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By P. Espinoza, I. Yutronic, J. Olivares

"Compañía Minera Doña Inés de Collahuasi's (CMDIC) Concentrator Plant is located at 4300 meters above sea level, approximately 2000 kms north from Santiago, Chile. The concentrator has three grinding lines. Each one, Lines 1 and 2, have one SAG mill at the primary grinding stage followed by a secondary grinding and classification stage consisting of one ball mill, two hydrocyclone batteries and a sump with two pumps. Line 3 has one SAG mill (Gearless Mill Drive), followed by a secondary grinding and classification consisting two ball mills, four hydrocyclone batteries and a sump with four pumps. Average throughput for these three lines is 150 Ktons per day. Peak values of 180 ktons per day have been achieved.Since late 2008 and throughout 2009 and 2010, Operations and Automation areas developed an optimization program for the grinding process, considering the implementation of Model Predictive Controllers (MPC) for the three SAG mills, and later, extending MPC to the secondary grinding and classification stages. The implementation of a MPC solution was based on stability considerations (high variability of the SAG process, high interaction between sump level control and classification efficiency, etc), requirements for an automated and consistent operation, the need to maximize throughput under timevarying scenarios and, in the long term, optimization of grinding process considering Specific Energy Consumption, among other parameters. MPC solutions had proven control robustness and strong optimizing capabilities in terms of helping in discovering new operational points and so achieving all of the aforementioned goals.On the first stage of the program, MPC implementation considered identification of dynamical process control models based mainly on historical data, in order to know the impact of a group of Manipulated Variables over Controlled Variables. Also, the impact of measured Disturbance Variables was modeled in order to take into account the high impact and interaction of these variables over the processes. At the end of this stage, 4 controllers were successfully operating for the 3 SAG mills and the secondary grinding and classification process. In response to challenges made by the working group, the second stage of the development of control applications in the grinding process required to achieve a design and implementation of a solution capable of governing the milling process, in addition to stabilization and optimization of both processes (SAG milling and secondary grinding), to integrate from a point of view of process control and automation. Therefore, in the second stage, MPC performance was continuously monitored and optimized through advanced tuning, inclusion of more process models (with proper step tests), redefinition of specific optimization criteria and adjustment to new process needs.Finally, successful MPC implementation at Collahuasi has proved that fast development and deployment of process control solutions is possible and allows immediate process stability, higher throughput, specific energy consumption reduction, and successful product size control, among many others."

Jan 1, 2012

By T. C. Chudy, C. C. Hamilton, G. R. Lane

The selection and provision of mineralogical services to assist in developing a GeoMetallurgical program for a Cu Porphyry Mine in South America is covered in this paper and presentation. The objectives of the study were to provide a rapid mineralogical scoping protocol to support and confirm results of sequential Cu leach data and to provide detailed mineralogical analysis of four composites and products on which locked cycle test-work were performed. Results showed good reconciliation was achieved between measured chalcopyrite and secondary Cu mineral contents with leach data using a linear intercept mineralogical procedure by TIMA (Tescan Integrated Mineral Analyser), as well as a generally positive correlation between pyrite and Cu-sulphide grain size. Size-by-size analysis of Composite feeds and test products was used to outline comparisons between feeds on a mineralogical basis and provide liberation and association data with which to diagnose and better explain performance. Identifying key mineralogical drivers to Cu recovery and concentrate grades using these results, a greater understanding of the role of chalcocite and pyrite on metallurgical performance was established. INTRODUCTION As a global supply of copper, porphyry copper deposits are typically very large, often in excess of a billion tons of ore (Keeney, Walters, & Kojovic, 2011; McMillan & Panteleyev, 1980) with grades typically between 0.2 and 1.2% Cu, and up to 2% Cu in places. They are also a source of gold. The complexity of copper porphyries arises from their classically zoned nature, the type of host rock, as well as whether or not they have been overprinted by oxidation, weathering and supergene enrichment (Sillitoe, 2011). Zones of alteration and other geological attributes give rise to varied mineralogy, ore mineral textures and material properties, all of which result in different processing characteristics. Due to the interdependencies of geological and processing characteristics, it is vital to study links between the geology at the macro-scale and mineralogical features down to millimeter and micrometer scale (mineral content and texture) and their influences on mineral processing behaviour (Cropp, Goodall, & Bradshaw, 2013). The growing field of GeoMetallurgy aims to provide an Orebody Knowledge framework by which to gather such information as it pertains to: (1) Understanding and predicting likely flotation response; (2) Guiding relevant test work to investigate and validate links, interdependencies, theories and data interpretation/evaluation; (3) Populate a block model with relevant information and; (4) Implement and refine the model for ongoing optimization and control. Specific questions addressed by using mineralogy often include flotation diagnostics as aids to interpretation, causes of dilution and losses to tailings, as well as guidance to sampling for test work (Cropp et al., 2013).

