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By Laszlo Imreh, Pierre Trudel, André C. Tessier

"The Siscoe gold mine is ranked as the fourth largest mine in the history of the Val d' Or mining district with a total production of 27.4 tAufrom1929 to 1949. Most of the gold mineralization occurred in quartz veins, hosted in a small (.69 km2) intrusion of altered quartz gabbro known as the Siscoe stock. The southern contact of the Siscoe stock is crosscut by a major shear zone (the ""K"" zone) trending 295/80 NE and from which gold mineralization was also exploited within the limits of the stock. Field and petrographic observations of the Siscoe stock combined with chemical analyses of major, trace and rare earth elements of the stock and its volcanic host rocks (Dubuisson Formation) suggest that:The Siscoe stock comprises four alteration facies formed from a homogeneous precursor: 1) the actinolite - Mg-chlorite facies, characterized by low FeOt,Na2O andTiO2 , and highAlp3 and MgO; 2) the albite- Fe-chlorite facies, containing high concentrations of FeOt and Nap; 3) the Mg + Fe-chloritefacies, which is transitional between the actinolite- Mg-chlorite and albite- Fe-chloritefacies; 4) the carbonate-chlorite facies, which consists of rocks that were subjected to extreme hydrothermal alteration, is spatially associated with mineralized veins and part of the K-zone. The first three facies are depleted in SiO 2, CaO and Au, and enriched in MgO, relative to fresh rocks of the Siscoe stock and the equivalent rocks of the Dubuisson Formation.Hydrothermal fluids introduced mainly along the K-zone ""channelway"" induced three types of element remobilization, resulting in the various alteration facies: 1) redistribution of elements within the stock from one facies to another (e .g., FeOt, Nap, Ti02 migrated from the actinolite- Mg-chlorite facies to the albite- F e-chlorite and Mg + Fe-chlorite facies); 2) remobilization of elements from the intrusive rocks into veins (e.g ., SiO 2 and CaO ); 3) introduction of elements by hydrothermal fluids in rocks of the Siscoe stock (e .g., Al2O3, MgO, H2O, CO2 and FeOt).The Siscoe stock and the Dubuisson Formation form part of the same tholeiitic suite. Similarly, the Bourlamaque batholith and the Val d'Or Formation belong to a single calc-alkaline suite. In view of this relationship, dissimilarities between the petrochemistry of the Siscoe stock and the Bourlamaque batholith suggest that the two intrusions are not co-magmatic.Finally, gold may have been leached from rocks of the Siscoe stock and remobilized into the quartz-carbonate mineralized veins. In order to account for the total production at the Siscoe mine, approximately 1.2 km3 of the Siscoe stock is thought to have been partially leached of its gold."

Jan 1, 1990

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 A. Y. Smith, R. J. Allan, E. H. W. Hornbrook, R. W. Boyle, W. Dyck

"HYDROGEOCHEMICAL METHODS in the Canadian Shield can be based on analyses of the following :(1) spring waters and their precipitates;(2) stream and river waters;(3) lake and muskeg waters;( 4) ground waters from drill holes, wells, faults, etc.;(5) waters issuing from diamond drill holes, faults, shear zones, etc. into mine workings and open pits; also analyses of precipitates associated With these waters.The first four can be utilized in both reconnaissance and detailed surveys. The last is suitable only for detailed work.For hydrogeochemical methods of prospecting to be effective, the following conditions must be present. 1. Natural waters must come into contact with mineral deposits and remain in contact with them long enough for the solution of detectible amounts of ore, gangue or indicator elements. 2. Dispersion patterns must be present in the hydrologic systems in order to trace the elements back to their source. These patterns may take the form of trains (e.g., increasing amounts of elements upstream toward t he source (s), Figure 1 ) or halos (e.g., increasing amounts of elements in bodies of water toward the source (s), F igures 2 and 3)."

