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By M. Melashvili

"The SART Process was introduced to the gold mining industry in the late 1990s and several commercial plants were built over the following 10 to 15 years. Most of these plants have operated well and enjoyed plant performance in line with expectations based on the underlying chemistry of the process. However, two of the plants have consistently suffered from high reagent consumption and poor product quality, which has motivated a re-examination of the process fundamentals.The primary product in the SART Process is synthetic chalcocite, Cu2S, and the stoichiometric requirement for sulphide addition is therefore 0.5 moles per mole of copper in solution. It is shown here that when sulphide is added in excess of this amount (generally with the express purpose of lowering residual copper in solution) it not only fails in its intended purpose, but can have a negative impact on the settling and filtration characteristics of the copper sulphide product, as well as the final product quality. Moreover, these problems are exacerbated the higher the selected operating pH in the pH 3 to 5 range, and the longer the residence time of slurry in this pH range.Freshly precipitated copper sulphide compounds are known to be very fine and slimy, with poor settling and filtration characteristics. It is therefore important to engineer the SART Process in such a way as to encourage growth in the average particle size of the precipitate crystals. This was the rationale for recycling a portion of the thickener underflow to the primary reactor when the SART Process was developed. However, the theory of crystallization predicts that compounds of extremely low solubility, such as chalcocite (Cu2S: KSP = 10-48), are unlikely to experience growth in crystal particle size through heterogeneous nucleation and are more likely to grow through aggregation. On this basis, the SART practice of recycling a portion of thickener underflow to the primary reactor in the commercial plants is unlikely to lead to crystal growth via nucleation, and could in fact promote breakage of the aggregates that formed in the thickener. Recycling thickener underflow could therefore have the opposite of the desired outcome. The result of testwork that is reported here lends credence to this theory and questions the wisdom of recycling in the SART process."

Jan 1, 2016

By F. R. Jones, B. W. Taylor, A. R. MacPherson, L. J. Lamb

"The Lake St. Joseph Iron Deposit is in Northwestern Ontario. Exploration to date by Steep Rock Iron Mines Limited has indicated an ore reserve of 683 million long tons. Metallurgical investigations have proved that suitable oxide pellets and/or superconcentrates can be economically produced from this ore. Studies of the Lake St. Joseph iron ore deposits have covered the period 1956-71, and have embodied applications of the improvement in mining and processing technology over this period. Market demand in the early 1960's for high-grade oxide pellets as a replacement for natural iron ores for the blast furnace first drew attention to the potential of the Lake St. Joseph deposit. The current market trend toward utilization of sponge iron to replace steel scrap as feed to electric furnaces and cupolas now suggests that the deposit can better be exploited as a prime source of superconcentrates."

