluxing, and the silica, alumina, sulphur, and other foreign matter are objectionable impurities which should be as low as possible, totalling not over S per cent and preferably below 3 per cent, since all of these so-called impurities must be fluxed by the lime and magnesia of the stone. When the lime and magnesia of the stone have to be used to flux its own impurities, there is that much less of these two oxides left for fluxing the impurities from the balance of the charge, and, as a result, more lime-stone must be used than is desirable with good practice.

Jan 1, 1939

By Paul F. Grundy

The ever changing conditions in the coal-mining industry and the increasing competition of other fuels such as oil, natural gas, wood refuse, etc., have made it imperative that the coal operator be in a position to supply a constantly uniform and clean product. The old idea of putting on the market "anything so long as it's black" no longer holds good. Today's users of fuel have to consider the amount of heat per dollar they obtain from the fuel they buy. Burning of coal has become largely mechanical, at least so far as industrial coal is concerned, and industrial coal is the kind the bituminous operator is for the most part called upon to supply. When we get into things mechanical, the matter of precision has a mighty bearing on efficiency. Precision in coal means standardization of quality, or uniformity of preparation in size and also in ash and moisture content. The trend of coal mining is ever towards more and more mechanization. Mechanization means coal-cutting machinery, loading machinery, conveying machinery, etc. Although such methods cut the cost, the cleanliness of coal so mined and loaded is not controllable to the same extent as when each operation is by hand labour. The operator has therefore the problem of mining his coal as cheaply as possible, but at the same time preparing a product which will to the utmost satisfy his customer. The answer is a mechanical cleaning plant. Before installing a cleaning plant, however, the operator must take into consideration the cleansability of the coal from his particular mine. This is accomplished by taking representative samples of the seam or seams to be treated, and subjecting these to screen tests, followed by float and sink tests on each size product. From these tests, it can be readily found to what degree of cleanliness the several sizes can be prepared advantageously, and also what will be the percentage recovery of cleaned coal. The accompanying tables give the results of representative tests on Michel coals of minus-2-inch size. All coal over 2-inc.h was intended to be hand cleaned. The coal-cleaning plants in general use at the present time come under two broad classifications, namely, wet and dry.

Jan 1, 1934

By A. Dubé

Introduction It is common experience that on re-heating a fully martensitic steel below the Ac, point, a softer and tougher structure is formed as the tempering temperature is raised. The mechanism of tempering of plain carbon and low-alloy steels has been the subject of many investigations. In a very comprehensive study, Cohen and associates have recently established three different tempering stages, the last of which reaches completion at about 700°F., with heating rates of 5°F. per minute. This latter stage has been established ta be the carbide-forming stage, so that the steel consists of a mixture of ferrite and carbide, which are the metastable equilibrium phases. Any subsequent changes which take place at higher temperatures below the Ac, point in the absence of graphitization or carbide lattice conversion normally consist of agglomeration of the carbide particles. This naturally results in the softening of the steel. However, in many cases the Ac, point, as determined by dilatometry, thermal analysis, etc., only gives a rough estimate of the temperature where the reaction, ferrite plus carbides to austenite, is initiated. This appears to be especially true in the case of alloy steels in which three phases can co-exist in certain temperature ranges. Bain has given data on the three-phase equilibrium in both manganese and nickel steels. In these steels, the formation of austenite at temperatures just above the AeIL apparently involves the diffusion of the alloying elements to the first pools of austenite, thereby making the process very sluggish. As a consequence, the critical Ac points determined by ordinary methods at normal rates of heating can be quite distant from the true equilibrium Ae temperatures.

Jan 1, 1946

By Marvin J. Udy

THE preparation of this paper is in response to many requests for in-formation on the mining and smelting developments which have been carried on for the past three years or more by the Chromium Mining and Smelting Corporation of Canada. By way of introduction, the following paragraphs, relating to the early history and present status of the chrome-mining industry in the United States, are of interest. They are quoted from Bulletin No. 76 of the California State Mining Bureau, published in August, 1918: 'Chromium was discovered about 1797 by a French chemist, Louis Nicolaus Vauquelin, in ore from Siberia. From that time until 1827, the Urals, in the region of Ekaterinburg, supplied the chromium used in . Europe. In the summer of 1827, as the historian advises us, Isaac Tyson, Jr., saw in a Baltimore market-place a cart containing a cider barrel, which was held from rolling about by some heavy black stones. He had made a study of such stones at the first-known American locality, the Bare hills near his father's home, six miles from Baltimore, and he recognized the black stones as chromite. On inquiry, he found the chromite came from the Reed farm in Harford county, 27 miles northeast of Baltimore. At this time, Tyson was one of the very few men in the country who knew the value of chromite, and in the next few years he took full advantage of this knowledge. He immediately bought the Reed farm. He found about 30 tons of float-ore on the surface, but there was no outcrop of ore in place. Never-theless, he sank a shaft, and at a depth of eight feet struck an ore-pocket which proved to be 80 ft. long, 25 ft. wide, and 8 ft. in maximum thickness. His next discovery was of placer chromite in the sands in the beds of brooks on an estate called Soldiers' Delight, 16 miles northwest of Baltimore. Thus far, Tyson's observations had shown him that chromite occurred apparently only in serpentine areas. His explorations were now leading him farther afield. In 1828, he found on the Wood Farm, in Lancaster county, Pennsylvania, a deposit of float chromite which led him to lease the ore-right of the farm, which he finally purChased in 1832. Here he developed the famous Wood mine, the largest single producer of chromite in the world.