Jan 1, 1971

By M. A. Tomlinson, A. Di Feo, T. C. Cheng

"Goldcorp is evaluating the potential for economic reprocessing of historic tailings at several sites, including the Aunor A facility located in Timmins. The overall project is investigating the ability to recover value, remove sulphide minerals and alternative tailings re-deposition strategies. Desulphurization of tailings will eliminate potential acid generation while improved tailings re-deposition could reduce long-term liabilities and closure costs. Through a task-share project between CanmetMINING and Goldcorp, tailings obtained from Aunor A site have been characterized using mineral association analysis and diagnostic separation techniques. The selected tailing samples contain about 2 g/t of gold. Mineralogical analysis has identified coarse agglomerates bound with gypsum and significant free gold, confirmed by high recoveries in cyanide leach tests. However, despite no evidence of surface passivation on the minerals, conventional flotation techniques were found to be ineffective in concentrating gold or sulphides due to the high fines content and highly agglomerated particles.INTRODUCTION Aunor Mine is one of the historic mines that belongs to Porcupine Gold Mines (PGM) in Timmins, Ontario. PGM has undertaken an extensive mine closure program to remediate the legacy and closed mines with the funding supported by operating mines since 2006. Over the course of 100 years of operation in the Porcupine camp, it is estimated (Stevens et al., 2013) that about 130 million tonnes of tailings have been generated. Aunor Mine produced 2.5 million troy ounces of gold between 1940 and 1984, while processing over 7.6 million tons of ore. The original gold grade of Aunor Mine ores was about 10 g/t. With 95% gold recovery at the time, an estimate of 0.5 g/t of gold remains in the Aunor Mine tailings. Aunor Mine has two tailings impoundment areas: designated as Aunor A and Aunor B. The two sites are similar in size, about 7 hectares each. The amount of tailings in the Aunor A site alone is about 3.2 million tonnes. In 2017, bulk tailings were sampled from five designated locations in Aunor A. They were characterized using mineral association analysis and diagnostic separation techniques. The bulk tailings samples contained about 10% total sulphur, 10% iron, 2% arsenic and 2 g/t of gold. The prime objective of this collaborative work between CanmetMINING and Goldcorp is to facilitate the timely development and deployment of innovative tailings reprocessing solutions, specifically in the areas of desulphurization, and recovery of metals. This work targeted specific tailings of interest to Goldcorp which goals are to improve its environmental footprint and to reduce long-term liability. Thus, it is imperative to look for innovative reprocessing options that can overcome not only technological challenges but also achieve economic benefits to offset the costs associated with reprocessing the tailings. Moreover, the greater goal is to demonstrate and deploy tailings reprocessing solutions to solve one of the industry’s most difficult challenges for a sustainable mining industry in Canada."

Jan 1, 2019

By P. B. Marchant

"Considerable gold and silver values are lost to flotation tailing from both the South ern Tail orebody and the Main Zone orebody,The underlying mineralogical assemblage of the two ore types is quite different. This paper describes the research strategy employed to determine an economical method to scavenge gold and silver from flotation tailing.Much of the unrecovered gold in the Southern Tail ore is associated with arsenopyrite while silver losses are probably as unliberated tetrahedrite and minor silver-bearing sulphosalts. Processing alternatives tested include bulk sulphide recovery by flotation and gravity, selective pyrite-arsenopyrite separation, cyanidation, and oxidation pretreatments such as pressure leaching and bioleaching to enhance cyanidation response. Gold losses in the Main Zone ore are more closely associated with silver-bearing minerals, pyrite, and siliceous gangue and consequently respond more readily to conventional tank cyanidation. The 5 300 tid carbon-in-leach circuit to treat Main Zone flotation tailing is described in detail.IntroductionEquity Silver Mines Limited's open-pit silver mine, a 70% owned subsidiary of Placer Development Limited, is located 37 km southeast of Houston in north central British Columbia. The deposit consists of two major ore zones, Main Zone and Southern Tail (Fig. I). These zones are within a sequence of Cretaceous volcanic and sedimentary rocks striking 015 degree s and dipping 45 degrees west. Feasibility ore reserves were 28 million tonnes grading 106 g A g/t, 0.96 g Au !t and 0.38% Cu a strip ratio of 2.2 to I. A detailed mineralogical description has been presented and a comparison of the two ore zones is shown in Table I."

Jan 1, 1986

"The United Nations, through the secretariat of the United Nations Economic Commission for Europe (UNECE) and its Ad Hoc Group of Experts on Harmonization of Fossil Energy and Mineral Resources Terminology, developed the Framework Classification.The UN Classification applies to fossil energy and mineral reserves and resources located on or below the Earth's surface. It aims to simplify reporting of energy and mineral reserves and resources in a globalized world and is intended to improve global communications regarding remaining quantities of non-renewable fossil energy and mineral resources. The UN Classification is not recognized as a Foreign Code under NI 43-101.SECTION I1.,The United Nations Framework Classification (UNFC-2009) applies to fossil energy and mineral reserves and resources located on or below the Earth' s surface. It serves the needs for classification at a global level (for energy and mineral supply studies), for governments (resources management and policy formulation), for industry (business process management) and for financial reporting.SECTION II2.,The UNFC is a generic system in which quantities are classified on the basis of the three fundamental criteria of economic and social viability (E), field project status and feasibility (F), and geological knowledge (G), using a numerical coding system. Combinations of these criteria create a three-dimensional system. Categories (e.g. El, E2, E3) and, in some cases, sub-categories (e.g. E1.1) are defined for each of the three criteria as set out and defined in Annexes I and II.3 The first set of categories (the E axis) designates the degree of favourability of social and economic conditions in establishing the commercial viability of the project, including consideration of market prices and relevant legal, regulatory, environmental and contractual conditions. The second set (the F axis) designates the maturity of studies and commitments necessary to implement mining plans or development projects. These extend from early exploration efforts before a deposit or accumulation has been confirmed to exist through to a project that is extracting and selling a commodity, and reflect standard value chain management principles. The third set of categories (the G axis) designates the level of confidence in the geological knowledge and potential recoverability of the quantities.4. The categories and sub-categories are the building blocks of the system, and are combined in the form of ""classes"". The UNFC can be visualised in three dimensions, as shown in Figure 1, or represented in a practical two-dimensional abbreviated version as shown in Figure 2."