Jan 1, 2019

By Wayne F. Luinstra, Frédérique Vincent, André R. Laplante

"A test to determine the amount of gravity recoverable gold (GRG) in an ore is described. Typical GRG results are given; possible uses are discussed. A case study of how the technology was used at Hemlo Gold Mines is presented.IntroductionUsing gravity to supplement either flotation or cyanidation is a well established practice in the gold industry. It differs from most recovery circuits in that most of the gold recovered by gravity would be recovered in any case by the circuit downstream, be it flotation or cyanidation. The economic justification of gravity is therefore based on small margins (for example. a net smelter return of gravity gold of 99%, as opposed to 94% by flotation). It was easily demonstrated when either flotation or cyanidation were relatively inefficient processes, and labour costs low (as gravity can be labour intensive). Over the past thirty years, the advent of better flotation machines (flash, column, high capacity), more effective collectors, and better control systems has increased flotation's metallurgical performance, thereby decreasing the incentive for gravity recovery. Cyanidation technology has undergone similar changes, with the advent of activated carbon, oxygen and lead nitrate addition, and improved impeller design.Today, gravity could remain an attractive option only inasmuch as it can be implemented with very low capital and operating costs. This has resulted in a relative shift away from gold gravity recovery (except for alluvial deposits), in the seventies and eighties. For example, as of the early nineties, gold gravity recovery has disappeared from the typical South African flow sheet. The advent of the Knelson Concentrator, at the beginning of the eighties, foreshadowed a resurgence of gravity recovery, as gold's grinding and classification behaviour (Banisi, Laplante and Marois, 1991) makes it possible to achieve adequate gold recoveries with very simple, Knelson based, gravity circuits (Laplante et al., 1994).Consider for example the Hemlo recovery circuit, made of a single 61 cm x 122 cm (2' x 4') screen, an automated 76 cm (30"") Knelson feeding a Gemeni table, to produce a 70-80% Au concentrate which accounts for approximately one fourth of the gold production (about 3 million grams per year). The low capital and operating costs make it possible to justify the gravity circuit on the basis of savings in other operating costs (e.g. fewer carbon stripping and reactivation cycles). Any improvements in recovery, which would at Hemlo be difficult to measure, would be additional benefits. Such a circuit is rapidly becoming a standard in Canadian and Australian operations.The justification for installing a gravity circuit must be first based on an estimate of how much gold it will recover (irrespective of what the potential benefits are). This is a function of the nature of the gravity circuit to be installed; however, Knelson circuits, by their very simplicity and efficiency, limit the options. This leaves the amenability of the ore to gravity recovery as the single most important parameter. Various approaches can be used to characterize this response (Woodcock, 1993). The present work sought a route which would fulfil the following criteria:"

Jan 1, 1996

By L. Gawlik

Since the beginning of the 20th century, coal has been the world?s main source of primary energy. In the last few decades, consumption of coal has gradually decreased in some European countries and elsewhere due to the introduction of other energy sources. Regardless, industry continues to use solid fuels for energy production despite of their association with carbon dioxide emissions (about 7 billion Mg). For global development going forward, it is important to estimate what level of energy consumption can be expected as far into the future as the next few decades. To answer this question, specialized global institutions such as the International Energy Agency, the World Energy Council, the World Coal Institute, the U.S. Energy Information Administration, the International Institute for Applied Systems Analysis, as well as oil companies (BP, Shell), conduct periodic forecasts of world energy consumption. The development of such forecasts is very complex because their reliability is influenced by many factors that are difficult to estimate, such as economic development, population growth, climate change, and public awareness. We are all aware that the global demand for electricity will rise significantly in the coming years as a result of anticipated population growth, industrial development, rising living standards, and the general development of civilization. The world?s mineable resources in 2010 were as follows: coal 665.6 million Mg, brown coal 195.4 billion Mg, oil 163 million Mg, and natural gas 185.5 billion m3. The annual global primary energy demand in 2008 was more than 12 billion toe (12,267 Mtoe). In 2006, electricity production in the world amounted to 18,921 TW?h. The structure of the primary energy mix in TW?h (2006) was as follows: coal 41.0%; natural gas 20.1%; nuclear fuel 14.8%; hydro energy 16.0%; liquid fuels 5.8%; and other renewable energy sources 2.3%. The European Union (EU) is one of the largest coal consumers in the world. In 2010, the EU-27 produced 396.1 million Mg of brown coal, and 132.9 million Mg of hard coal; 188 million Mg of coal was imported from outside the EU. The largest hard coal producers in the EU-27 are Poland, Great Britain, Germany, the Czech Republic, and Spain; major brown coal producers are Germany, Greece, Poland, the Czech Republic, and Hungary. For many years, the Polish energy sector has been based on domestic fuel sources ? hard coal and brown coal. The characteristic feature of the Polish balance of primary energy resources is a very high share of coal ? much higher than the world average. Pursuing the decarbonisation of the Polish economy may jeopardize Poland?s energy security, leading to a significant increase in electricity prices, which would hamper the country?s economic development. Price volatility and supply uncertainty are now the dominant features of the international markets for oil and natural gas. The high share of these fuels in meeting energy needs is the reason for unrest in many countries, including those in the EU. Under these circumstances, coal may become the stabilizing element of EU energy security.