Jan 1, 1972

By S. R. McCutcheon, D. R. Lentz

"Abstract - The 12.1 Mt Brunswick No. 6 Zn-Pb-Cu-Ag massive sulfide deposit (New Brunswick, Canada) lies between two subaqueous felsic volcanic formations near the base of the Middle Ordovician Tetagouche Group. The footwall comprises rhyodacitic pyroclastic rocks (Nepisiguit Falls Formation), whereas the hanging wall comprises a sequence of rhyolitic flows, breccias, and hyaloclastites (Flat Landing Brook Formation), with an age difference of less than 3 m.y. The Brunswick No. 6 deposit is a proximal autochthonous deposit with a well-zoned massive sulfide body and a basal Cu zone that, at depth, develops into a stockwork stringer sulfide (feeder) system. The massive sulfides are capped sharply by a layered magnetite-chert unit that extends regionally beyond the deposit. The deposit has a keel shape (sheath) formed by the F1F2 interference pattern. The heterogeneous ductile deformation and upper greenschist-grade regional metamorphism has transposed many of the epigenetic stockwork structures and alteration within the host sequence. Hydrothermal alteration is much more extensive, both vertically and laterally, in the footwall (a few 100 m) than in the hanging wall (<100 m). Anomalies farther up in the hanging wall sequence seem to be associated with independent alteration systems related to the rhyolite domes. The least-altered footwall units exhibit keratophyric alteration (albite or adularia) and weak Mg enrichment (chlorite). More intense micaceous (sericite and Mg-rich chlorite) alteration occurs around the footwall sequence. The transposed stockwork stringer sulfides are typically composed of Fe-rich chlorite (± sericite, ± silica) with pyrite, pyrrhotite, chalcopyrite, arsenopyrite, and sphalerite. The (Fe2O3T+MgO)/(Na2O+K2O) and base metal alteration indices are the best practical lithogeochemical vectoring tools at this deposit. Sommaire - Le gîte de 12.1 Mt de sulfures massifs de Zn-Pb-Cu-Ag Brunswick No.6 est localisé entre deux formations volcaniques felsiques sous-marines situées près de la base du Groupe de Tetagouche d’âge Ordovicien Moyen. Le mur consiste en roches pyroclastiques rhyodacitiques (Formation de Nepisiguit Falls) tandis que le toit consiste en une séquence de coulées, de brèches et d’hyaloclastites de composition rhyolitique (Formation de Flat Landing Brook); la différence d’âge entre les deux est de moins de 3 m.y. Le gisement Brunswick No. 6 est un gîte autochtone proximal comprenant un corps de sulfures massifs clairement zoné ainsi qu’une zone cuprifère basale qui passe progressivement à un stockwork de filonnets de sulfure en profondeur (feeder). Les sulfures massifs sont abruptement surmontés d’une unité laminaire de chert et de magnétite dont l’extension par delà le gisement est régionale. Le gisement affecte la forme d’une quille (fourreau) résultant d’un patron d’interférence F1F2. Plusieurs des structures épigénétiques du stockwork et des patrons d’altération de l’encaissant ont été transposés par la déformation ductile hétérogène ainsi que par le métamorphisme régional au faciès des schistes verts. L’altération hydrothermale est beaucoup plus importante, tant verticalement qu’horizontalement, dans le mur (quelques 100 m) que dans le toit (<100 m). Les anomalies situées plus haut dans le mur semblent être reliées à des systèmes d’altération indépendants associés aux dômes de rhyolite. Les unités les moins altérées du mur une altération kératophyrique (en albite ou en adulaire) ainsi qu’un faible enrichissement en Mg (chlorite). Une altération en mica (séricite et chlorite magnésienne) plus intense est présente autour de la séquence du mur. Les sulfures du stockwork de filonnets transposés consistent typiquement en chlorite ferrifère (± séricite, ± silice) accompagnée de pyrite, de pyrrhotite, de chalcopyrite, d’arsénopyrite et de sphalérite. L’indice d’altération (Fe2O3T+MgO)/(Na2O+K2O) et l’indice des métaux usuels sont les outils d’exploration les plus appropriés dans ce gisement."