Jan 1, 2003

By David E. Pearson

"Western Canadian coking coals are located in the Rocky Mountains arid foothills of British Columbia and Alberta. They are typical of non-marine coals, characterized by total sulphur contents of less than one per cent and with ash contents of 10-30 weight per cent. This mineral matter is composed principally of kaolinite and quartz.The petrographic compositions of most of the coking coals which constitute the 2.065 billion tonnes of British Columbia's measured reserves, and the compositions of the metallurgical coal currently exported from Alberta, show them to be rich in the inertinite macerals, macrinite, fusinite and semifusinite. When the mean maximum reflectance of vitrinite in oil/Romax) is used as the coal rank parameter, this high inertinite content is reflected in the lower volatile-matter yields, but the relatively inert character is also displayed by lower free swelling indices (PSI), lower maximum dilatations, lower maximum fluidities and lower coke strengths than other coals of the same rank, but richer in the reactive macerals, vitrinite, exinite and semifusinite. Despite these reduced values in the so-called caking tests, inertinite-rich coals produce strong coke.A system of classification of coking coals is introduced which is based on rank, as indicated by Ro max, and either petrographic composition, maximum dilatation, maximum fluidity or, to a lesser extent, PSI and volatile-matter yield. Six principal coal groups can be distinguished among coking coals using this versatile classification. Inertinite-rich coals are assigned to the Keystone Group (GI), Balmer Group (G3) or Moura Group (G4), each group being named for a coal typical of that population ofcoals. Reactive-rich coals are assigned to the Pittston Group (G2), Kellerman Group (G5) or Big Ben Group (G6). A multiple-regression analysis of coal quality versus price shows how the value of coking coal in the Japanese market appears to be related to reflectance and the free swelling index, and confirms the economic viability of the classification.British Columbia's measured reserves of coking coal contain representatives of four of these groups; Balmer Group (73%), Moura Group (/6%), Pittston Group (3.4%) and Kellerman Group (7.6%).This analysis of the quality of Western Canadian coking coal indicates that despite the need for large proportions of inertinite-rich coking coal in blends produced in the Japanese steel industry, there is nevertheless a demonstrated overabundance of Balmer Group and Moura Group coals available to the international coking coal market."

Jan 1, 1980

By Noko S. Phala

The energetics of CO and CO2 hydrogenation to methanol were studied on a gold surface via periodic and cluster DFT calculations. CO hydrogenation appears to be much easier than CO2 hydrogenation on an Au(111) surface. The use of a cluster model does not seem to improve the energetics of CO2 hydrogenation. It is concluded that an active support (e.g. ZnO) will be required to increase the productivity of a gold surface towards synthesis gas conversion. The energetics of CO and H2 adsorption are improved on the cluster model compared to a slab model of an Au(111) surface, suggesting that nanosized gold particles may be beneficial in CO hydrogenation.

Oct 1, 2003

By E. Ghali, N. Sorour, G. Houlachi

"The binary Fe53Mo47 and ternary Fe54Mo30P16 coatings have been electrodeposited from citrate-based electrolyte on mild steel (MS) substrates. The electrodeposition has been conducted for 6 hours at 20 mA cm-2 and 30°C. The electrochemical tests were carried out in chlorate solutions containing 300 g L-1 NaCl and 4 g L-1 K2Cr2O7 at pH 6.4 and 80°C. A number of surface characterization techniques such as SEM-EDS, XRD and EMP were used to study the microstructure of the alloys. The electrocatalytic activity of these coatings was investigated using steady-state polarization and galvanostatic technique. This study revealed that the Fe-Mo-P coating could be considered as a promising cathode material for the hydrogen evolution reaction (HER). The results showed an important decrease of hydrogen overpotential for Fe-Mo-P alloy as compared to that of Fe-Mo and especially MS which is a traditional cathode in chlorate industry. The HER overpotential for Fe54Mo30P16 at current density of 250 mA cm-2 was decreased by 30% in comparison to that of MS (1029 mV), and by 16.5% to that of Fe53Mo47 coating (857 mV).INTRODUCTIONSodium chlorate is used for manufacturing of several daily consumption products. This chemical is mainly employed for the production of chlorine dioxide (ClO2) that is a bleaching agent used in the pulp and paper industry. The ClO2 is also consumed for generation of perchlorates, chlorites and other chlorates, as an herbicide and in uranium processing (Cornell, 2002). Sodium chlorate is produced industrially via following overall formula in un-divided electrochemical cell (Nylén & Cornell, 2009):"