Aug 1, 2013

By The Mill Staff

Professor H.E.T. Haultain: Table 4, on page 131 of this paper, gives the assays of sized mill products. I am sure that there are many others besides myself who would be very interested to learn how the sizes 0-3 and 3-13 microns were obtained. * * * Mr. Del Davis, Mill Superintendent, Flin Flon Concentrator, on behalf of the authors of the paper, has kindly sent us the following information in reply to Professor Haultain's enquiry. Sub-Sieve Sizing by Sedimentation Procedure (As used by R. J. Digney & J. E. Ryan, 1940) Fractionation of Sub-Sieve Sizes General Method: A 1,050-gram sample of the material was pulped to 3 liters; 10 c.c. of sodium silicate and 0.2 grams of Na2C03 was added and thorough agitation given for two hours previous to settling. The material was assumed to consist wholly of particles of the same density as the most abundant particles. The time required for a particle of the upper limit of the lowest size-range to be separated, to fall a specified distance was calculated according to Stokes Law. The pulp was poured into the settling apparatus and, when the calculated time had elapsed, the suspended material was drained. The settled sediments and the process repeated until no more suspended material was obtained. The next larger size fraction was obtained in a similar manner, using the settled sediments which had been stripped of all smaller sizes as above. The fractions were arbitrarily set at 0-3, 3-13, 13-23, 23-33, and 33-43 microns (43 microns = 325 mesh).

Jan 1, 1946

By Mehmet F. Taner

"Abstract - "Les Mines Selbaie" is a large, generally low-grade, polymetallic (Zn, Cu, Ag, Au) "giant" volcanic-associated ore deposit in the Harricana-Turgeon belt, which is part of the northern Abitibi Subprovince of Quebec. Economically recoverable mineralization occurred in three zones: the A1 (open pit), A2, and B Zones (underground). The pre-mining resource was 54.9 Mt, including 8 Mt of sub-economic, locally silver-rich massive pyrite. In the Selbaie region, calc-alkaline andesitic volcanics form the "basement." The Brouillan Volcanic Complex formed within an islandarc setting on this regional basement. Evolution of this complex involved the formation of a felsic magma chamber and the development of a large caldera structure over the Brouillan batholith. Second- order linear and sub-parallel faults and rifts developed as a result of back-arc rifting/extension. The edges of these rifts were feeders for linear felsic volcanism (similar to fissural volcanism). During felsic volcanism, the feeder zones acted as conduits for hydrothermal fluids with associated pervasive potassic alteration. The alteration is characterized both vertically and laterally by well-defined quartz, sericite, K-feldspar and/or, locally, biotite halos, to produce a mushroom-shaped morphology. This K-feldspar alteration is not common in most other volcanic-associated ore deposits. Mafic dikes cut all of the geological units and appear to occupy earlier synvolcanic structures. Regional greenschist- assemblage metamorphism and pervasive deformation affect all units. Thrust faults moved part of the Brouillan batholith and basement units over the Mine Sequence. Although there is no relationship between the earlier mineralizing events and these late thrust faults, the thrust faults probably played an important role in location of ore lenses. Three types of mineralization are present: 1) copper-rich veins and hydraulic breccias; 2) disseminated and stringer, low-grade and high volume zinc-copper mineralization; and 3) finely laminated, massive, locally silver-rich pyrite mineralization within structural basins. At the bottom of these basins and at the contact of welded acid tuff (WAT) units, massive pyrite zones are locally enriched in zinc, and contain the possible vestiges of black smoker-like material. The ore metals were probably derived from leached basement andesitic volcanics in a hydrothermal system powered by one or more heat sources within the Brouillan Volcanic Complex. The Selbaie deposits are mainly characterized by: (1) the presence of small volcano-tectonic subsidence basins, which host chemical sediments and volcaniclastic materials; (2) pervasive potassic alteration; (3) locally high-grade silver mineralization (up to 1-2 kg/t Ag); and (4) the presence of random zones of Pb mineralization. The Selbaie deposits can be closely compared with the copper-rich Murgul deposits, which belong to the eastern Pontides metallogenetic province in Turkey. If conventional classification criteria for the volcanogenic massive sulfide deposits (e.g., Noranda, Mattabi types, and recent subsea floor deposits) are taken into account, the Selbaie deposits still exhibit many differences in terms of alteration assemblages, physical volcanology and some styles of mineralization (low-grade and large volume). Consequently, the "Selbaie" deposits belong to a new sub-type of volcanic-associated sulfide deposit, which may be used in future exploration programs."