Jan 1, 2006

By Lawrence D. Meinert

Abstract-Most large skarn deposits are zoned in both space and time relative to associated intrusions. Zonation occurs on scales from kilometers to micrometers, and reflects infiltrative fluid flow, wallrock reaction, temperature variations, and fluid mixing. The most spectacular examples of skarn zonation usually occur at the skarn-marble contact, where transitions between monomineralic bands can be knife sharp. Other small-scale examples occur in zoned veins and individual mineral crystals. Although, visually striking and scientifically interesting, in mineral exploration these small-scale variations are less useful than deposit- or district-scale zonation. In most skarn systems there is a general zonation pattern of proximal garnet, distal pyroxene, and vesuvianite (or a pyroxenoid such as wollastonite, bustamite, or rhodonite) at the marble front. As well, individual skarn minerals may display systematic color or compositional variations within the larger zonation pattern. Such patterns are reviewed for 14 well-studied examples of Cu, W, Sn, Au, and Zn-Pb skarns. In addition, many deposits have endoskarn or other alteration of the associated intrusion, and recrystallization or other subtle changes have occurred in the surrounding wallrocks. Copper skarns, such as Mines Gaspé in Quebec and Big Gossan in Irian Jaya, have high ratios of garnet:pyroxene and are zoned outward from the intrusion, to garnet, to pyroxene, to massive-sulfide replacement and vein deposits. Garnets in Cu skarn are Fe-rich and change from dark red-brown near the intrusive contact to paler brown, green, or yellow in distal locations. Pyroxenes in Cu skarns are pale and diopsidic near the intrusion, and become darker and more Fe- and Mn-rich away from the intrusion. Tungsten skarns, such as Salau and Costabonne in France and Pine Creek and Garnet Dike in California, have intermediate ratios of garnet:pyroxene, are more extensive vertically and along strike than perpendicular to the intrusive contact, and have zonation patterns commonly complicated by overprinting of metamorphic lithologies. In W skarns, garnet is commonly subcalcic and the pyroxene is Fe-rich, reflecting particularly reducing wallrocks or great depth of formation. Tin skarns, such as Dachang in China and Moina in Australia, also can have subcalcic garnet and Fe-rich pyroxene, but this reduced mineral assemblage typically is due to an association with reduced S-type granites. Tin skarns differ from most other skarn types in having a late greisen stage that may replace earlier Sn-bearing calc-silicate minerals, thus liberating Sn to form cassiterite. Many high-grade Au skarns, such as Hedley in British Columbia and Fortitude in Nevada, have low ratios of garnet:pyroxene and are associated both with reduced plutons and reduced wallrocks. Goldrich zones occur in Fe-rich, pyroxene-dominant, distal skarn. Zn-Pb skarns, such as the Yeonhwa- Ulchin district in Korea and Groundhog in New Mexico, have low ratios of garnet:pyroxene and generally form distal to associated intrusions. These skarns also are zoned from proximal garnet to distal pyroxene and pyroxenoid (bustamite-rhodonite), with significant zones of massive sulfides within and beyond skarn. Manganese enrichment of most mineral phases, particularly pyroxene, is characteristic of distal zones. Fundamental controls on skarn zonation include temperature, depth of formation, composition and oxidation state of associated plutons and wallrocks, and tectonic setting. Most W skarns form at relatively great depth, 5 km to 20 km, with extensive high-temperature metamorphic and metasomatic mineral assemblages. In contrast, most other skarn types are relatively shallow, <10 km and mostly <5 km, with limited, lower temperature metamorphic aureoles. Differences in oxidation state correlate well with different skarn zonation patterns, particularly garnet:pyroxene ratios and compositions, and can be used in both classification of and exploration for s

Jan 1, 1997

By C. E. Rodgers

WHEN I was asked to prepare, for presentation at the 1938 Annual Meeting of the Institute, some data concerning low-discharge grinding at the Sylvanite mill, it seemed entirely reasonable to expect to have at my disposal the results of at least four full months&apos; operation after the last unit, as then proposed, was installed. Delays in delivery of equipment, nevertheless, shortened the interval, as will be shown herein. In submitting the data contained in this discussion, I wish distinctly to state that the intention is not to claim that final conclusions can be reached at this time concerning our operation. Test periods, in certain instances at least, have been too short for this purpose, although some, where so stated, have been extended through March and April. It is hoped, however, that an advantageous trend in practice can be shown and that the experience related may be of interest to others who are now operating, or who are considering the operation of, low-discharge grinding mills. In the plant under review, four mills are in use: one primary and three secondary. Present mill capacity (grinding, as shown by screen test, about 99 per cent through-200-mesh and approximately 88 to 89 per cent minus-325-mesh with infrasizer giving about 90 per cent of the latter) averages about 525 tons per day. A brief description of each unit will be given and comments will also be made relating to operation. This will be followed by certain test data from which, in conclusion, operating results are summarized.