Jan 1, 2016

By R. Vracar, D. Sinadinovic, Ž. Kamberovic

The paper presents a study on the oxidation of polymetallic gold ore by oxygen in an autoclave. It presents optimum parameters of autoclave leaching of sulphide refractory ore in self-generated sulphuric acid solutions aiming to get products with favourable technological characteristics from which precious and other useful metals can be recovered. A high partial oxygen pressure (1.5 MPa to 2 MPa), high leaching temperature (190°C to 200°C), pulp density 80 g/L to 240 g/L, ore fraction size 76 µm and leaching time up to 4 hours yield good results in rich ores. The full extent of copper and zinc conversion as well as a high extent of iron conversion has been achieved, while lead, gold and silver and a minor part of iron remain as a solid residue among gangue minerals.

Jan 1, 2001

By R. Lasmanis

"Washington State's diverse geological setting is reflected by its several types of porphyry Cu-Mo deposits. All are associated with calc-alkaline intrusions, and they range in age from Jurassic to Miocene.The largest is the Mount Tolman Mo-Cu deposit located in the Omineca Belt. It contains 2.18 billion tonnes of material averaging 0.093% MoS2 and 0.09% Cu. There are intermediate-size Mesozoic deposits in the Intermontane Belt. The Cascade volcanic arc hosts small- to medium-Tertiary deposits. Among these, of particular interest is the Sunrise breccia pipe, which has been explored over a vertical distance of 1000 m. It contains reserves of 64.6 million tonnes averaging 0.319% Cu and 0.071% MoS2 with significant tungsten values. The Glacier Peak deposits have potential resources of 1.7 billion tonnes averaging 0.334% Cu and more than 0.015% MoS2 and also contain tungsten. The Margaret deposit can be classified as a porphyry Cu-Au type; it has reserves of 200 million tonnes averaging 0.434% Cu, 0.0175% MoS2> 0.27 g!t Au and 1.92 g!t Ag. Drill intercepts of the North Fork deposit also show significant gold values.Mineralization and hydrothermal alteration of these deposits are typical of porphyry deposits, although some differences can be noted. The Mount Tolman molybdenum deposit contains topaz alteration zones and is transitional between a plutonic type and a Climax-type molybdenum system. Other deposits, such as the Kelsey, Mazama, North Fork, and Margaret, are low-sulphide systems that contain significant amounts of magnetite."

Jan 1, 1995

By J. Shibata, N. Murayama, H. Yamamoto

As one of effective usages of wastes discharged in the non-ferrous metal industry, the syntheses of AlPO4-5 and hydrotalcite were carried out using aluminium dross as a raw material. AlPO4-5 is an aluminium phosphate condensate with properties of zeolitic materials and large pore size. Hydrotalcite is an inorganic anion exchanger and has a layered structure of complex hydroxide. They were synthesized in an autoclave at high temperature and pressure. Various physical properties such as crystal structure, surface texture, and specific surface area were measured. AlPO4-5 can be synthesized from aluminium dross under hydrothermal conditions at 180°C-200°C, and 3h. At the same time, AlPO4-34 and non-porous AlPO4 were also formed as by-products. The crystals of AlPO4-5 have a hexagonal structure. Lt is desirable to heat the reaction product at around 550 °C to remove tri-ethyl amine. The specific surface areas of the reaction product before and after heat treatment at 550°C were 17.9m2/g and 358m2/g, respectively. Hydrotalcite can be obtained by a co-precipitation method from the aluminium dross and waste MgCl2 solution discharged in the aluminium regeneration process. Small amounts of SiO2, Al2O3, and Fe2O3 remain in the reaction product as an impurity component. The interlayer distance of the obtained reaction product is about 0.3 nm. Hydrotalcite changes to Mg-Al oxide during calcination at 500 °C for 3h, and then hydrotalcite is formed again by a rehydration operation. The physical properties of the reaction products obtained from the wastes were very similar to those from fresh reagents.