Jan 1, 2000

By K. Negishi, H. Izui, Y. Komiya

"Aluminum has the disadvantage of low strength. In order to obtain the high strength composites, crystals were refined in the composites. Grain refinement was approached at two process points, generating new grains and grain growth. It is important for refinement how to control the grain growth. Spark plasma sintering (SPS) technique and TiB2powder addition prevent grain growth. To investigate effects of strain and additives to the alloy, aluminum powder was sintered at 70 MPa and 80 MPa, changing the content of TiB2additive powder by SPS method. As the results, the crystal grain size was finer in the composite sintered at higher sintering pressure and with larger amount of additive, though over 1.0mass% TiB2addition dissuades grain refining. These grain refinements improve hardness and tensile strength of sintered aluminum alloys.INTRODUCTIONAluminum matrix composites have low density, high specific strength, and low cost. Their applications are growing continuously in the field of automotive and aerospace industries. Recent studies to improve their good mechanical and tribological properties use different ceramic materials such as Al2O3(Das, Munroe, & Bandyopadhyay, 1996; Ashwatha & Anthony Xaviorb, 2014),SiC(Hassani, Bagherpour, & Qods, 2014), B4C(Ehsan, Masoud, Touradj, Amir Hossein, & Ali, 2015),TiC(Ehsan, Ali, Masoud, Kamyar, & Touradj, 2017), and TiB2(Fei & Zhong-Hua, 2013).On the other hand, grain refinement is also an important techniques for improving the properties of aluminum alloys. Finer grains improve the mechanical properties of the alloy. For cast and wrought alloys, addition of grain refiners in molten alloys accelerates generation of nucleus and suppresses grain growth. Al-Ti-B alloy is generally added as grain refiners in the molten aluminum alloys(Easton & Stjohn, 2001). Especially, TiB2has very high melting point, low density, and excellent hardness. Therefore, TiB2was selected as additions into aluminum alloys. Liquids reaction leads to high cost, oxidation, discoloration of molten alloy, and clustering of additive powder as compared with solid reaction such as sintering. However, sintering of aluminum alloy powder with ceramic powders is difficult for oxide layer surface. Recently, we take attention to spark plasma sintering (SPS) as advanced sintering process. SPS process can easily sinter a high quality specimen at a lower sintering temperature and in a shorter time than conventional sintering processes. By SPS process, we have sintered titanium powders that are as difficult to be sintered by any other processes as aluminum powders, and improved mechanical properties by adding hard ceramic powders TiC and TiB(Kamegawa, Izui, Komiya, Kobayashi, & Arimoto, 2015; Sampei, Izui, Komiya, Shigimura, & Suzuki, 2015). Furthermore, it is important to consider not only pure aluminum but alloys for solid-solution or precipitation strengthening. Actually, pure Al does not have enough strength compared with the alloys. Al-Cu alloy and Al-Mg alloy demonstrate adequate strength for industrial and commercial sectors. Al-Mg alloy retains good formability, and has high resistance to corrosion. It is therefore used for many applications. We must consider alloying processes. The most cost effective process is blended elemental technology where alloying elements are added to base powder as elemental or master alloy powder (Smugeresky & Dawson, 1981; Yasue, Radjai, Miwa, & Sakaguchi, 2003; Bolzoni, Esteban, Ruiz-Navas, & Gordo, 2012)."

Jan 1, 2018

By Vincent Jourdain

Located in the James Bay Territory of Quebec (figure 1), the Clearwater property is jointly owned by SOQUEM Inc. and Eastmain Resources Inc. Exploration on the property since the early 70's has resulted in the discovery of several gold showings. Exploration work by the joint venture since 1995 has led to the definition of the Eau Claire Zone containing an indicated resource of 540 000 tonnes grading 11.1 g/t Au (8.7 g/t Au cut at 34.28 g/t Au). Hosted by the archean Eastmain Greenstone Belt, the property geology is dominated by mafic volcanics containing thin horizons of felsic volcanics and/or sedimentary rocks (figure 2). The sequence is metamorphosed to the amphibolite facies and contains two small toalitic intrusions. Regionally, the structural trend is oriented E-W but in the area of the Eau Claire Zone a second trend oriented NE-SW is also present. The first exploration work on the property was conducted by SEREM and SDBJ from 1973 to 1976. It consisted of an airborne electromagnetic survey and a ground follow-up program on several grids. This program, comprising geological mapping, geophysics and drilling, led to the discovery of the Natel showing where a drill intersection returned 8.89 g/t Au over 2.65 m. From 1978 to 1985 the area was withdrawn from staking because of the hydroelectric projects of the James Bay. In 1985, Westmin Resources Limited and Eastmain Resources began an exploration program based on a helicopterborne electromagnetic survey. Once again, a ground follow-up program was conducted on several grids. This program included a soil geochemical survey that identified a strong (up to 10 g/t) gold anomaly. Drilling of the anomaly led to the discovery of the Eau Claire Zone where resources were estimated at 270 000 tonnes grading 6.1 g/t Au at the end of 1990. In 1995, SOQUEM Inc. took an option to gain a 50% interest on the property and obtained its interest in 1998. Since 1995, two principal mineralization zones were identified on the property. The "Rosemary Zone" consists of veinlets and disseminations of pyrite and chalcopyrite within a magnetite-rich amphibolite. The mineralized zone was identified by stripping within an area of 1000 by 500 metres. Drilling indicates copper grades up to 0.32 % Cu over 49.2 m. with locally grades of 0.75 % Cu over 5.9 m. The mineralised zone is associated with QFP dykes and corresponds to a strong magnetic anomaly and a strong copper and molybdenum soil anomaly. The "Eau Claire Zone" consists principally of a set of gold bearing quartz-tourmaline veins oriented E-W and dipping at 45-50??to the south (figure 3). More than ten different veins are recognised in the zone and up to eight veins were observed in a single trench (figure 4). The host rock between the veins shows a degree of deformation similar to the regional volcanosedimentary rocks. Each vein is hosted by a unit of actinolite-tourmaline-biotite schist up to one meter thick. The veins and their hosting schist clearly cross-cut the schistosity of the hosting lithologies and the contacts between amphibolites and felsic to intermediate rocks. Usually, the veins are less than a meter thick and are composed of quartz, tourmaline and calcite. The proportion of quartz and tourmaline varies from 0 to 80% within a single vein. The veins commonly show massive, banded or brecciated texture. Sulphides, mainly pyrite, pyrrhotite and chalcopyrite, generally form less than 5% of the vein material.