Jan 1, 1938

By A. G. Roach

HISTORY AND LOCATION SEAL HARBOR GOLD MINES, LIMITED, was incorporated in January, 1934, to take over 116 acres of mining lands in the Stormont gold district, Guysboro county, Nova Scotia. The nearby ten-stamp Mulgrave mill was also purchased, and since that time 974 acres of adjoining properties have been secured from the Government by direct lease. The property is served by a motor road and is about two miles from the village of Goldboro, situated on the shore of Isaac&apos;s harbour. The bulk of ?the supplies are brought in by coastal steamer from Halifax, a distance of about 100 miles, since the nearest railway connections are at Antigonish, some 40 miles distant. EARLYY WORK The Seal Harbor properties include the old Partington and Beaver Hat mines. A small tonnage of ore was mined from each and treated on the properties in stamp mills, which were later torn. down . A survey of old records shows that this ore averaged about $8.00 gold per ton. Development work by the present Company was confined to the 100-foot level until February, 1935, but after that rime the incline shaft was deepened to 425 feet and the 300-foot level was explored east and west of the shaft. Six belts of ore were intersected to the west of the shaft. The ore from these belts was hand sorted and the ore portion was put through the 10-stamp Mulgrave mill to defray operating expenses and to serve as a sample of the deposits. The average recovery per ton from this sorted ore was $6.10. Samples of the ore were sent to the testing laboratories of the Mines Branch, at Ottawa, and of Canadian Industries, Limited, at Toronto. Recommendations from both of these authorities were for removal of the coarse gold by concentration and barrel amalgamation, followed by cyanidation of the residue.

Jan 1, 1937

By P. F. Berry

X-ray fluorescence analytical techniques are fairly well established in the mining and metallurgical industries. In recent years, instruments utilizing radioactive isotopes have been developed which, owing to their more compact size, greater reliability and independence of substantial power supplies, are often preferred to X-ray tube installations. Analyses can now readily be carried out in the field, on the factory floor and in otherwise difficult locations. This paper describes many examples of radioisotope X-ray fluorescence for the analysis of metallic minerals, finished metals and alloys and also for the continuous measurement of metallic coatings, such as tin and zinc, on steel. Radiation scatter techniques are also described for certain analyses and for thickness and density measurements.

Jan 1, 1967

By Emilie Williams

Introduction Mouska Mine, Doyon Division site Background and timeline Obstacles to standard paste delivery system Components of Mouska paste system Technical considerations Operational considerations Mouska Mine Underground gold mine 30 km east of Rouyn-Noranda, Quebec Part of Doyon Division, 100% owned by IAMGold Commercial production since 1991 Cumulative production over 850,000 oz Conventional mining until 2009 485 m shaft from surface 560 m internal shaft 500 tpd (ore and waste) Off-site milling Approximately 150 employees

Oct 1, 2010

By Daniel Hodouin, Claude Bazin

"This paper examines the economic advantages of adapting the production grade and recovery targets of a flotation circuit as a function of metal prices. The study is periormed using a steady state simulator for a Cu-Zn differential flotation circuit operating under the following strategies of 1) Maintaining constant operating targets, 2) Maximizing individual circuit Net Smelter Return and 3) Maximizing the global Cu-Zn circuit efficiency. Results are compared to assess whether or not one of the strategies produces benefits significantly higher than the others. A discussion of the tools required to achieve such a kind of control completes the paper.Paper not Available at Time of Publication"

Jan 1, 1997

By A. L. Mular

"IntroductionThe mining industry has made slow but steady progress toward the automation of mineral processing plants over the last 35 years. The stimulus to automate is associated with important economic and hidden benefits, so that today, the latest in instrumentation and technology is being employed.This chapter assesses the status of instrumentation justified economically, defines levels of control and certain fundamentals of feedback control loops, describes typical sensors employed in concentrator circuits, summarizes control strategies employed therein, and briefly highlights trends in integrated plant control.Current Status of AutomationAs early as 1948, the mining industry began to show interest in automation, probably because of its close association with the petroleum industry. By 1958, some regulatory control had been achieved in various sections of a concentrator, particularly in crushing circuits. Successes were attributed primarily to the reliability of sensors employed in the corresponding control scheme. By 1961, computers had appeared on the scene, initially, for data logging purposes.From 1958 to 1965, very few new control strategies had been implemented. Guidelines in vogue at this time appeared to be:- Install proven automation, unless a good instrumentation department exists at the concentrator.- Do not bite off too much instrumentation at one time.- Do not purchase on price alone; instrumentation must function in the concentrator environment.- Do not purchase automatic control equipment to be fashionable.Such guidelines were employed1 because of a lack of: (a) quantitative process knowledge, (b) reliable primary sensors, (c) initiative to try something new, (d) trained personnel, and (e) understanding that older concentrators could justify automation in many situations."