Jan 1, 2004

By G. M. Ritcey, K. E. Haque

"Laboratory-scale sulphuric acid leach tests were conducted on Elliot Lake and Rabbit Lake ores in the presence of various oxidants, namely: Fe2(SO4)3, CI2, H2O2, NaClO3, MnO2, H2SO5, HNO3, Oxon e, K2Cr2O7, KMnO4, O3 and O2. Uranium extraction is significantly accelerated when an oxidant is present in the leach slurry. Uranium extraction of 96-99% can be achieved by maintaining the oxidation potential of the leach slurry in the range of400 to 500 m V (S.H.E.). A rough cost comparison of all the oxidant s indicates that the use of Fe2(SO4)3, 02, O3, C12, HNO3 or MnO2 is far less expensive than NaClO3 and the uranium extraction rates are similar to or better than with NaCIO3. For uranium extraction, it is concluded that gaseous oxygen (i.e., air or oxygen) would be the most effective and economical oxidant for the 24- to 36-hour leach of Elliot Lake ore; whereas for the 8- to I2-hour leach of Rabbit Lake ore, a strong chemical oxidant (e.g. NaCIO3, H202, 03, K2Cr2O7, H2SO5 or HNO3) would be a logical choice.IntroductionThe common practice for the treatment of uranium o res is leaching with sulphuric acid or with a carbonate-bicarbonate mixture. Generally, the latter process is ad opted when an excess (> 18%) of acid-consuming components (e.g., carbonate) is present in the ore. Irrespective of the nature of the leachant, the addition of an oxidant to the leach mixture is necessary to obtain a high uranium extraction.The maximum extraction of uranium can only take place when all the uranium is in the highest oxidation state (i.e.,U +6-uranic ion) . Because a certain percentage of uranium present in the ore is invariably in the lower oxidation state (i.e.,U+4-uranous ion) , the addition of an oxidant to the leach slurry is necessary to oxidize uranous ion to the uranic state."

Jan 1, 1982

By Floyd B. Bowen

Pumping methods are used, almost exclusively, to transport materials in the mining, beneficiation and chemical processing of phosphate rock in the Florida phosphate pebble field. Matrix pumping systems capable of pumping 15,000 g.p.m., transporting 15,000 long tons (dry) per hour, are now in use. They utilize 16-by 48 Yz -in. pumps, driven by 1.250-h.p. motors, discharging into pipelines 20 inches .in diameter .and sometimes several miles in length. As an average, it is necessary to pump 3 tons of matrix to the beneficiation plants, and to dispose of 2 tons of sand tailings and clay as waste, in order to produce 1 ton of shipping-grade rock. The washers and flotation plants, where the clay and phosphate and silica particles are separated, require the movement of large volumes of slurries with densities ranging from 2-3 per cent to 45 per cent solids, by weight. Extensive waste disposal systems require the pumping of dilute clay slurries in quantities of up to 50,000 g.p.m.,

Jan 1, 1963

By N. E. Chaouai, M. Lavigne

"Indicator kriging is one of the non-parametric geostatistical techniques, ideal for estimating the reserves of irregular mineralizations (e.g. gold, uranium, platinum, diamonds, etc.). Such mineralizations are characterized by variograms which are very sensitive to extreme values (outliers) and very difficult to model.The classical approach to this problem in the industry is to trim off these high grades using arbitrary empirical upper limits established either from practice or subjectively. Contrary to this simplistic approach, indicator kriging discretizes the histogram of the grades (usually log normal) in several classes and carries out interpolation separately for each class. This paper gives an overview of the differences between ordinary and indicator kriging along with the steps required to carry out indicator kriging. The paper concludes with a case study involving a gold deposit estimation.IntroductionParametric geostatistics (ordinary kriging) has given us satisfactory results over the years in estimating the reserves of well behaved deposits with coefficients of variation close to 1 (e.g. sedimentary and porphyry type deposits). However, numerous problems have been encountered in estimating irregular mineralizations (e.g. small vein deposits of precious metals) with high coefficients of variation in the range of 2 to 5. Such deposits are characterized by a very small proportion of extreme values (outliers) which create distortions on the variograms (pure nugget effect) and render them useless(1,2).The solution often followed by the mining industry today is to trim off these values based on empirical limits. These limits are derived either by comparison with production results (or bulk samples) or by geological interpretations. Needless to say that such manipulations can lead to serious errors in estimating ore reserves (usually underestimation of total metal quantity), especially when production results or bulk samples are absent.On the other hand, the solution followed by the geostatisticians in the past was to smooth out the data using transforms (e.g. normal-score or natural logarithms). Such transforms are nonlinear and they required non-linear estimation techniques such as disjunctive kriging (DK) or multivariate gaussian kriging (MG) which require normal related hypotheses, not necessarily true in geology.Nonparametric geostatistics (indicator and probability kriging) have been -developed in order to handle these problems related to highly skewed distributions."