May 1, 2001

By R. D. H. Wigmore

"The title of this paper probably sounds quite simple and self-explanatory. The problem, however, was not a simple one, and proved somewhat baffling and expensive over a period of years. After outlining the problem, and the methods which were used to overcome it, we will present our solution, or perhaps it would be more accurate to say the partial solution which we reached.We have in the cyanide circuit at Upper Canada two Dorr Single Tray Thickeners with 30 ft. diameter wood stave tanks 12 ft 4 1/ 2 in. deep. In the years prior to 1948, when the tonnage was 250 to 280 tons per day, two stage thickening was practised. As the tonnage was increased to 450, 500, and finally 560-600 the thickeners were used in parallel.Some settling problems were encountered, and in those days the means to combat these problems were very few. The only remedy was to dump in pounds and pounds of lime. Then along came a few reagents, and the writer remembers one in particular, - 11 Carboxel11 (carboxymethlycellulose). When water or solution was added to this reagent, the result was a gooey mass somewhat resembling tapioca. One drawback was that it stayed in this state, and didn't go anywhere. A method was finally developed of containing the mass (or mess) on a fine screen, and dissolving some of it was a fine spray of solution. The only thing the writer recalls was that it was not a very effective settling agent.Shortly after this time, along came the new flocculents, first a few, and then many, and our settling problems were over.But when the settling problem was solved at Upper Canada, some other problems were creeping up on us."

Jan 1, 1965

By S. Sagar

In September 1996, Dofasco commissioned a 165 Mt twin shell EAF furnace and an associated ladle fleet. During and after this startup, refractory performance in the ladles was optimized on a holistic approach through several initiatives that increased ladle refractory retention times from 7000 to over 16000 mins (160 heats). This dropped the overall refractory costs by 40% while ensuring a consistent and predictable ladle performance. One initiative that drove enhancements was the proper brick selection for each particular refractory area and it's process conditions. The work concentrated on the selection of a resin bonded alumina magnesia carbon (AMC) brick in the barrel area of the ladle, upgrades to the 2 piece monolithic alumina cast ladle bottom and increasing the slagline refractory brick quality with the appropriate zoning/paneling. The adoption of a magnesia alumina carbon (MAC) brick in the transition area of the barrel/slagline has also improved performance and predictability. The design developments have lowered the overall wear rates and balanced the service life of the entire ladle. Safety was improved and breakout potential reduced by changing the safety lining from a high alumina brick to a state of art technology - monolithic shotcrete lining. This initiative decreased the overall manpower required to brick a ladle and increased shop ladle availability. The ladle bricking area environment was also improved by adopting a post combustion device to burn phenol fumes from ladle preheating practices. This has improved the overall health & safety of the reline teams. The consistent and predictable performance of the ladle on a whole has allowed the shop to decrease the number of working ladles in operation from 5 to 4. This further decreased wear rates from over-oxidation and excessive thermal cycling, saving approximately $500,000 in natural gas from pre-heater operations.

Jan 1, 2004

By H. Matsuoka, M. Kawata, K. Mitsuhashi, C. Tokoro, T. Kato

Copper sulfide minerals such as chalcocite, bornite, and chalcopyrite exhibit high floatability, thus they can be recovered as a froth product by adding only a collector in flotation process. However, when their surface is oxidized, the floatability decreases and the simple addition of collectors is no longer effective. In this case, flotation activators are required to improve the floatability. Sodium hydrosulfide (NaHS) acts as an activator for chalcocite and bornite by re-sulfurizing their oxidized surface portions. Nevertheless, the flotation conditions still require optimization in order to achieve a satisfactory recovery of copper minerals. This work addresses the study of the effect of pH and NaHS dosage on the flotation of a complex Chilean copper ore containing chalcocite, bornite, and chalcopyrite. Flotation experiments were conducted using a Denver-type flotation cell and by investigating the effect of two NaHS dosages (50, 100 g/t) and three pH values (8, 9, 10). Following this process, the surface interaction between the minerals and NaHS were analyzed using zeta potential measurements. INTRODUCTION In recent years in the mining area of metal resources, the complexity of ores and the fineness of grind required to liberate minerals have become serious problems. When ores are complex, several valuable minerals may interact with each other and reactions may become more complicated compared with an ore containing only one valuable mineral. Therefore, more processing steps may be required than for a simple ore. When the valuable minerals exist as fine grains, it is difficult to liberate them and more grinding time is required. Furthermore, the probability of collision of fine particles with bubbles in the flotation process declines and it takes much time to recover them. Flotation is one of the most applied methods to separate valuable minerals from unwanted ones using the difference of hydrophobicity (or hydrophilicity). In the flotation method, the air is fed to a cell filled with the pulp of the mineral of interest and agitation is implemented to introduce air and generate bubbles. The hydrophobic particles attach themselves to the bubbles and float along with the bubbles while the hydrophilic particles remain in the pulp. Copper sulfide minerals such as chalcocite (Cu2S), bornite (Cu5FeS4), chalcopyrite (CuFeS2) are widely known abundant copper resources. Generally, flotation is used to recover and concentrate these minerals. Copper sulfide minerals usually have high floatability. However when their surfaces are oxidized, their floatability becomes lower. When the sulfide surfaces are oxidized, activators are added before flotation to modify the surface property of the Copper sulfide minerals so that they can be floated. Sodium hydrosulfide, NaHS, is known as one of the activators. Zhao (2018) reported that NaHS is often used as a sulfurizer for oxidized sulfide minerals (Becker (2014), Clark (2000), Newell (2006), and Newell (2007)). In earlier work, the authors performed flotation tests on a sample of Chilean copper ore which contained chalcocite, bornite, and chalcopyrite using a Denver-type flotation apparatus. The sample was ground to P80 70 µm with a rod mill and ball mill. In these tests, the following conditions were used: natural pH (about pH 8.5), flotation time of 45 min, 60 g/t AP3477 (sodium diisobutyl dithiophosphate) as collector, and MIBC (methyl isobutyl carbinol) at appropriate dosage as frother. NaHS was applied at the dosage of 0, 67, and 120 g/t, respectively at the conditioning. The concentrates were collected at 6, 18, 30, and 45 min. The weights of the concentrates and tailing were measured and they were assayed for copper by ICP-AES (Optima 5300 DV, PerkinElmer Inc.). The flotation recoveries of the copper minerals were calculated from the results of MLA (mineral liberation analyser) (MLA 650, FEI) measurements.