Jan 1, 1938

By A. Shaak

The town of Hedley is some 200 miles due east of Vancouver and lies in the bottom of the Similkameen River valley, at the junction of Twenty Mile creek, at about the 1,700-foot elevation above sea level. The Nickel Plate mine camp is on Nickel Plate mountain at an elevation of about 5,600 feet, four miles slightly east of north from Hedley, to which it is connected by 9 miles of road, and by 10,000 feet of tramway and electric railway. The latter is used both to transport the ore to the 275-ton mill at Hedley and to furnish the mine with supplies. History The Nickel Plate claim was staked in 1898 to cover the sole outcrop of the Nickel Plate system of orebodies. Soon afterwards it was optioned by the Marcus Daly interests of Butte and underground exploration commenced. Early exploration was successful, and a power plant, tramway, and mill were constructed. The mill began operations in 1904. Production under Daly control ensued from June, 1904, to August, 1909, and amounted to 167,000 tons. The Hedley Gold Mining Company was organized and commenced operations in 1909. The capacity of the mill was increased by stages to 200 tons and a steady production, except for short periods in winter, was maintained until 1930. By 1930, gold to the value of nearly $10,000,000 had been produced since the discovery in 1898. The South American Development Company optioned the property in 1932. Active exploration was conducted during 1933 and 1934 by the Kelowna Exploration Company, a subsidiary of the South American Development Company. This work was sufficiently successful that, coupled with the increase in the price of gold from $20.67 to $35.00 per ounce, production re-commenced in December, 1934, and continued uninterrupted until July, i946, when, along with all B.C. gold mines, a strike closed operations. Under the present ownership, the mill capacity was increased to 275 tons.

Jan 1, 1946

By A. R. Chambers

Malagash received its name from the Indians many years ago on acco.unt of its turbid (milky) waters, although,' the red men did not appreciate the significance of this turbidity, and were, of course, entirely ignorant of the crystal beds of pure salt that lay beneath their feet. The salt lake of the Mississippian, or Lower Carboniferous period, in which these beds occur, was apparently a large one, for there is evidence of its existence at intervals all the way from Sydney, Cape Breton, to St. John, New Brunswick, 300 miles as the crow flies, and there is no knowing what its other dimensions may have been, as much of this horizon is now buried on the one side beneath the waters of the Gulf of St. Lawrence, and on the other underneath the Atlantic ocean. About midway of the area and underlying the counties of Colchester and Cumberland, there occurs a synclinal basin of which the Malagash salt outcrop forms only a small portion of its northern limit. This basin has been described very fully by A. C. Hayes. (1) The operations carried on at Malagash since Hayes wrote his report have brought to light very little in the way of new geological data, except to confirm his pre-dictions as to the probable faulting of the salt strata, and the effect which these faults would have on mining operations.

Jan 1, 1924

By D. C. McKechnie

To the examining engineer, the ever-recurring problem involving the valuation of ore in a prospect presents at different times highly divergent aspects- some entirely physical or mathematical, others essentially personal. The majority of the problems discussed in this paper are obvious to the engineer, and need little elaboration. By the prospector and the general public, however, a number of these questions are often and completely over-looked, or else approached in such a way as to produce final results which are, to say the least, misleading and at times decidedly erroneous. While the purpose of this paper is to deal with prospects or potential mines, many of the problems considered apply equally well to mines with developed ore reserves. As the usual showings on a prospect consist chiefly of a few open-cuts and a limited amount of underground development work, an estimation of the actual ore reserves is a question of minor import, and seldom enters appreciably into an estimation of the probable value of the particular property. The physical valuation of a prospect mainly involves the consideration of such subjects as the general geological features exposed, the strength and type of mineralization, the width and persistence of any associated veins, shear zones, fractures or contact zones, together with the possible grade, size, and continuity of any ore-bodies that may there occur.

