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By Xuanxue Mo, Zengqian Hou

"Yidun Island Arc in Sanjiang Region is one of the important Kuroko-type volcanogenic massive sulfide deposits (VMS) districts in southwestern China. Tectonically, the region is situated in the eastern portion of the Tethys orogenic belt, and is a late Triassic extensional island-arc. The island-arc underwent earlier arc construction, intra-arc rifting, later arc construction and backarc spreading during its evolution and formed a complete trench-are-basin system, i.e., the Ganzilitang suture zone, forearc basin, outer are, intra-arc rifting zone, inner arc and back-arc basin from east to west across the island-arc.Kuroko-type VMS deposits and occurrences discovered so far are limited to within the intra-arc rifting zone of the Yidun Island Arc and tend to occur in groups or clusters distributed in separate faultbound basins; e.g., Zengke, Gacun and Xiangcheng basins, in which submarine volcanic rocks, especially bimodal volcanic suites, were extensively developed. The VMS deposits formed in the waning stages of volcanic eruption in the late Triassic and are hosted in the uppermost section of acidic volcanic rocks (rhyolitic-dacitic) within a bimodal suite. The orebodies are associated with these host rocks: rhyolitic-dacitic pyroclastic rocks and exhalites and chemical sedimentary rocks.All VMS deposits in Sanjiang are of the Zn-Pb-Cu type. The architecture of the deposits is twofold, consisting of concordant lenses and layers of massive sulfides and an underlying discordant stockwork. All deposits show metal zoning. The hydrothermal alteration related to mineralization is strong and asymmetrical, and is found dominantly within the stockwork zone and host rocks surrounding the ores. There are two alteration patterns: (1) chloritization - sericitization - silicification from outer sides to core of the alteration pipe; and (2) silicification + sericitization - silicification + carbonate from the lower to upper part of the alteration pipe.The mineralization temperatures of ore-forming fluids vary in the range 200°C to 335°C. Thermal evolution of the fluids displays temperature profiles that first increase and then decrease with progression of mineralization processes. The majority of black-ore minerals (Groups I and informed during intensifying stages of hydrothermal activity, and the majority of semi-yellow-ore minerals (Group III) formed during a thermal maximum. There followed the formation of fine-grained black-ore minerals (Group IV), which interacted with barite layers during decreasing temperature phases.Oxygen and hydrogen isotope studies show that the sources of ore-forming fluids for the Kurokotype VMS deposits are two-fold: seawater and magmatic hydrothermal solutions, represented by the Kuroko deposits in Japan and Kuroko-type VMS deposits in Sanjiang, respectively. In the latter case the hydrothermal solutions reacted with host rocks and mixed with cold seawater percolating downward, and circulated through volcanic rocks resulting in development of alteration zones and formation of ore minerals in Kuroko-type VMS deposits."

Jan 1, 1993

By John Lamb

"The French Mine (formerly the Oregon) is 4 miles by road south of the Nickel Plate Mine. It lies on the steep eastern slope of the Similkameen River valley at elevation 4,000 feet, between Cahill and Winters Creeks.The property was discovered in .1901 and bought 4 years later by Messrs. French and Deardorf. It was developed in a modest way by the owners and by several companies until 1947, without success. The spectacular showings of bornite and chalcopyrite in the outcrop drew attention to its possibilities for copper rather than gold. Diamond drilling by the owners in 1947 yielded several high-grade gold intersections and created new interest. In 1949 the property was purchased by Kelowna Mines Hedley Limited. Exploration drilling and geological mapping outlined sufficient ore to commence production in 1950. Except for a few months each winter the operation has been continuous, and to the end of 1954 had produced almost 30,000 tons of high grade ore. The ore is crushed at the mine and hauled 8 miles by truck to the Company mill at Hedley, where it is treated together with the ore from Nickel Plate.Regional GeologyThe country rock consists of brown tuff, flows, limestone, and cherty formations belonging to the Nicola group (Rice 1947) of Triassic age. West of and just below the mine is a large mass of granodiorite which is part of the same body that underlies Nickel Plate Mountain. Bedding attitudes in the mine area are locally complicated, but dips are generally less than 45 degrees. To the south and below the mine the beds stand almost vertical, continuing downward with this attitude for 2,000 feet to Winters Creek. The main granodiorite contact is also steep, conforming to the sedimentary formations, whereas northward towards Nickel Plate it dips gently to the northwest, slightly flattter than the thick westerly dipping pile of sediments above it. The interpretation of this structure by Billingsley (1936) is that the French mine lies near the crest of the large northeasterly trending asymmetrical anticline which contains in its gently dipping western limb the Nickel Plate orebodies. The granodiorite body conforming roughly to this shape seems to occupy the core of the anticline. Sill-like prongs of granodiorite extend upward to the east from the main mass, outcropping over wide areas at the head of Winters Creek (Figure 1). There is a possibility that overthrusting from the west, similar to that on Nickel Plate Mountain, occurs not far below the French mine.An interesting feature of the granodiorite-sediment contact is that where it dips gently westward no alteration exists in the sediments, but where it has a steep front pronounced heat effects are indicated by the presence of a thousand-foot-wide zone of reddish garnet skarn."