Jan 1, 2019

By J. L. Jambor, E. J. Reardon, E. C. Appleyard, J. A. Cherry

Since the 1957 initiation of deposition of mill tailings at the Heath Steele mine, New Brunswick, periodic detailed studies of the tailings geochemistry and mineralogy have been conducted, thereby providing a unique opportunity to examine the geochemical evolution of a sulfide rich mine>tailings impoundment. When deposited, the tailings contained about 85 wt% sulfide minerals, 1 wt% to 3 wt% carbonate minerals, and 12 wt% to 14 wt% alumino silicate minerals. Mineralogical examination of samples of the tailings from deep in the impoundment indicates that little oxidation occurred prior to tailings deposition. The mill process-water deposited with the tailings was neutral to slightly basic and contained low concentrations of Cu, Pb, Zn and As. Sampling of the tailings five years after deposition indicated that sulfide-oxidation reactions, occurring near the tailings surface, had decreased the tailings-water pH and increased the concentrations of Fe, S042-, Cu, Zn and Pb (up to 10 000 mglL, 25 000 mg/L, 2.5 mg/L, 200 mg/L, and 4.5 mg/L, respectively). In 1975, ten years after decommissioning, the tailings could be divided into three zones: an upper oxidation zone in which sulfide oxidation is actively occurring, a hardpan layer, 15 cm to 20 cm thick, composed of tailings cemented by secondary Fe minerals, and an underlying reduction zone in which the tailings are relatively tin oxidized (Boorman and Watson, 1976). Chemical analyses indicated the presence of low-pH waters containing very high concentrations of SOi-, Fe, and other metals (up to 63 600 mg/L SO42-, 28 600 mg/L Fe, 10 mg/L Pb, 600 mg/L Cu and 4000 mg/L Zn), suggesting that there was a period of intense sulfide oxidation between five and ten years after tailings deposition. More recent sampling, conducted some 20 years after the end of tailings deposition, indicates that relatively little change has occurred in the last ten years, thereby suggesting that the hardpan layer has provided an impediment to the downward movement of dissolved constituents. Analyses of the tailings solids show increased accumulations of many metals at or near the depth of the hardpan layer. Evaluations of the long-term evolution of the tailings can be enhanced by comparison of descriptions of the tailings with descriptions of the gossan that overlay the massivesulfide orebody prior to mining. At Heath Steele, this comparison suggests that many of the mechanisms controlling the distribution of metals in the tailings will probably persist over geologic time.

Jan 1, 1992

By Julian Boldy

"A review of lI5 significant volcanogenic massive sulphide discoveries in the Canadian Precambrian during the past 65 years reveals that only 7% resulted from surface exploration methods aimed at depths below the effective reach of current geotechnical systems. This has been due to the relative ease of shallow discovery, resulting in the harvesting of ore deposits within 100 m of surface. This surficial layer has been more or less effectively supersaturated by routine exploration. In view of the fact that a mining district's metal content is not likely to be confined to a surficial layer, the time has come to explore the vertical dimension, prospecting for ore at depths in excess of 150 m in productive districts.Essential to this task is an understanding of the Precambrian volcanogenic ore deposit model. The deposits generally have .a tendency to be associated with the terminal phases of felsic volcanism, and to be located along certain time-stratigraphic horizons, marked by the products of volcanic exhalation. The significant massive sulphide deposits are discrete entities, ranging from the top 5% of deposits greater than 45 million tonnes in size, with a surface area of 0.30 km2, to median-sized deposits greater than 1.2 million tonnes in size with a surface area of 0.02 km2. These significant economic ore deposits range in value from $5.5 billion for the top 5% of deposits to $125 million for median-sized deposits. In terms of their frequency of occurrence, the former group has a 4% distribution and the median-sized deposits have a 15% distribution.An historical review of many of the principal Precambrian mining districts, such as Noranda and Flin Flon, shows that major tonnages are usually discovered during the initial phase of application of each new geotechnical search system, until carried to its ultimate limit of effective depth penetration. With the exception of the Noranda district, the mean depth of ore discovery in the other districts is still generally confined to within 50 m of surface.Targeting deep ore is difficult. One effective prospecting method is the detection of fracture-controlled mercury leakage haloes in outcrop (or core) above and peripheral to blind deposits. This technique has proved to be a useful pathfinder, and focuses attention on deposits buried at depths in excess of 700 m in some instances. Down-hole EM systems also promise to be effective in locating deposits previously bypassed by drilling. In addition, trend surface analysis programs of trace element geochemistry should be used as an aid to targeting ore by directing exploratory drilling along the plane of ore-hosting stratigraphy.By definition, mining districts have a proven ""favourability"" for the occurrence of ore. Discovery probabilities in these districts will be considerably enhanced if future search efforts are concentrated below a depth of 150 m, and more use is made of ""deep"" ore search systems coupled with sound geological reasoning. Deep drilling programs to a depth of800 m should initially be designed to search for the top 25% of deposits greater than 3.3 million tonnes in size, with a surface area of0.05 kmt. :a value of$310 million and a 12% frequency of occurrence; This can be accomplished for an over-all cost approximating $40,000 per drill hole. If the search depth is limited to 400 m, the over-all cost of each drill hole is $20,000, equivalent to expenditures routinely allocated to grassroots exploration drilling in the remote greenstone belts of the Precambrian. Depending on a company's objective, prospecting for deep ore offers a stimulating challange - though risky, it has its rewards."