Jan 1, 1989

By Sarah Bélair, Denis Beltrami, Valérie Weigel

"Mabounié is a polymetallic deposit in Gabon containing niobium (1.2%), tantalum (0.03%), rare earth elements (REE) (1.4%) and uranium (0.03%) that are mostly carried by Pyrochlore minerals. Classical route for niobium recovery is a pyrometallurgical process that consists in pyrochlore flotation, dephosphorization of the niobium concentrate and aluminothermy to obtain ferro-niobium. Pyrochlore concentrates can also be treated by a hydrometallurgical process with a mixture of sulfuric and hydrofluoric acid leaching followed by solvent extraction to separate niobium and tantalum. However, the first process does not allow recovering REE and the second one is not environmentally friendly because of the use of HF.ERAMET is looking for new opportunities on strategic metals market and has recently developed a hydrometallurgical process for the recovery of all these metals based on a sulfuric acid treatment of the ore. After solubilisation of the valuable elements, niobium / tantalum are precipitated as a concentrate (12% w/w Nb and 0.3% w/w Ta) while REE (2.7 g L-1) and uranium (90 mg L-1) remain in solution. Nearly 90% of REE are recovered from the solution by conventional REE double salt precipitation [NaREE(SO4)2.xH2O) method but some REE still remain in solution as this method does not allow to quantitatively recover medium (Sm, Eu, Gd, Tb, Dy) and heavy (Ho, Er, Tm, Yb, Lu, Y) rare earth elements. From the different process routes studied (solvent extraction, precipitation, ionic exchange etc.), an innovative method has been developed and patented to recover REE residual concentration in solution. This is based on an aluminum removal step followed by precipitation of REE as REE phosphate concentrate (REEPO4) by phosphate addition into solution. The phosphate concentrate is then purified according to the conventional monazite treatment, i.e. caustic conversion and REE solubilisation in acidic media."

Jan 1, 2016

Project Details ? Located 420 km northeast of Yellowknife, Northwest Territories ? Jericho kimberlite discovered in 1995 ? Feasibility study showed +3,000,000 carats recovered from 2.6 million tonne reserve ? Nine year mine life ? Approx. 100 person workforce in full production ? First mine for Tahera, First new mine for Nunavut Completed Ambitious Construction Schedule During 2005 o Feb ? Mar/05 Winter road mobilization o Mar ? Dec/05 Pre-stripping, earthworks o Mar ? Apr/05 Camp facility, occupied May 3 o May ? Dec/05 Process plant o Jun ? Dec/05 Maint. shop, fuel storage, power plant o Jan ? Mar/06 Commissioning, start-up o April 2006* COMMERCIAL PRODUCTION

May 1, 2006

By S. McPhail, D. Richardson, G. J. Simandl, C. Akam, S. Paradis, M. Yakimoski, A. Teucher, Y. Cui, F. Ferri

Historically, extracting dissolved solids from formation water was economically motivated. Today, main drivers are environmental benefits and regulations, as large volumes of water needed for hydraulic fracturing require treatment before reuse or disposal. The concentration of dissolved solids in flowback water increases with each use and may require treatment prior to disposal. Element recovery from produced water is not profitable unless market conditions are favourable, recoverable elements are abundant, and economical recovery methods are established. Sale of recovered materials could offset recycling and disposal costs. This paper describes markets for Cl, Na, Ca, K, Sr, Mg, Br, Ba, I, B, and Li, and their concentrations in produced water from oil and gas wells in British Columbia. It recommends revising sampling protocol for produced water.