Jan 1, 1949

"Industry Guide 7 contains the Securities and Exchange Commission’s basic mining disclosure policy. It provides guidance for Registrants on various matters including the definition and disclosure of proven and probable reserves and the definitions of exploration, development and production stages. The definitions and disclosure instructions contained in this Guide apply to all public mining entities and their public disclosures. As NI 43-101 is law in Canada, the SEC allows mineral resources and mineral reserves prepared under CIM Definition Standards for issuers that are reporting in Canada but are also registered with the SEC. The SEC places conditions on the disclosure of these estimates. US shareholders are cautioned that the requirements under NI 43-101 and the CIM Definition Standards differ significantly from those of Industry Guide 7.Note: this Guide contains the Commission's basic mining disclosure policy. The definitions and disclosure instructions contained in this Guide apply to all public mining entities and their public disclosures pursuant to the rules of Regulation S-K. Questions regarding the application of these definitions and rules to specific situations should be asked of the Commission's mining engineers or geologists at either the Office of Engineering, Division of Corporation Finance, 202-272-3252, or the Branch of Geology, Central Regional Office, 303-844-1000.(a) Definitions. The following definitions apply to registrants engaged or to be engaged in significant mining operations:(1) Reserve. That part of a mineral deposit which could be economically and legally extracted or produced at the time of the reserve determination.Note: Reserves are customarily stated in terms of DoreD when dealing with metalliferous minerals; when other materials such as coal, oil shale, tar, sands, limestone, etc. are involved, an appropriate term such as Drecoverable coalD may be substituted.(2) Proven (Measured) Reserves. Reserves for which (a) quantity is computed from dimensions revealed in outcrops, trenches, workings or drill holes, grade and/or quality are computed from the results of detailed sampling and (b) the sites for inspection, sampling and measurement are spaced so closely and the geologic character is so well defined that size, shape, depth and mineral content of reserves are well- established.(3) Probable (Indicated) Reserves. Reserves for which quantity and grade and/or quality are computed from information similar to that used for proven (measured) reserves, but the sites for inspection, sampling, and measurement are farther apart or are otherwise less adequately spaced. The degree of assurance, although lower than that for proven (measured) reserves, is high enough to assume continuity between points of observation.(4)(i) Exploration StageDincludes all issuers engaged in the search for mineral deposit (reserves) which are not in either the development or production stage."

Jan 1, 2003

By T. I. . I. Sibbald, D. E. Jiricka

The Goldfields area on the north shore of Lake Athabasca , was active in the thirties and early fortie . In 1934, Tom Box discovered gold mineralization in quartz veins in pyritized 'granite' at the ite of what wa to become the Box Mine of the Consolidated Mining and Smelting Company of Canada Ltd . In the rush of activity that followed numerou other gold occurrences were found. the most significant of the e being the Athona and Frontier prospects which underwent underground development bur did not reach production. Probable reserve identified to date in the Box and Athona deposits amount to some 3.5 million ton grading an average 0.ITT oz gold/ton . Some 10 to 15 million tons of lower grade, po ible reserve are also reported at the Box. From 1939 to 1942, the mine produced, from underground , 64,066 ounces of gold from ore grading 0.0452 oz gold/ton at a cost of between $1.5 and $2.0 a ton. The Goldfield area is underlain by upracrustal rock intruded by mafic bodies and invaded by ·granite'. The upracru tal . thought to be of Lower Proterozoic age are characterized by a helf-type ucce ion metamorpho ed under lowermiddle amphibolire facie conditions into quartzites, quartz-mica chi ts. dolomitic marbles, diop idic calc-silicates and cordierite-anthophyllite-biotite schi t . Amphibolites of presumed igneou origin form concordant to subconcordant units up to a few ten of metre thick in the upracrustals and occur as le regular masses in the 'g ranite·. Compositionally they vary from gabbroic to granodioritic. with the more mafic varieties being most common. Granitic rocks exhibiting a variety of textural , structural and compositional characteristics underlie most of the area. Four main units are recognized: foliated (older) 'granite', 'leucogranite', feldspar augen 'granite· and pyritic mine 'granite', however, in geological mapping the e are somewhat arbitrarily differentiated due to transitional relationships and admixture of types. Although there is no unambiguous age data available for the e 'granite ' adjacent northshore 'granite' of geologically similar age yield Rb-Sr isochron dates around 2000 Ma. The area is characterized by two major deformation events of which the earliest gave rise to the dominant regional foliation and to rare minor and outcrop scale folds. This foliation which in general parallels compositional layering is folded into the major Goldfields syncline, whose axis plunges weakly to moderately outh. Related open minor folds have subvertical axial surfaces which are locally paralleled by a weakly defined schistosity. Most rock display a pronounced rodding fabric, coaxial with second fold axes and apparently the product of this deformation event. However, a note of caution is introduced into this interpretation, as the rodding fabric is also parallel to first fold axe .

Jan 1, 1986

By R. J. McLeod

"A new backfill plant was constructed in 1966 adjacent to the mill of Giant Yellowknife Mines. The reasons for the addition were to increase the storage capacity for sand fill, provide facilities for mixing cement with the fill for mine stope floors and provide a system for pumping fill to B -3 raise, a horizontal distance of 4700 feet from the mill. At the same time it was hoped to increase the density of the fill pumped underground.The plant consists of two only steel storage tanks of approximately 400 tons capacity of sand fill each. They are 17 feet in diameter and have conical bottoms. The cylindrical section is 28 feet high and the conical section 14 feet high. Each tank is fitted with water nozzles around the conical sections to remove solids at the required density. The water pressure on the nozzles is 150 lbs per square inch. The cyclones used to produce the sand back fill are located on top of the storage tanks. Five model Dl0B Kreb cyclones are used for this purpose. The vortex finders are 2 1/ 4 inch and the apex 1 1/4 inch. The cement silo or storage tank is located in the same building. It is a steel tank 17 feet diameter by 41 feet high with a conical bottom. The silo holds 250 tons of bulk cement, which is unloaded by air from a bulk highway carrier using 12 to 15 lbs air pressure. The cement is conveyed to the top of the silo, a distance of 70 feet from the ground by 450 lbs. of air.The mixing tank is a rubber lined steel tank 14 feet in diameter by 14 feet high. It is equipped with a Lightning Mixer supplied by Greey Mixing Equipment Ltd. The mechanism is driven by a 40 HP motor. The lower shaft is unsupported and is equipped with two 36 inch diameter axial flow turbine type rubber covered impellers. The backfill is discharged from the storage tank into the mixing tank by a 5"" x 4"" SRL-C pump powered by a 20 HP motor. A level control in the mixing tank regulates the rate of discharge of this pump through a Grinnell-Saunders automatic control valve. The cement silo is situated at such a height above the mixing tank to allow for a Rota-Lock, a small hopper and a cement feeder. The feeder is an Omega gravimetric belt type dry feeder with a Sensogram controller. The capacity of the cement feeder is 20 tons of cement per hour. The cement and backfill are mixed in the mixing tank at whatever ratio is required for the mine stope floors."