Jan 1, 1930

By A. M. Christie

Acknowledgements The following account has been compiled mainly from records and maps of Eldorado Mining and Refining (1944), Limited, hereinafter called the Crown Company. This work was done under the supervision first of R. F. Murphy, Chief Geologist, and later of E. B. Gillanders, Eldorado Mine Manager and Manager of Eldorado Exploration Division. W. F. James and B. S. W. Buffam are consulting geologists. Since 1946, S. N. Kesten has supervised geological work for the Company in the Goldfields area. Where maps of Eldorado properties have been included in this paper, the name of the geologist who examined the property is acknowledged on the map. Outline of General Geology The geology of the area is described in the Goldfields and Martin Lake preliminary maps of the Geological Survey and accompanying Paper 49-17. In brief, the Tazin series of Precambrian age, consisting mainly of sediments with minor interbedded volcanics and basic intrusives, is intruded and replaced by granite, now generally gneissic. The late-Precambrian Athabaska series of conglomerate and arkose, with interbedded basalt and andesite flows, rests unconformably on the older rocks. All these rocks are cut by small diabase dykes. The geological terms used in this paper and in the accompanying figures are standard with the exception of the term 'mafic rocks'. These include all the basic or dark coloured rocks of the Tazin group, such as amphibolites, biotite schists, and chlorite-epidote rocks (greenstones), whether they be of extrusive, intrusive, or sedimentary origin.

Jan 1, 1949

By J. Sliwinski, H. Kikutani, J. Atwood, D. Larimer, J. Pyecha

"The Pogo mine start-up milling targets of 2005 are vastly different than the targets of today. Improvements have been driven by unintentional design oversight, continual circuit optimization or optimization due to changing mineralogy. The more significant process improvements to date have occurred in the gravity, flotation and leach/CIP unit operations. Details of the significant improvements are presented along with the impacts to the process.INTRODUCTION Deposit Overview Pogo mine is located along the Goodpaster River, southeast of Fairbanks, Alaska. This deposit sits in the center of the Tintina Gold Province (Figure 1), a region of similar Intrusion Related Gold deposits in Interior Alaska, the Yukon and British Columbia Canada (McCoy et al., 1997). The Pogo deposit is hosted in a rock sequence of amphibolite-grade gneisses intruded by mid-Cretaceous granitic plutons and dikes, which in turn, were cut by and host the auriferous quartz veins of the Pogo deposit.The Pogo deposit is comprised of three distinct zones of auriferous quartz veins spatially adjacent to each other. These are the Liese Zone, North Zone, and the East Deep Zone. The Liese Zone consists of three stacked, massive quartz veins, designated the L1, L2, and L3. Vein thickness averages 3.0 m, with variations from 0.3–18 m thick. The North Zone mineralization is comprised of four veins with more consistent thickness, between 1 and 1.5 m. The East Deep deposit lies northeast of the Liese Zone ore body. It is separated by a post mineralization, quartz diorite pluton. The East Deep Zone is comprised of two stacked massive quartz veins with similar orientation as the Liese Zone."

Jan 1, 2017

By Dwayne D. Tannant

The Fox Pit at EKATI Diamond Mine reached a depth of approximately 150 m as of September 2005. The pit is designed with 11 m wide benches excavated using a double benching technique leaving 30 m high bench faces. A series of blasting tests were conducted to improve catch bench reliability and increase the percentage of benches that meet an 11 m catch width. The design changes focussed on reducing damage to the bench crest because maintaining the bench crests is the most important factor in achieving good bench reliability. The most significant improvement was obtained by raising the bottoms of the lower trim blastholes by 0.5 m. This change decreases the depth of broken rock below the final bench elevation, and consequently reduces damage to rock that forms the bench crest below. Another improvement was the use of smaller diameter (165 mm vs. 270 mm) blastholes in the A and B rows for the upper trim blast. This change reduces the charge per delay during detonation, reduces the powder factor slightly, and reduces the energy affecting the final pit wall and crest. As mining and the blast testing progressed, the portion of the bench with crest achievements (>11 m) increased from 57% to 75%.