Jan 1, 1990

By Mehmet F. Tauer, Pierre Trudel

"The metamorphosed mineralized Archean Bourlamaque batholith, dated at 2710 +-54 Ma, is a homogeneous body, composed mainly of quartz diorite; it is cut by late dioritic, mafic and aplitic dykes, and contains metasomatized ultramafic xenoliths. A geochemical investigation confirms the calc-alkaline affinity of the batholith, and the REE profiles are similar (Lal Yb = 3.9 to 8.1) to those of the intermediate to felsic metavolcanic rocks at Val d'Or into which the batholith was intruded and to those of the calc-alkaline members of the Blake River Group at Rouyn.Gold distribution within the batholith is heterogeneous and anomalous. Although the background Au content of the batholith is estimated to be 3.2 ppb Au by RNAA , the estimated median gold content of the batholith is 10 pp b (with P84 =32 and PI 6=3.1 ppb), much higher than that of ""normal"" granites (I to 5 ppb A u). The gold mineralization exhibits a close relationship with shear zones (e.g. Belmoral Mine) or with diorite dykes (e.g. Sullivan, Perron and Courvan mines).The following sequence of geological events is proposed: (I) pretectonic emplacement of the batholith, which is cogenetic with its volcanic host rocks; (2) regional metam orphism and deformation with localized shearing; and (3) hydrothermal alteration and quartz veining. We conclude that: (1) the Bourlamaque batholith is a gold-enriched intrusion, and (2) the determination of average gold content may be an effective method f or distinguishing ""sterile"" intrusions from those with potentials for gold mineralization.RESUMELe batholite archéen minéralise de Bourlamaque, daté de 2710+-54 Ma, est une intrusion homogène, constituée principalement de diorite quartzique, qui a été recoupée par des dykes tardifs de diorite, d'aplite et de composition mafique. Il contient aussi des enclaves de roches ultramafiques metasomatisees. L 'étude géochimique confirme le caractère calco-alcalin du bath olite et les profits des ETR sont semblables (La/ Yb=3.9 a8.1) a ceux des volcanites intermédiaires afe tsiques de la région de Val d'Or dans lesquelles le batholite fait intrusion et à ceux des unites calco-alcalines du Groupe de Blake Ri ver a Rouyn. La distribution de Au dans le batholite est à la fois anomale et hétérogène. Bien que le bruit de fond du batholite ait été estime à 3.2 ppb A u par AANR, sa médiane a été estimée a10 ppb Au (avec P84=32 et PI6= 3.1 ppb) , ce qui est nettement plus élevé que la valeur des roches granitoides ""normales"" donnée dans la littérature (1 a 5 ppb A u). Dans la batholite, la minéralisation aurifère est intimement associée aux zones de cisaillement (e.g. mine Belmoral) au aux dykes de diorite (e.g. mines Sullivan, Perron et Courvan).La séquence des évènements géologiques suivante est proposée: (1) emplacement pre-tectonique du batholite, lequel est cogenetique avec les roches volcaniques encaissantes; (2) deformation et metamorphisme regional avec des cisaillements tectoniques locaux; et (3) altération hydrothermale et développement des veines de quartz. Nous concluons que: (1) le batho lite de Bourlamaque est une intrusion enrichie en Au; (2) la determination de la teneur en Au peu être une méthode efficace pour distinguer les intrusions stériles de celles qui possèderaient un pot entiel pour des minéralisation s aurifères."

Jan 1, 1989

By K. B. Hodgins, P. Ouellette

"The North American mining industry faces challenging emissions regulations in the coming years with the introduction of new standards for off-road diesel engines. In order to remain competitive, industry stakeholders must invest in new technologies that are economically and environmentally sustainable.Westport Innovations Inc., Vancouver,British Columbia, has developed high-pressure direct injection (HPDI) technology that allows diesel engines to run primarily on natural gas with no intrusive engine modifications while reducing emissions of nitrous oxides (NOx), particulate matter (PM), and CO2 with no loss in performance. This is accomplished through late-cycle, high-pressure direct injection of natural gas into a diesel engine combustion chamber at the end of the compression stroke. A small amount of diesel fuel is injected first to initiate combustion, followed by the natural gas fuel main injection. The diesel cycle and its operating characteristics are fully retained including high combustion efficiency, lowspeed torque, high-power output, no engine knock, and fast response to load changes. Natural gas has not been an attractive fuel alternative in mining operations due to the reduced performance associated with spark ignition engines and the added weight and space required for compressed natural gas tanks. However, liquefied natural gas (LNG) can be stored with less impact on weight and space than compressed natural gas (CNG) by virtue of its increased energy density over CNG.Westport has applied HPDI technology to the Cummins QSK series engines for high horsepower applications such as power generation and mine haul trucking. HPDI can also be applied to other engine manufacturer platforms and market applications. Westport is interested in demonstrating the unique characteristics of HPDI in mine haul trucks to show diesel performance with low emissions levels using low-cost LNG fuel.Retrofitting a Tier 1 engine with HPDI would provide emissions reductions below the EPA 2006 Tier 2 standards while maintaining the performance and efficiency of the Tier 1 platform. The longer-term target is the Blue Sky voluntary standards of 3.8 g/kW-hr (2.8 g/bhphr) of NOx + non-methane hydrocarbons (NMHC), and 0.12 g/kW-hr (0.09 g/bhp-hr) of particulate matter (PM) and future standards. Typically, an HPDI engine is capable of reducing NOx emissions by 40% over diesel operation at the same efficiency. On a mining truck meeting Tier 2 standards, this NOx reduction would translate into NOx emissions approximately 14 t per year lower than under diesel operation. PM emissions are typically reduced by some 60% with HPDI fueling, which would translate into a yearly reduction of 650 kg. Similarly, engine-out greenhouse gas (GHG) emissions are reduced by 21%, resulting in a yearly reduction of approximately 950 t per vehicle.Above those of emissions reduction, efficient natural gas usage in mine haul operations would offer significant benefits in terms of fuel cost savings, which may be further combined with monetary emission credits. HPDI technology would enable operators to exercise environmental stewardship while gaining a competitive advantage in advance of new Tier 2 and Tier 3 emissions regulations. The advantages of natural gas use among mine haul truck operations can be replicated across a broad range of engines and applications."