Jan 1, 1968

By K. Woo

"With ever-increasing global demand for mineral resources, mining companies are considering developing deeper, more complex and lower grade ore bodies. Recent years have seen the transition from surface to underground mass mining operations to access deeper resources, with mass mining methods such as block and panel caving being favoured due to the tonnages and economics achievable when dealing with lower grade ore. However, the rock engineering interactions involved with such plans are complex and challenging given the significant disturbance caused to the surface environment. If not properly accounted for, the unexpected propagation of a cave breaking through to surface or the differential surface displacements otherwise produced may threaten the safety and integrity of overlying mine and civil infrastructure.This paper reports the findings of a comprehensive research programme focussed on investigating and integrating new developments in remote sensing characterization and monitoring technologies with state-of-the-art numerical modelling to better manage the challenges and geo-risk associated with the transitioning from open pit to underground mass mining. Procedures were developed to integrate discrete fracture networks (DFN) into finite-element/discrete-element analyses to simulate the influence of rock mass fabric on cave propagation and the evolution of caving-induced surface subsidence and step-path failure of large open pit slopes through the progressive failure of rock bridges. The results clearly showed that caving-induced surface deformations tend to be discontinuous and asymmetric due to large movements around the cave controlled by geologic structures, rock mass heterogeneity and topographic effects. Although parallel investigations of empirical design charts used to predict caving-induced deformations were found to be moderately sufficient for assessing the extent of macro deformations directly above the undercut (i.e. caving collapse features), numerical modelling was found to provide the only means to confidently estimate the lateral extent of smaller strain subsidence.These results were further supported by the use of high-resolution synthetic aperture radar (SAR) data to monitor the ground deformations associated with block caving production, and to use these to calibrate and constrain a series of predictive 3-D numerical models focussed on Rio Tinto's Palabora mine in South Africa. The InSAR data was obtained from Canada’s recently launched RADARSAT-2. The close fit achieved between the predictive 3-D numerical model and InSAR monitoring data demonstrates the significant value of InSAR calibrated 3-D numerical models applied to subsidence prediction.Collectively, the results from this research programme have helped to further the characterization, assessment and understanding of mass mining-induced subsidence, and its evolution, by addressing existing limitations in the use of empirical and numerical design methods. The limitations and uncertainty arising from mine site data were detailed, specifically the representation of mine geology, rock mass properties, in-situ stresses and cave propagation, together with means to constrain these inputs and calibrate sophisticated 3-D numerical models through back analysis and integration with high resolution monitoring data."

Jan 1, 2015

By S. C. Tadolini

"Roof bolts are the primary means of roof support in underground mines in the United States. Presently, post installation loads are measured on about 10% of the expansion anchor and combination bolts by using standard torque wrenches that Unit the accuracy of the measurements to about ±30%. No system has been available for measuring the loads generated on full-column resin grouted bolts. Mine roof bolts are long, slender, non-uniformly sized, roughly forged, and placed in a corrosive environment, creating obstacles for researchers trying to utilize ultrasonic reflection techniques to measure strain. A complete measurement system has now been developed by the U.S. Bureau of Mines, Denver Research Center. This system is being used successfully in determining loads measured on expansion anchor and combination bolting systems, examining the transfer mechanics of full-column, resin-grouted bolts, and measuring loads generated on bolts in an underground roadway support system.IntroductionMillions of roof bolts are used each year to provide structural support in underground U.S. metal, non-metal, and coal mines. Approximately 45% of the bolts used are expansion anchor, 45% are resin-assisted combination or full-column resin-grouted, and 10% are friction anchored. The effectiveness of these bolting types is dependent on a number of factors including rock density, bolt length, in situ stresses, rock properties, bolt pattern, installation tension, etc. When implementing expansion anchor and resin assisted combination bolts , installation and post-installation tension, or load, is a major determinant of support quality and subsequent behaviour of the reinforced roof. The method by which these systems carry roof loads is well understood and documented(1,2).The importance of these installation and long-term bolt tension measurements is well recognized and are required in the federal government regulations(3). A statistical sample of the bolts, lout of 4 during installation and lout of 10 inby the last open crosscut, is required to be measured using current torque wrench technology. The disadvantages of this system are the disturbance and weakening of the anchorage and limitations of accuracy to approximately ± 30%. The method by which resin-grouted bolting systems carry roof loads is much less understood, mainly because testing procedures used to evaluate point-anchor systems are not effective on resin-grouted bolts.Due to these problems, the U.S. Bureau of Mines designed a measurement technique that provided a highly accurate, low cost-per- bolt system that led to the development of ultrasonic pulse echo instrumentation. The original system was designed to measure the tension only on expansion anchor bolts. Recent developments and system modifications have permitted measurements on resin-assisted and full-column resin-grouted bolts. The use of these types of support systems has been increasing steadily in underground coal mines.The development of this instrumentation has provided laboratory and field results that help to explain the mechanics and relationships between the applied loads and the interaction between the grouted bolt system consisting of the bolt tendon, grout, and rock."