Jan 1, 2018

By Brad Ross

? Value in number of children ? Polygamy common ? Support of extended family and friends ? HIV/AIDS ? Not a culture of savings ? Minimal mechanical experience ? Supportive ? Unintended hindrance ? Mining claims ? Export Permits Lack of funding

May 1, 2011

By M. Girard, E. Bourricaudy, M. Ashbury, M. Canizares, A. Mezei

Liquid-solid separation and slurry handleability relate directly to commercial plant ramp-up success or lack thereof. The paper provides a well-proven approach for testing, scale-up and commercial application of difficult processes. Emphasis is placed on practical applications of rheological concepts developed both academically and empirically. The cause-effect relationship between the process chemistry and possible physical implications on the metallurgical plant operability is presented in a simple, direct form. Examples include recent results in areas of mineral processing and extractive metallurgy. Technical process risk factors and their mitigation are included in the discussion.

Jan 1, 2011

By Penny C. Stewart

Narrow vein underground stope design engineers in Australia have expressed significant interest in the requirements and methodology for developing site-specific stability charts. This is largely due to a concern that existing stope stability charts inadequately take into account factors influencing narrow vein stope stability, such as the effect of backfill abutments and the effect of in stope pillars. This paper presents conclusions from a comprehensive statistical examination of case history requirements for a site specific stability chart and critiques literature arguments for site-specific charts. A new statistical analysis technique and a large comprehensive stability database enabled the authors to quantify case history requirements and any site-specific effects the model exhibits. The analyses indicated that a reliable stable-failure boundary requires at least 150 case histories, of which a minimum of 10 percent should be unstable stope surfaces. Marginal site-specific effects were observed for the operating conditions captured within the database. The authors conclude that the apparent site-specific effects noted in previous literature are attributable to operating conditions inadequately represented in the database. Such operating conditions could induce erroneous stability predictions at any site, and are therefore not truly site-specific. Following from these conclusions, a strategy for taking into account operating conditions particularly relevant to narrow vein mining has been proposed. In particular, the development of a chart that accounts for the effect of backfill abutments on stope stability. Backfill abutment effects are particularly relevant to mechanised narrow vein mining methods such as modified Avoca. Collaboration from at least five sites is required to obtain access to relevant data and underground drives for geotechnical mapping. Preliminary estimations indicate that approximately three months fieldwork is required to meet case history requirements. CMS data in conjunction with 3-dimensional design software will be used to calculate equivalent linear overbreak/slough (ELOS). In contrast to stable, failure, major failure and caving categorical variables, ELOS is an objective and continuous stability variable and is therefore, more conducive to mathematical modelling.

Oct 1, 2001

By G. B. Harris

In recent years, increasing attention has been focused on the hydrometallurgical treatment of base metals feeds, especially nickel laterites and polymetallic sulphides. One approach that has received some attention is the use of high?concentration chloride solutions (brines), which are now showing some promise after initial setbacks. Iron is usually the major component in the base metal feed to such processes, and if physical means cannot be used to remove it, then the iron concentration in the pregnant leach solution is always significant. This paper presents some of the background theory on the hydrolysis of iron in high-strength (>3M) chloride brines, and discusses the roles of temperature, solution concentration and free acidity. It is shown that the system is complex, but that crystalline and readily-filterable hematite can be generated, and that the system has a low energy requirement.