Jan 1, 2005

By John G. Payne

The Pebble deposit contains three contiguous zones. In the Pebble West zone, granodiorite stocks intrude gently warped volcanosedimentary rocks that had been intruded by diorite sills and two felsic intrusion breccias. In the Central zone, volcanosedimentary rocks are cut by diorite and granodiorite sills; the latter are genetically associated with the granodiorite stocks. In these zones, widespread potassic alteration consists of brown biotite and K-feldspar, with abundant quartz-carbonate-sulphide vein stockworks. Chalcopyrite/pyrite ratios range from 2/1 to 1/20. Diorite hosts the highest-grade mineralization, with much of it being disseminated. Patchy, later phyllic alteration is moderate and commonly depleted in copper. The Pebble East zone occupies a granodiorite stock and intruded host rocks below an eastwardly thickening wedge of barren Tertiary volcanic and sedimentary rocks up to 600 m thick. Brown biotite alteration and copper mineralization are strong, with chalcopyrite/pyrite ratios commonly over 1/1. The zone extends to a depth of at least 1200 m below surface; some drill holes bottomed in strong mineralization. Later, weak to moderate sericitic alteration removed some chalcopyrite and locally remobilized gold. The Pebble deposit has a Measured/Indicated resource of 3 billion tonnes grading 0.28% Cu, 0.32 g/t Au and 0.015% Mo (= 0.5% copper-equivalent) and an Inferred resource of 1.13 billion tonnes grading 0.30 g/t Au, 0.27% Cu and 0.015% Mo (0.55% copper-equivalent). The Pebble East zone has an Inferred resource of 2.13 billion tonnes grading 0.55% Cu, 0.34 g/t Au, and 0.037% Mo (0.97% copper-equivalent). The Pebble East zone is open to the east and to a lesser extent to the north, south, and to depth.

May 1, 2006

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

In this study, the dissolution behavior of copper and arsenic from five different copper concentrates containing 21-34.7 %wt. of Cu, 1.3-7.3 %wt. of As and 15.6-27.5 %wt. of Fe were investigated under high pressure oxidative leaching conditions. Leaching experiments were conducted using different leaching reagents such as ferric sulfate, sodium chloride, deionized water, sulfuric acid and ferric sulfate-sodium chloride mixture in order to compare the effectiveness of each on the dissolution of copper and arsenic from copper concentrates. The effects of ferric sulfate (0.1-1 mol/L), sodium chloride (0.1-0.5 mol/L) and leaching temperature (100-180C) were investigated under 1 MPa autoclave pressure , particle size >45 m, pulp density of 100 g/L, stirring speed of 750 rpm and leaching time of 60 min. Results showed that the most effective leaching reagents for the dissolution of copper are 0.34 M NaCl and a mixture of 0.1 M Fe2(SO4)3 - 0.34 M NaCl, in which As extraction did not exceed 7%. The copper concentration in the pregnant leach solution obtained after leaching using 0.1 M Fe2(SO4)3 - 0.34 M NaCl was 13-30 g/L, while the concentration of As and Fe were in a range between 0.3-1 g/L and 0.2-10 g/L at 160C. INTRODUCTION Copper sulfide minerals, such as chalcopyrite (CuFeS2), chalcocite (Cu2S) and covellite (CuS) are used for the production of copper in mining and metallurgical industries worldwide. The conventional pyrometallurgical processing followed by electrorefining is a common process for the production of high purity copper (Moskalyk, & Alfantazi, 2003; Padilla, Aracena, & Ruiz, 2012). However, the degradation of copper content and an increase in the amount of metal impurities, especially arsenic (As), antimony (Sb) and bismuth (Bi) in copper concentrates cause serious problems to the production of copper and environmental pollution (Basha, Selvi, Ramasamy, & Chellammal, 2008; Long, Peng, & Bradshaw, 2012). Hydrometallurgical processes are used to recover metals from refractory ores, concentrates and tailings by leaching (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018). During the last decades, hydrometallurgical processing for the recovery of copper from enargite-bearing copper ores and concentrates has been extensively investigated due to its relatively more environmentally benign and less capital costs than the conventional pyrometallurgical processes (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018; Choubey, Lee, Kim, & Kim 2018). However, several issues facing the hydrometallurgical processing are incomplete extraction of copper associated with the coprecipitation of iron compounds, difficult stabilization of toxic impurities and production of a contaminated copper product (Dreisinger, 2003; Safarzadeh, Moats, & Miller, 2014). Hence, a number of studies have been conducted on the selective dissolution of copper from copper ores and concentrates or selective removal of arsenic leaving a clean copper residue using various leaching reagents or lixiviants (Safarzadeh, Moats, & Miller, 2014; Mihajlovic, Strbac, Zivkovic, Kovacevic, & Stehernic, 2007; Tongamp, Takasaki, & Shibayama, 2009). More extensive studies on the selective extraction of copper from enargite- bearing copper ores and concentrates are much needed to improve copper recovery.