Jan 1, 1990

By V. Maazallahi, A. Rouleau, A. Saeidi

"The geometry of the terrain topography before and after the failure are the essential factors for analyzing a landslide and for the design of remedial works. This paper presents a method to determine these factors, considering that the determination of the slide surface profile is difficult and costly using traditional methods and the pre-slide topography data is often not available. The presented method to locate the 3-dimensional geometry of the slide surface is based upon determining the slide surface in 2- dimensional vertical sections. This method utilizes post-slide topography of the area to determine the pre-slide topography and uses equations to estimate the pre-slide situation using a statistical process applied to points located on the sliding mass. This method is tested on the Maskun landslide in Iran. The results indicate the applicability of the proposed method to obtain the key factors for the back analysis of a landslide in a cost effective manner.INTRODUCTIONProbably the most reliable method of determining the strength of soil and rock mass is to back analyze a failed or failing slope (Caia et al. 2007; Wyllie & Mah, 2004). This method is used in geotechnics for estimating geotechnical parameters of the rock and soil mass (Sakurai et al., 2003; Frayssines and Hantz, 2009; Cepuritis et al. 2010), but it requires that the failure mode is well established and that there is complete and precise information available on the slide surface and the sliding mass (Sjöberg, 1996). Measurement of the slope profile before failure provides the basic information required for a back analysis (Hock & Bray, 1981). The post-slide topography is another key factor. The back analysis process for determining the actual strength parameters of soil or a decomposed rock mass after a landslide requires knowledge of the sliding mass geometry, both before and after the sliding event. Many studies are reported in the literature about the application of back analysis for determining geotechnical and geomechanical parameters of rock and soil mass (Sakurai et al., 2003; Cepuritis et al. 2010; Frayssines and Hantz, 2009, Hussain and Stark, 2011), but they generally present little information about the geometry of the slide surface, and the weight of rock and soil mass that slid, although these factors are the most important parameters for back analysis process. Liang and Xue-song (2012) determined the slip surface having the minimum factor of safety (called critical slip surface in deterministic analysis of slopes) and the slip surface with the minimum reliability index (called critical reliability slip surface), but they have not considered the actual profile and geometry of the slide surface. Rocscience (2013) proposes the back analysis option in the Slide software for slope stability study, but it requires an imposed slide surface as input; alternatively, one can estimate by a finite element analysis the surface with the minimum factor of safety. However, this estimated surface is very likely different from the actual sliding surface, considering the uncertainties in input data of the finite element method."

Jan 1, 2015

By K. Swedburg, M. Samuels

"The expansion of the concentrator at New Gold’s New Afton copper-gold mine in British Columbia, Canada, completed in 2015, included an increase in the required retention time of the cleaning circuit. On site pilot testing of a Woodgrove Staged Flotation Reactor (SFR) had shown that a high grade copper concentrate would be produced with excellent recovery. The SFR was selected rather than conventional tank cell technology based on the much smaller footprint (which rendered a separate and expensive building extension unnecessary), high concentrate grades demonstrated in testing, and the lower energy requirement. Three SFRs, each capable of taking a nominal feed of 700 cubic meters per hour were installed and commissioned 10 months after the receipt of order. The SFRs operate in series and all of the cells produce a final concentrate typically grading 32% Cu, while the overall SFR stage recovery averages 55%. The tailings from the SFRs are treated in the original conventional three-stage cleaner circuit where the remaining values are recovered in a concentrate grading 28% Cu. Design production was achieved within days of the initial start-up of the SFR circuit.INTRODUCTIONThe New Afton concentrator, located near Kamloops, British Columbia, Canada processes a copper-gold ore with a designed head grade of 0.96% Cu, 0.72 g/tonne Au, and 2.22 g/t Ag.The New Afton Mine is a block cave mining operation. Ore is undercut from the bottom up and removed at the extraction level from a grid of draw points. Ore is trammed to the ore passes by underground loader and dropped to a haulage level for transport by trucks to the gyratory crusher. From the crusher, the ore is conveyed to the mill via a 4.5 kilometre long conveyor system. Block caving was chosen for the New Afton deposit because the underground infrastructure could be built below the historical open pit of the Afton Mine (a previous operation of Teck Resources) and this method is favorable to large-scale low-cost mining.The New Afton deposit is a copper-gold, alkalic porphyry system. Hypogene mineralization is characterized by discontinuous copper sulphide veinlets and disseminations, principally chalcopyrite and minor bornite, along with pyrite as the second most dominant sulphide. Gold particles vary in size from 5 µm to 40 µm and occur as native gold and electrum, associated mostly with the chalcopyrite grains and never with the pyrite. In some instances, a second alteration phase has resulted in the copper sulphides being replaced by tennantite-tetrahedrite locally and along faults that transect the mineralized body. The design head grade of arsenic is 244 PPM, although the presence of tennantite is inconsistent and can vary from trace levels to 15% of the total copper. No arsenopyrite is present in the hypogene ores. Supergene mineralization, consisting of native copper with accessory chalcocite occurs in minor amounts along faults and highly oxidized near-surface portions of the deposit."