Jan 1, 1981

By D. E. M. Plaumann

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

Jan 1, 1988

By Bert Taylor

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

Jan 1, 1949

By M. N. Pons, E. Benizri, B. Laubie, A. Bani, Université de Lorraine LRGP, P. Kidd, A. J. M. Baker

"Serpentine (i.e. ultramafic) outcrops in Europe cover more than 10,000 km2 and have a low-fertility and low-productivity, making them unattractive for traditional agriculture. In general, these areas are slowly abandoned by local farmers, affected by strong rural exodus and landscapes undergo a process of closure. However, ultramafic landscapes have the potential to provide multiple ecosystem services that can contribute to Europe’s goals towards insuring food security, production of renewable raw materials and renewable energy. Phytomining (agromining) aims at cultivating metal hyperaccumulator plants that are able to accumulate Ni and other strategic elements from metal-rich soils and transport them to the shoots (>1%), which can then be harvested as a bio-ore. Such metal-rich biomass can be used to recover highly valuable metals, e.g. nickel (Ni) and also to produce energy (heat, electricity). Nickel agromining can thus offer an eco-efficient alternative to classical pyro- or hydrometallurgical processes without disturbing the soil cover and the geology from soils and rocks that would be considered as low grade ores or sterile material for conventional mining methods.Phytomining agroecosystems can lead to better soil resource efficiency and can offer a fully integrated, new agromining agriculture that could cover thousands of km2 in Europe, benefit local communities with a sustainable rural development and provide a wide array of ecosystem services which will need to be optimised under a life cycle assessment methodology:1. Production of bio-sourced metal products2. Production of energy biomass3. Restoration of soil fertility for conventional agriculture (Ni stress)4. Limiting Ni intake by local populations and farmers5. Use and conservation of local biodiversity (plant and soil fauna and microflora) as cultivated ecosystems6. Store carbon in cultivated soils7. Enhance pollination activity at landscape level.Nickel agromining could therefore change the economical balance in historically-disadvantaged ultramaficregions of Europe"

Jan 1, 2016

By I. G. L. Sinclair

"ABSTRACTTill geochemistry was the principal exploration tool employed by Selco in the investigation of a zinc prospect in Lanark Township, Ontario. The geochemical studies were undertaken after the discovery at surface of marble boulders containing bands of sphalerite and lesser amounts of galena.The source of these mineralized boulders was sought by means of B-horizon soil geochemistry, followed, where appropriate, by back-hoe trenching and till sampling. The thickness of the tiff in this part of Ontario seldom exceeds 3 m or 4 m and it was generally possible to sample complete profiles of the till and thus to obtain detailed information concerning the dispersion of zinc. Drilling and trenching at the up-ice extremity of the zinc dispersion trains showed the source of the float to be sub-economic bands of sphalerite in the Grenville marble. IntroductionTill geochemistry was used extensively by Selco during the period 1975 through 1980. It was the principal exploration technique employed in a program seeking carbonate-hosted lead-zinc deposits of the Balmat-Edwards type in the Grenville rocks of southeastern Ontario. This paper describes the use of the method to locate the bedrock source of a group of sphalerite boulders found at the surface near the village of Hopetown in Lanark Township (Fig. I) . Geochemical investigations at this locality were the subject of a previous publication by the author (Sinclair, 1979).All of the analytical work related to this project was done by Bondar-Clegg & Co. Ltd. in their Ottawa laboratory. In all cases the -80 mesh fraction of the till was analyzed. The total zinc content was determined by atomic ab sorption spectrometry after hot aqua regia digestion; readily extractable zinc was determined by atomic absorption after samples had been cold-leached for 18 hours in an ammonium citrate hydroxylamine hydrochloride solution, the pH of which was adjusted to 8.5"