Jan 1, 2019

By M. B. Kime, E. K. Kanowa

"Heap leaching of low-grade copper-cobalt oxidized ore and solvent extraction of copper, cobalt, manganese, and zinc were investigated. Heap-leaching parameters were optimized using agitated and tray batch leaching tests, capped by tests on a pilot semicontinuous heap-leach plant. The optimum pregnant leach solution yielded 2,350 mg/L Cu, 3,800 mg/L Co, 3,600 mg/L Fe, 130 mg/L Mn, 25 mg/L Zn, and 0.58 g/L solids in suspension, with pH and acidity of 1.27 and 3.66 g/L, respectively. Copper, cobalt, and impurities (Mn and Zn) were extracted using (v/v) 30% CP-150, 20% IONQUEST® 290, and 20% D2EHPA, respectively; overall recoveries of copper, cobalt, and impurities were 98, 99, and 70%, respectively.RÉSUMÉ La lixiviation en tas des minerais oxydés de cuivre et de cobalt et l’extraction par solvant du cuivre, du cobalt, du manganèse et du zinc ont été étudiés. Les paramètres de lixiviation en tas ont été optimisés à l’aide d’essais de lixiviation en discontinu sur un plateau, complétés par des essais relatifs à une usine pilote de lixiviation en tas en semi-continu. La liqueur-mère optimale a produit 2 350 mg/L de Cu, 3 800 mg/L de Co, 3 600 mg/L de Fe, 130 mg/L de Mn, 25 mg/L de Zn et 0,58 g/L de solides en suspension, avec un pH et une acidité respectifs de 1,27 et 3,66 g/L. Le cuivre, le cobalt et les impuretés (Mn et Zn) ont été extraits avec un pourcentage volumique (v/v) de 30 % de CP-150, de 20 % d’IONQUEST ® 290 et de 20 % de D2EHPA, respectivement. Les récupérations totales de cuivre, de cobalt et d’impuretés étaient de 98 %, 99 % et 70 % respectivement.INTRODUCTIONBoss Mining S.P.R.L. is a privately owned hydrometallurgical plant located in southeastern Democratic Republic of Congo. It treats oxidized ores of copper and cobalt that are recovered by sulphuric acid heap leaching (Trujillo, Cisternas, Gálvez, & Mellado, 2014; Kime, Kanowa, Mulaba- Bafubiandi, & Diyambi, 2016) and solvent extraction processes (Crane, Urbani, Dudley, Horner, & Virnig, 2009; Fisher, 2011; Kime et al., 2016). The ore to the plant is primarily sourced from the Luita and Mukondo mines. Geological prospecting work by Gecamines and Boss Mining showed that the Mukondo mine contains a very large cobalt reserve not yet exploited, but with grades as low as 0.2%.Currently, Boss Mining is focusing on the production of copper, but is investigating methods to efficiently produce both copper and cobalt. A challenge in the production of cobalt from this ore is its presence in the third oxidation state. This requires the addition of reducing agents, which increases the overall operating cost (Kime & Makgoale, 2016). Cobalt solvent extraction is also very sensitive to impurities such as nickel, zinc, manganese, and aluminum, which are often co-extracted (Swartz, Donegan, & Amos, 2009). This sensitivity to impurities requires the optimization of operational parameters to lower impurity levels as much as possible in the electrolyte."

Jan 1, 2017