Jan 1, 2016

By Rodrigo Serna, Nora Schreithofer, Zoltan Javor, Kari Heiskanen

"Mineral floatability is a function of mineral properties such as size, liberation, shape, surface composition, reagent coverage on the surface, level of oxidation on the surface, etc.; pulp chemistry: Eh, pH, temperature, DO, reagent type, etc. and hydrodynamic conditions: turbulence, gas dispersion, D32, etc. The biggest challenge with predicting the floatability of an ore is finding a measurement method that would take into account the effect of all of the variables on the floatability of the minerals. The particle-bubble attachment time measurement is considered as an alternative way to predict the floatability of minerals instead of using expensive and time consuming flotation tests because it takes into considerations both surface chemistry and hydrodynamic parameters present in the froth flotation. The existing particle-bubble attachment time method is suitable for studying some significant flotation variables such as pH, temperature, particle size, reagent type and concentration. Nevertheless, measuring the attachment time as a function of collector coverage or surface oxidation is much more challenging. This paper introduces a new induction timer designed to measure the effect of galvanic interaction on the attachment time of particles. The new setup employs an array of needles to create 6 bubbles with the average size of 2.1 mm. Imaging system is used to set the initial distance between the particle bed and the array of bubbles. The attached particles are collected and sent for further analysis. One of the main benefits of the new experimental setup is that the attached particles can be physically separated from the unattached particles therefore many of the particle properties can be analysed and evaluated.INTRODUCTIONSchumann (1942), and Tomlinson and Fleming (1963) proposed that the flotation rate constant is dependent on four probabilities: particle-bubble collision (PC), particle-bubble attachment (PA), particlebubble aggregate stability in turbulent environment (PE), and particle-bubble aggregate stability in froth phase (PF). Any change in any of these probabilities will affect the overall flotation rate constant. All of the above-introduced probabilities are influenced by stream- variables and mechanical variables but their extent of influence vary. The stream variables are particle properties that have been shown to affect flotation rate constant such as size, shape, liberation, mineralogical composition, reagent surface coverage, surface oxidation, etc. The mechanical variables are the direct and indirect properties of the flotation unit affecting flotation rate constant such as cell and impeller geometry, air rate, impeller speed, air holdup, bubble size, residence time, hydrodynamics, froth depth, etc. For instance, the probability of collision is affected by many mechanical variables (bubble size, hydrodynamics, bubbles holdup etc.) but only one stream variable (particle size) has a major effect on it. On the other hand the probability of attachment is affected by much more stream variables: particle size, shape, liberation, surface composition, particle surface oxidation and collector coverage, Eh, pH, etc."

Jan 1, 2016

By G. D. Hibbard

New regions of material property space can be accessed by combining microstructural design at the nm-scale with architectural design at the µm- or mm-scale. In the first case, large strength increases associated with grain size reduction to below 50 nm have driven extensive research efforts into the development of nanocrystalline materials [e.g. 1-3]. For many potential structural applications, however, the density of a nanocrystalline material is just as important as its strength. In fact, reducing the density is more important than increasing the strength for certain weight specific materials performance indices, and it is especially critical for applying structural nanomaterials in the aerospace and automotive sectors. We have developed a new type of structural nanomaterial wherein the effective density of the parent metal is reduced by more than an order of magnitude by incorporating an internal periodic cellular architecture of open space. In one example a low density cellular nanocrystalline material was created by electroforming nanocrystalline Ni around a rapid-prototyped acrylic photopolymer microtruss, Figure 1 [4]. Microtruss materials have periodic cellular architectures that are specifically designed to undergo stretching-dominated deformation as opposed to the bending-dominated deformation in conventional open cell metal foams [5,6]. This new cellular nanocrystalline hybrid material combined the structural efficiency of microtruss architectures with the ultra-high strength that can be achieved by grain size reduction to the nm-scale. Although it played a critical role as a cathode support during the initial stages of nanocrystalline electrodeposition, the polymer core did not contribute significantly to the inelastic buckling resistance of the composite metal/polymer struts and it may therefore be desirable to remove it post deposition by means of chemical dissolution or thermal decomposition [4]. Nanocrystalline electrodeposition can also be used to reinforce conventional cellular metals (Figure 2), creating new types of cellular composites such as metal/metal microtruss hybrids [7-10] and metal/metal foam hybrids [11,12]. For example, nanocrystalline Ni-Fe alloy sleeve thicknesses of up to 400 µm were used to reinforce an aluminum alloy microtruss that had been fabricated using a stretch-bend approach; the nanocrystalline sleeves had the effect of increasing the microtruss strength by up to a factor of twelve and the specific strength by a factor of three [8]. Alternatively, much thinner sleeves can be electroplated to produce structural coatings. In one example, a nanocrystalline Ni coating was designed to provide both corrosion protection and inelastic buckling resistance [10]. Because the ultra-high strength material was optimally located at the furthest distance from the neutral bending axis, only a thin coating of 50 µm was needed to double the inelastic buckling resistance of the 1.13 mm x 0.63 mm cross-section plain carbon steel struts [10]. Overall, the mechanical performance increase that can be obtained through electrodeposition depends on a complex interaction between the starting cellular architecture and the nanocrystalline sleeves.