Jan 1, 1986

By R. V. Typliski, G. J. Labarre

"Hudson Bay Mining and Smelting Co., Limited, Flin Flon, Manitoba, has developed a metallurgical wastewater treatment system employing two-stage lime neutralization, sludge recirculation and conventional mechanical thickening.Laboratory investigations, initiated in 1970, led to the construction of a pilot-scale minewater treatment plant, at Schist Lake Mine, in early 1971. This plant, treating 600 cubic metres /day of wastewater containing 30-60 mg/1 zinc. demonstrated metal extraction efficiencies of 99% under normal operating conditions.On the basis of this pilot work, a second pilot-scale treatment system was established in 1975 at the Flin Flon metallurgical plant to treat zinc refinery process waste water. Through numerous improvements, including automatic con¬trol equipment, the system has evolved to a full-scale operation capable of treating 800 cubic metres/day of wastewater containing 4-11 gm/1 zinc. Treated effluent contains 0.2-20 mg/1 zinc. The hydroxide sludge produced returns to the metallurgical process for the recovery of metals while clarified effluent discharges to waste.With the system, Hudson Bay Mining has achieved its original goal: the reduction of environmental contamination, while realizing a return from metal recovery.IntroductionWith the goal of reducing environmental contamination, Hudson Bay Mining and Smelting began laboratory testwork, during the late 1960's, into the methods and technology for neutralizing and removing heavy metals from acid mine drainage. Based on lime treatment, early tests showed much promise and culminated in the construction of a pilot-scale test plant at Schist Lake Mine in early 1971.This pilot system utilized technology developed in laboratory studies, and related to work then underway by state authorities in Pennsylvania coal mines. The basic operation consisted of a lime slurry tank, baffle reactor and a free-settling basin. Through three summers of operation, the Schist plant, capable of treating 600 cubic metres/day of minewater, produced a heavy-metal extraction of 99% under normal conditions. Although successful in proving the treatment chemistry, the plant displayed deficiencies in sludge densification and handling."

Jan 1, 1980

By D. T. A. Symons, M. M. Kapnistos

"Gravity profiles were established over five known Middle Silurian pinnacle reef s in Lambton County, southwestern Ontario. The profiles were approximately 3 km long, with stations every 60 m. Four fairly recent technological developments were used which now make the gravity method relatively efficient and inexpensive compared to the techniques available two decades ago. These are: (1) a La Coste-Romberg model G gravimeter to measure the gravitational attraction with a sensitivity of 0.01 mgal; (2) a Sokkisha model B-2 semiautomatic engineers level to measure the station elevations with an accuracy of 0.5 cm; (3) a Huntec FS-3 hammer reflection seismograph to determine the overburden thickness to an accuracy of ± 7%; and (4) a modern microcomputer to process the data sets and mathematically filter the gravity values. Thus, two men can measure and analyze approximately 1 km of profile per day and obtain the residual gravity profile to an accuracy of ± 0. 05 mgal. Positive gravity anomalies of 0.17 to 0.72 mgal were observed along all five profiles that approximately delineate the dimensions of the known pinnacle reefs, which have an average height of 120 m, diameter of 1500 m and depth to top of 575 m. The relative ease of the gravity survey method makes it a cost-effective technique for hydrocarbon exploration in southwestern Ontario. IntroductionPinnacle reefs are important sources of oil and gas. Although the pinnacle reefs in southwestern Ontario are small, they may contain commercial pools of oil and gas because of their proximity to refineries. They are also useful for subsurface storage of gas to meet the winter surge in gas demand.At present, geophysical exploration for new pinnacle reefs is done primarily by relatively expensive seismic methods, which are also limited in southwestern Ontario by prohibitions on blasting. Gravity methods of exploration were attempted more than two decades ago in this region. (Pohly, 1954); however, the level of technology at that time made the method marginal in sensitivity and quite expensive. Several subsequent developments- gravimeters with greater sensitivity and operating ease, ""self-levelling"" levels for elevation control, hammer seismic reflection units for determining overburden thickness and portable microcomputers for efficient data analysis-have significantly improved the precision and cost effectiveness of the gravity method. Five known pinnacle reefs were selected in Lambton County, southwestern Ontario (Fig. 1), to test the effectiveness of integrating these developments to locate and define additional reefs. Isolated reefs were chosen to avoid the problem of the gravity effect s of one reef interfering with those of a neighbouring reef. The surveys were facilitated by choosing reefs with a concession road passing directly over them ."

Jan 1, 1980

By G. M. Proudfoot

IN June, 1940, a .Longyear 3420 blast-hole drill rig was purchased by the Hudson Bay Mining and Smelting Company for the purpose of experimental blast-hole drilling. The early work showed that the diamond drill could compete successfully with the Leyners then in use and by January 1st, 1941, the diamond drills were placed on contract. They may be considered to have gone into production at that date. Prior to the introduction of the diamond drill, our bench drilling was clone with Ingersoll-Rand S-70 and S-83 Leyner drills. Holes up to 40 feet were drilled, using coupled 1 1/4-inch round steel. Bits with a gauge of 3 inches were used in starting these holes, with a drop of 1/8-inch gauge for each 2-foot run. The holes normally ended with 1 1/2 or 1 5/8-inch gauge, though bits down to 1 1/8 inches were available for holes longer than normal or for holes drilled in abnormally abrasive ground. This method of drilling was very satisfactory and was a distinct improvement on the three-bench method of drilling used previously (1). During 1937 and 1938, when experienced miners were available, Leyner long-hole drilling was carried on with great efficiency. However, in 1939 and 1940, labour turnover in-creased and experienced long-hole drillers became scarcer. This resulted in a marked decline in the efficiency of the drilling operations, with increased costs and less tonnage drilled per shift. EXPERIMENTAL WORK The initial diamond drilling test work was carried on with a Longyear 3420 air-driven blast-hole drill rig, drilling a 1 3/16-inch hole. Tests indicated that diamond drilling could be successfully adapted to local conditions and that overall costs compared favourably with costs obtained with the Leyner long-hole method. Leyner long holes on stope benches usually bottomed at a larger diameter than the 1 3/16 -inch diamond-drill holes. More powder could therefore be introduced into the Leyner holes than could be placed in the diamond-drill holes. To achieve the same fragmentation, additional footage was required when diamond drilling. Diamond-drill holes of larger diameter were indicated.

Jan 1, 1943