May 1, 2006

By A. S. Baillie

THE subject of taxes appears to engage our attention altogether too lightly. At least it would seem that, notwithstanding the increase in the cost of government, and its concomitant increased taxation, little constructive work is being done with a view to solving the problem of increased taxes. I believe that the inattention of the public generally to the subject of increased taxation is due largely to the fact that shareholders are not advised of the amount of taxes paid by industry. Were companies to advise their shareholders the amount of monies paid or payable out of each year's earnings for taxes in one form or another, and were such amounts compared with the amounts available for distribution to shareholders, the effect would be a general demand for more equitable taxation methods. While I am not unmindful of the burdens and inequities in taxation confronting industry generally, I believe that the mining industry in this Province does not fully appreciate the burden represented by the costs of taxation. My reason for this belief is based partly upon1the fact that no serious effort has so far been made to devise some means By which the mining industry in British Columbia might protect itself against the ever-increasing burden of taxation. In addition to this fact the widely diverse nature of mining in British Columbia makes the accumulation of comparable data a problem which, so far as I can ascertain, has never been attempted in a manner which would be authoritative. It should not be necessary to cite the effects of this burden upon any one mining company or upon any branch of the mining industry, such as coal mining, placer mining, base-metal mining, or the precious-metal mines. In my opinion, the operation of our mines by large corporations is conducive to the payment of taxes with much less protest than would be the case were the taxes being taken out of the earnings of just a few partners or a very few shareholders. We appear to be too unmindful of the taxes, paid by mining companies, of a kind that are not so prominently brought to our attention as are the taxes which are calculated upon the taxable income of the company at the end of its fiscal year. British Columbian mining companies pay many other taxes besides those which are calculated upon taxable income. Among these are school taxes, property taxes, gasoline taxes, mineral taxes, coal royalty, to say nothing of fuel-oil tax, sales tax, and excise taxes.

Jan 1, 1935

By F. McCarthy, R. McDonald, G. Woodbridge, G. Brock, D. Robinson

"The Direct Nickel process is a hydrometallurgical technology being developed to treat nickel laterite ores by leaching at atmospheric pressure. The leaching process uses nitric acid in a relatively large excess, is thus non-selective, and much of the soluble iron goes into solution. The first stage of solution purification is iron hydrolysis and the Direct Nickel process utilises the colligative properties of the pregnant leach solution (PLS) to enable the removal nitric acid of varying concentrations via distillation. Once sufficient acid has been removed, the boiling point of the solution is increased allowing high temperatures (up to ~170°C) to be achieved under atmospheric pressure. The higher temperatures initiate the precipitation of the iron primarily as hematite and allow the removal of the acid generated by the hydrolysis reaction. The rejection of the iron is near complete with solution tenors typically falling from 65 g/L Fe, or more, to less than 50 mg/L Fe. The separation of clean nitric acid from the PLS at this stage allows for relatively simple acid recycling. A hematite product with iron content above 60% could also provide an extra revenue stream.INTRODUCTION The Direct Nickel (DNi) Process is a novel hydrometallurgical processing route designed to treat all types of nickel laterite ores in a single flow sheet using nitric acid under atmospheric conditions. A simplified flow sheet is shown in Figure 1.Between 2007 and 2012 the process was developed and refined at Direct Nickel’s laboratory within the CSIRO Minerals’ facilities in Perth, Australia, and in 2013 a series of pilot plant trials were run under continuous operating conditions. The process is designed to treat both limonite and saprolite in the same flow sheet, and after comminution, the blended ore is mixed with nitric acid and leached just below boiling point in agitated atmospheric leach tanks. The acid is added in excess, so typically for a blend of limonite and saprolite the acid addition can be in the order of 1.8-2.0 t acid/t dry ore. After a residence time of between 4-6 hours the majority of the soluble metals have leached into solution and the insoluble leach residue is separated from the pregnant leach solution (PLS) via counter-current decantation."

Jan 1, 2016

By D. J. MacKinnon

"The Canadian Copper Company, formed in 1885, began mining the nickel-copper ores of Sudbury in 1887 and producing nickel-copper matte from its smelter in 1888(1). Although it was quickly appreciated that the separation of nickel and copper (i.e. refining) was critical to the production of a marketable nickel product, refined nickel was not produced in Canada until about 1918. From the beginning of the industry, there was considerable political agitation to refine nickel in the Province of Ontario(2). The companies, however, claimed that the existing processes, viz. the sodium sulphide ""tops and bottoms"" process operated by the Orford Copper Company of New Jersey, U.S.A.(1,3,4) and the carbonyl process(5,6) developed by the Mond Company at Clydach, Wales, were too costly to operate in Canada. Nickel was first refined in Ontario about 1918 when the International Nickel Company opened a nickel refinery at Port Colborne, Ontario(6). Initially, however, the refinery at Port Colborne employed the Orford sodium sulphide process, and the first electrorefined nickel in Canada was produced in 1920 by the British America Nickel Corporation whose mine and smelter were located in Sudbury, Ontario, but whose electrorefinery was located at Deschenes, Quebec, near Ottawa, Ontario. The purpose of this article is to outline the early history of nickel electrorefining and to trace the significant developments which led to the first successful commercial electrolytic nickel refinery in Canada. The development of an electrolytic process for refining nickel is due largely to the efforts of four individuals, viz. Carl Hoepfner, Hans A. Frasch, D.H. Browne and ultimately N.V. Hybinette, and the contribution of each is discussed separately below."

Jan 1, 1999