Jan 1, 2010

By Carmen M. Neculita

Optimizing the long-term efficiency of passive systems for the treatment of highly contaminated acid mine drainage (AMD) is still a challenging issue. The optimization lies, among others, on improving the understanding and quantification of metal removal mechanisms in order to predict and enhance the stability of neo-formed minerals in secondary phases. Sorption, and precipitation as oxides-hydroxides, carbonates and bio-sulfides are likely the most important metal removal mechanisms involved in passive biotreatment of both slightly and highly contaminated AMD. In the same time, the preferred mechanism is precipitation of metal sulfides that are more stable and less pH dependent than carbonates and oxides-hydroxides. However, results of relatively recent research studies performed at laboratory scale showed that sulfides accounted only for up to 15% of total metals removed in passive bioreactors treating highly contaminated AMD dominated by iron (around 500 mg/L). Several approaches were used in the aforementioned studies to understand and quantify metal removal mechanisms in spent reactive mixtures collected from long-term (12-15 months) operating column bioreactors. These included: thermodynamic equilibrium modeling and various chemical extractions and mineralogical analyses (such as single and sequential extractions, acid volatile sulfides-simultaneously extracted metals determination, scanning electron microscopy, X-ray diffraction, and thermo-gravimetric analysis). The large proportion of metals (85%) removed by sorption and precipitation as oxides-hydroxides and carbonates suggests a relatively high mobility of the neoformed minerals, especially in low pH conditions. However, no data is available on the long-term stability of spent reactive mixtures and, on the implications for the sustainable management of this type of solid waste. Lately there is almost a consensus that the efficiency of highly contaminated AMD treatment is warranted by the use of multi-step passive systems, including both chemical and biological units, installed sequentially. However, very few well-documented studies are available on the design, construction, operation and long-term efficiency of these multi-step systems. A leading example is the tri-unit passive system installed on the abandoned Lorraine mine site (Québec) to treat a highly contaminated AMD dominated by iron (about 2,500 mg/L). This tri-step system (160 m3), consisting in two bioreactors separated by a unit filled with wood ashes, achieves to date a satisfactory removal of iron (up to 99%). Moreover, based on the data collected during laboratory testing, iron was mainly removed in the wood ashes unit (from around 3,200 mg/L to 10 mg/L) by precipitation of oxides-hydroxides as predominant mechanism and only partially by sorption. However, no data is available on the actual metal removal mechanisms in the field. Overall, theses multi-step systems are more complex and their long-term performance is even less predictable than the single unit systems. This paper presents the current knowledge and the research needs on metal removal mechanisms in passive bioreactors and multi-step systems for the long-term treatment of highly contaminated AMD. KEYWORDS

Aug 1, 2013

By R. E. Whitehead

"The earliest gold-mining operations in Ontario included the Cordova, Deloro, and Sophia mines north of Lake Ontario (see for example, Miller and Knight, 1914). The Cordova deposit, worked intermittently between 1891 and 1940, is reported to have yielded 22 774 oz of gold from 127 670 tons of ore averaging 0.19 oz/ton Au (Gordon et al., 1979). The property remained inactive for many years. In 1988, Gunnar Gold Inc. and Mill City Gold Inc. conducted a drilling program along the main Cordova vein. The present study is based on core from three drill holes through the Cordova gabbro (Fig. 1) and the carbonate alteration zones enveloping the quartz-carbonate veins. The relatively narrow width of the alteration zone and the availability of continuous sections of core provided an opportunity to conduct a compact mineralogical-chemical study of carbonate alteration and its effect on the redistribution of rare earth elements (REEs). Although the Cordova deposit is not of great economic importance, this study has revealed a number of interesting features which are interpreted as resulting from processes that may be applicable to gold-bearing quartz vein deposits elsewhere.The Cordova gabbro is layered in places, varying from fine-to coarse-grained and in color from light gray (feldspathic) to dark gray-green (amphibolitic). However, layering within the section of the gabbro considered here and which hosts the Cordova gold deposit is mainly textural rather than compositional. Samples of unaltered gabbro selected for analyses were medium- to coarse-grained; amphibolitic and anorthositic phases were avoided as being less representative of the general chemical composition of the gabbro.The Cordova intrusion has undergone regional metamorphism and consists dominantly of actinolite and saussuritized plagioclase with smaller amounts of granular clinozoisite and flaky chlorite; minor quartz, ilmenite, titanite, apatite, and calcite are typically present. Some actinolite grains have irregular rims of pleochroic green hornblende that is optically continuous with the actinolite. Both chlorite and clinozoisite occur in two modes. Chlorite occurs as an alteration product of actinolite and as separate grains in the groundmass. Clinozoisite and albite occur together as regional metamorphic alteration products of original plagioclase crystals and as discrete grains in the matrix.The main gold-bearing quartz-carbonate occurrence occupies an easterly trending, steeply southward-dipping carbonatized shear zone up to 15 m wide and 700 m long. The gold occurs principally in irregular lenses and veins of quartz-carbonate up to a foot thick occupying the lower part of the schistose shear zone within the gabbro. This zone consists of chlorite, biotite, and carbonate and is flanked on both sides by chlorite-carbonate schist lacking biotite. The chlorite-carbonate schist in turn grades into regionally metamorphosed actinolite-bearing undeformed gabbro.The dominant sulfide within the quartz-carbonate lenses and in the surrounding schist is pyrite which occurs mainly as small cubes (1 mm to 2 mm) generally making up less than 10% of the rock. Rare blebs of chalcopyrite are present. Magnetite is common throughout the schist."