Jan 1, 2006

By E. G. Charlesworth, M. J. Robertson

"Abstract - Gold mineralization in the Main Reef Complex at Sheba Gold Mine is hosted within shear zones in a sequence of shales and greywackes of the Fig Tree Group. Competency contrasts between these rocks and a steeply dipping chert unit provided a locus for the development of right lateral steeply dipping shear zones. A set of flatter dipping shear fractures crosscut these zones with their intersections defining a system of en échelon ore shoots.Mineralization is both of vein type and disseminated strata-controlled replacement type. The latter occurs mainly within the wall rocks to the shear zones. Arsenopyrite and pyrite are the major sulfides, with arsenopyrite being confined largely to greywacke beds and pyrite to the more iron-rich shale units. Gold was introduced, together with minor sphalerite, pyrrhotite, chalcopyrite and early arsenopyrite, toward the end stages of a first pyrite phase. The mineralizing fluids were introduced along dilated cleavage planes and, emanating from these, produced a series of replacement fronts. Mineralization was syntectonic to late tectonic and occurred under conditions of progressive deformation, resulting in sulfides which are concentrated in the hinges of shear-related folds and semi­ massive sulfides within tension fractures which are folded, attenuated and fractured .An analysis of the structures shows a sequence of events in which an initial ductile simple-shear event with a steep sense of movement was succeeded by a more brittle strike-slip event. Mineralization is temporally related to the latter. This evolution in structural style was probably related to reactivation of the Sheba Fault as a transcurrent fault during emplacement of large granitoid bodies."

Jan 1, 1994

By J. E. Armstrong

THE Pinchi Lake mercury belt is in central British Columbia, about fifty miles north of the Prince Rupert branch of the Canadian National railway (Figure 1). Until 1936 little was known about the geology of this area. Jn 19~6 and 1937, J. G. Gray and the writer mapped the Fort Fraser area (G.S.C. maps 630A and 631A) for the Geological Survey of Canada. During the field season of 1937, cinnabar was discovered by J. G. Gray in the Permian limestones on the north shore of Pinchi lake on what have since proved to be the main showings of the Pinchi Lake mercury mine. Claims were staked in the spring of 1938 and later that year were optioned to the Consolidated Mining and Smelting Company. Development work revealed cinnabar deposits of commercial size and grade, and a small reduction plant was erected, production of mercury commencing in June, 1940. Since 1940, this plant has been enlarged by additional units. The ore treated is reported to contain between 0.5 and 0.75 per cent mercury. The production of mercury from this deposit is now far in excess of Canadian requirements and this country is able to supply the United Kingdom with part of its needs. In 1940 and 1941, several other occurrences of cinnabar were found northwest of Pinchi lake. During 1941, the writer made a geological investigation of the known cinnabar occurrences and of the belt of country in which they are located. The present paper summarizes the results of this work. The writer is indebted to officials of the Consolidated Mining and Smelting Company, and to Mr. E. Bronlund of the Pinchi Lake mine in particular, for their generous co-operation in the work.

Jan 1, 1942

By Oliver C. Ralston

DURING the development of froth flotation for the treatment of sulphide ores, a certain amount of information was collected about a few of the nonmetallic minerals, and occasionally the gangue minerals in the sulphide ores showed an embarrassing tendency to float. However, for many years efforts were expended to develop reagents that would cause wetting and depression of these minerals. Among the nonmetallics, the minerals sulphur, graphite, and talc probably should be classified as &apos;natural floaters&apos;, and perhaps sericite. They respond to almost any type of flotation reagent, and their beneficiation by flotation had been attempted years ago. Flotation of sulphur is now common practice, and graphite was concentrated during the World War in many mills, most of which have now closed down. A talc mill has just been constructed, and another may soon be erected. Second to the &apos;natural floaters&apos; come the &apos;easy floaters&apos;, cassiterite, fluorite, barite, mica, colloidal gangue, and paper pulp from white water of paper mills. That these materials responded to some of the early flotation reagents, as oleic acid, has long been known; occasional references to their behaviour are found in the very early flotation literature and in that of today. In general, however, nonmetallic minerals, valued at $2.00 to $15.00 per ton, have for a long time been regarded as cheap products and not worth much beneficiation. Consequently, only high-grade natural deposits suitable for direct mining and sale have been gouged out. Furthermore, the mining of these minerals has too frequently been in the hands of &apos;super-practical&apos; men-men of no technical or other training, willing to work for a very low compensation. From the social standpoint, nonmetallics has been the field for small, untrained enterprisers with little money to spend on plant or equipment. But times have been changing. The impact of technically trained men on the old industry has revealed appalling losses and inefficien-cies and proved the value of processing materials discarded or avoided by the old regime.

Jan 1, 1937