Jan 1, 2001

By B. Krysa, A. Mitsui, T. L. Koehler

The presence of arsenic, even at trace levels, in copper processing leads to the risk of arsine gas where reducing conditions exist that can reduce any soluble arsenic present to its minus 3 valance state and generate hydrogen gas from the solutions. This risk is understood to be significant within the industry, particularly in the case of liberator cells where copper concentrations in solution are drawn down to remove the copper and impurities from the bleed stream. However, there are numerous other potential areas where the arsine gas risk exists including contact between active metal such as iron, zinc or aluminum and arsenic bearing solutions in pyrometallurgical dust treatment, refining and acid plant operations. In turn this breaks down to different mitigating strategies dependent on the risk scenarios considered; for example, how this risk interacts with Change Management, Confined Space Entry or Contractor Training. The ISO 45001 is an emerging standard meant to improve employee safety, reduce workplace risks and create safer working conditions all over the world. The paper shows a case study of how a framework consistent with the tenets of ISO 45001 has been used to provide a consistent assessment of this risk for a wide variety of copper operations and the expectations of how this may be used to effectively improve the effectiveness of barriers against future incidents. INTRODUCTION Arsine gas (AsH3), first identified in 1775, is a highly poisonous, colorless, non-irritating gas with a mild garlic odor that is heavier than air. Arsine gas is generated when reducing conditions are generated in the presence of hydrogen ions and arsenic compounds. Present permissible occupational exposure limit as defined by Occupational Safety and Health Administration (OSHA) as an 8-hour time-weighted average is ppm (0.2 mg/m3). Further increasing the risk, conventional respiratory protection utilizing half or full masks with gas filtering cartridges provide minimal protection against arsine. In copper processing plants, the risk of arsine generation is well understood in liberator cells, which are part of the electrolysis process. In this process, solutions containing copper are reduced in copper concentration removing not only the copper, but also arsenic, antimony and bismuth. The residual solution can be bled to control impurity build-up in the refinery. In this process, if the copper concentration is reduced to the point where As3- can form, arsine gas is generated via reaction with hydrogen ions generated during the plating process. The copper concentration at which arsine gas is generated depends on copper concentration at the cathode surface, the temperature and the acidity. As a result, there is always a risk of arsine gas generation in the liberator cell. Process measures to minimize or mitigate arsine evolution are tight control on final copper concentration, warm temperatures, fresh feed distribution and internal acid recirculation rates. Operator exposure protection measures are hooding, ventilation, arsine detection systems, restricted personnel access, automated shutdown controls and pre-defined evacuation plans.

Jan 1, 2019

By X. J. Guo, H. X. Li

The biological heap leaching at Zijin copper mine is currently the largest industrial application of biohydrometallurgy and one of the earliest mines to develop biological metallurgy technology in China. Since 2001, through demonstration plant and different expansion stages, it has been developed to produce 20,000 tons cathode copper annually. However the ore grade of heap leaching gradually decline year by year with copper grade average at around 0.23% now. The copper mineral is mainly the secondary copper sulfide, and symbiosis with higher content pyrite, the sulfur and iron contents is about 15 times as copper in the ore (S/Cu>15), so the acid and iron treatment of leachate, extractant and acid mine waste water is a great challenge. Combined with located in the rainy area of southern China, huge amounts of over 10 million cubic meters acid drainage need to be treated every year, so the recovery of copper from waste water as well as low grade ore is one of the characteristics of this process. This paper in general introduced the operation of heap leaching, solvent extraction, stripping and electrowinning process in Zijin mine, summarized the experience of production and technical practice, and discussed the methods of heap construction, leaching rate enhancement, third phase formation and treatment, mine effluent treatment, resource recovery and heap slope reclamation. INTRODUCTION The Zijinshan gold and copper mine is located in southeast China, Fujian province, and the typical distribution characteristics of gold and copper deposit are "upper gold and lower copper", both are large porphyry deposits. By the end of 2017, the accumulated proven reserves of available gold resources were 318 tons, copper are 2.15 million tons, and the prospective reserves of copper are more than 5 million tons. The gold mining is open pit while the copper mining is open pit and underground. The cut-off grade of gold mining is 0.15 g/t, and that of copper mining is 0.15%. A large number of very low-grade gold ores (grade <0.30 g/t) adopt a combined ore-dressing process of “partial crushing/ore washing/reprocessing/heap leaching (overlying)/ carbon slurry" independently developed, while flotation and bio-heap leaching are adopted for copper mine. Among them, copper ore with more than 0.25% copper content enters the flotation system to produce copper concentrate, while that of less than 0.25% copper content is processed by bio-heap leaching, solvent extraction and electrowinning. Zijin biological heap leaching plant is the first mine in China to achieve the scale of annual output of ten thousand tons of cathode copper since 2006. In 2008, the solvent extraction and electrowinning system were expanded, now with a daily processing capacity of 20,000 tons of ores, an annual output of 20,000 tons of cathode copper, the expanded plant is the largest biohydrometallurgy plant in China.

Jan 1, 2019

By L. M. Nolan

Chemical Mechanical polishing, or CMP, has four components in the polishing interface; namely the polishing pad, the polishing fluid, the abrasives in the fluid, and the surface being polished. Understanding the way that load is transferred between these components is key to understanding and predicting wafer polishing. Numerous models of CMP and the different load sharing arrangements which occur between these four components have been proposed. These include models by Luo and Dornfield (1) and Fu, Chandra and colleagues (2). Other models focus on just one or two loading arrangements, such as `hydroplaning&apos; when all load is transmitted by the liquid in the case of Runnels and Eyman (3). In this work, we consider four possible combinations of load transfer from pad to wafer in CMP. These combinations are listed below in Table 1 and are illustrated schematically in Figure 1. It is important to note the role of scale in this process; the illustrations in Figure 1 are drawn to scale from actual profilometer scans obtained over a length of 5 with particles of 150 nm in diameter. Table 1 - Load Transfer Combinations Load Regime Illustration Wafer - Fluid - Pad Figure la Wafer - Particles/Fluid - Pad Figure lb Wafer - Particles - Pad Figure lc Wafer-Pad Figure ld The interest in load-sharing between the system components is because some configurations are anticipated to achieve higher polishing rates than others. For instance, in hydroplaning, or hydrodynamic lubrication, the lateral force exerted on the polished surface by the particles can be no higher than the fluid shear force developed in the gap between the pad and wafer. This is not expected to generate high rates of material removal from the surface (4). For contact that is largely between the pad and the wafer, such as the regime depicted in Figure ld, the lateral load is transmitted to the wafer by the pad and is hence limited by the pad&apos;s low rigidity. This regime is not expected to generate high material removal rates either. The highest levels of material removal are expected from regimes where the mechanical load exerted on the system is transmitted by the particles, in combination with either the fluid or the pad. Under what circumstances does the polishing regime change from one to another? What must occur in order to initiate or terminate contact between the system components? This is an ongoing question in the field of tribology, and has traditionally been viewed in terms of the pressure and velocity applied to the system, the thickness of the fluid film between the two surfaces, and the roughness of those surfaces (4). In this paper, we propose a new mechanism ? that the load on the particles may be sufficient to totally embed them in the polishing pad, at which point they are no longer able to transmit load and polish the wafer surface. This is illustrated schematically in Figure 1(d).

Jan 1, 2010

By H. Kalsi

"IntroductionRoyal Oak Mines Inc. is currently operating the Parnour No.1 mine located 10.5 miles east of Tinunins, adjacent to Highway No. 101. The operation produces gold ore at a rate of 3300 tpd. A variety of mining methods including bulk open stopes with ITH drilling, narrow vein panel, longhole and open pits contribute to produce 1.2 rnillion ore tons annually. The milling process requires water at an average rate of 1 rnillion gallons per day (700 U.S. gpm). The Pamour No.1 mine has historically had two major problems with regard to water control. The flfSt is the underground pumping system&apos;s inability to handle excess inflows of water generated from spring thaws. When the thaw takes place over a short period of time, large inflows of water are created which quickly migrate underground and sometimes result in the loss of the lower loading pocket. During the spring runoff the pumps at Parnour No.1 work continuously for 24 hours at a rate of 1200 U.S. gpm. The system was further taxed in 1991 when the Hoyle mine was connected to the Parnour No. 1 mine and whatever excess capacity the pumping system had for the spring runoff was taken up handling the inflow of mine water generated by the Hoyle.The second problem facing the mine was the lack of water for the mill during the severe cold winter months and dry summer months. At those periods inflows of water into the mine decline and the underground water normally pumped to the mill is severely decreased. In times of dry spells mill personnel increase the water diversion from the Three Nations Lake in order to meet water demands and maintain tonnage throughout. In essence, the mine was going through an annual ""feast or famine"" situation and a solution was needed.A number of options were considered but were narrowed down to the two most practical. The first was to modify the existing pumping system and increase its capacity in order to handle the excessive runoff during spring thaw. This involved a heavy capital expenditure and would have caused additional operating costs (i.e. increased hydro). The option chosen by the management was to divert the spring runoff into an unused section of the mine called The West End. A bulkhead placed at the 2400 level would provide an ultimate underground water storage potential of 150 million gallons. This would also cut down the load on the existing pumping system, save electric consumption in pumping and provide water storage which could be used for milling during the two dry spells. Pumping stored water from the mine could be scheduled for off peak hours to take advantage of lower power rates. To meet these objectives an underground bulkhead capable of withstanding high water pressure up to 2000 ft head (866 psi) was constructed in the spring of 1992. The bulkhead site was at the 2400 level at the Pamour No.1 mine. A major consideration was the requirement to construct a safe bulkhead and to minimize water leakage from the bulkhead.Figure 1 shows the bulkhead location."

Jan 1, 1994

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 Adrianus (Adrian) Halim, Tomas Bolsöy, Stina Klemo

"This paper describes the ventilation system used in underground hard rock mines in Sweden and Finland, which is very different than the system used in other major mining countries such as Australia, Canada, South Africa, and USA. The system utilizes auxiliary fans with Variable Speed Drive to distribute primary airflow to working levels instead of regulators that are commonly used in other countries. Descriptions about Ventilation on Demand in this system and ventilation regulations in Sweden and Finland are presented in this paper. This paper also quantifies a comparison between the Nordic and the system used in other countries. KEYWORDS Auxiliary fan, Regulation, Ventilation on demand (VOD), Variable speed drive (VSD) INTRODUCTION The primary objective of an underground mine ventilation system is to provide sufficient airflow to dilute atmospheric contaminants that are produced by mining activities. Typically, a mine ventilation system consists of one to four primary fans, regulators to control primary airflow distribution to levels or panels, and auxiliary fans that distribute airflow to dead-end workings via ducts. Depending on circuit geometry, underground booster fans are sometimes used. This system is widely employed in many countries such as Australia, Canada, USA, United Kingdom, and South Africa. However, mines in Nordic countries such as in Sweden and Finland employ a different system, in which auxiliary fans instead of regulators are used to control primary airflow distribution, as well as distributing airflow to dead-end workings. This paper describes the general layout of Nordic ventilation system, its VOD and heating system, its comparison with the system that is widely used in other countries which is referred as “Non-Nordic ventilation system” in this paper, and the current ventilation regulations in Sweden and Finland. The comparison was quantified by doing simulations using Ventsim Design 5 software. DESCRIPTION OF NORDIC VENTILATION SYSTEM General System Layout Figure 1 shows a schematic of typical ventilation system used in Swedish and Finnish mines. The schematic was drawn by Franzen, Myran, Larsson, and Rustan from Stiftelsen bergteknik forskning (Swedish Rock Engineering Research Foundation) and the English translation is written next to the Swedish terminologies. The system is generally a push-pull system. None of Swedish and Finnish mines use exhaust system. As shown in Figure 1, the primary airflow is distributed to levels by auxiliary fans that are bolted to a bulkhead located in the access to intake raise/shaft. These intake fans are attached to ventilation ducts that distribute airflow to all working faces. The exhaust auxiliary fans, which are not attached to any ducts and also are bolted to a bulkhead located in the access to exhaust raise/shaft, then flow the exhaust airflow into the exhaust shaft. A small quantity of this exhaust air is directed into the ramp, which acts as another main exhaust airway. The deepest level is also equipped with booster fan(s) that draw air from the intake shaft/raise and direct it into the ramp to provide fresh air in the ramp. There is no free split in this system. Figure 2 shows a schematic of a level ventilation layout in the Nordic system. Figure 3 shows a photo of these auxiliary fans."

Jan 1, 2019