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By J. Oliver, B. Wilson, 10

"Red Dog is one of the world’s largest zinc producing mines and a significant producer of lead and silver. While some of the operational challenges relate to the mine’s location in a remote part of Alaska, there are also challenges in the mineral separation process. Recent processing upgrades have included the installation of two M5000 IsaMills to improve the zinc regrind circuit and of one Jameson Cell to improve the pre-flotation circuit. Relatively fine grinding is required to achieve efficient mineral separation at Red Dog, but while fine grinding aids in the production of high grade concentrates, it also creates challenges in the efficient recovery of fine minerals. Over half of the lead lost to the lead circuit tailings, as well as almost half of the zinc lost to the final tailings, is in the finer size fractions (<10 µm). The high pulp density experienced in the pre-flotation and lead rougher circuits also causes significant entrainment loss of lead and zinc to the pre-flotation concentrate and of zinc to the lead concentrate. To help improve fine mineral recovery, Red Dog has recently plant tested and installed magnetic conditioning technology in the flotation circuits. The equipment aggregates fine paramagnetic sulphides and selectively increases their flotation recovery. This paper outlines the testing and positive results at Red Dog.INTRODUCTIONRed Dog is the largest zinc mine in North America and represents approximately 29% of North American zinc concentrate production. The mine is located in northwest Alaska, approximately 100 miles north of the Arctic Circle and 50 miles inland from the Chukchi Sea. The remote location makes the mine accessible only by air or seasonally by ocean-going barges. Red Dog is a partnership between NANA Regional Corporation Inc, the land owner and Teck Alaska Incorporated, the operator of the mine. Red Dog Ore and Mineralogy The Red Dog deposit is a rich sedimentary exhalative (sedex) zinc-lead-silver deposit. The main minerals present in order of abundance are: quartz, sphalerite, pyrite, barite and galena. Production began in 1989 with mining of the Main deposit. Between 1989 and 2012, approximately 55.6 million tonnes of ore with an average grade of 20.5% Zn, 5.6% Pb, and 97 g/t Ag was mined from the Main deposit. Subsequent to the development of the Main pit reserve, two additional open pit deposits were discovered: Aqqaluk and Qanaiyaq. The first blast in Aqqaluk occurred on June 4, 2010. As of December 31, 2014, the Aqqaluk deposit reserve is an estimated 45.7 million tonnes of ore with an average grade of 15.1% Zn, 3.9% Pb and 71 g/t Ag."

Jan 1, 2016

By Francisco Tamargo

"This paper refers to a new process for the full recovery (above 98%) of zinc and other valuable metals (Pb, Ag, Au) accompanying zinc in its concentrates while iron is separated as a marketable product free of any harmful impurities and suitable for sale to the cement industry. The process has been especially designed to achieve the following goals: maximizing the recovery of zinc and those valuable metals that are present in most zinc concentrates (lead, silver, gold); allowing the recovery of other valuable metals such as indium or germanium that could be present in some zinc concentrates; avoiding or minimizing the need for residue disposal in safe areas as it is the common practice nowadays in the zinc industry. Finally, the process achieves the previously related goals at a competitive cost, in line with the existing more efficient operations while using conventional equipment working under atmospheric conditions. The new process can be easily implemented in those smelters using the jarosite or goethite processes as well as in those smelters using direct leaching as a part of the processing facilities. A study has already been conducted on the use of jarosite in the cement manufacturing process, in the raw mix preparation as well as for a set controller. It shows that use of clean jarosite in the manufacture of ordinary Portland cement can result in improved quality of the clinker. Jarosite showed compatibility to influence the burning characteristics of the cement raw mixes. The presence of SO3 bearing mineral phases such as natrojarosite showed the suitability of clean jarosite for use as a set controller in cement production.INTRODUCTION Iron is by far the most abundant impurity accompanying zinc in its concentrates, and consequently, iron residue disposal is a major environmental concern for many zinc smelters in the world. This is especially true during the last 30 years when environmental restrictions became more and more strict. Much research work has been carried out during the last 30 years to try to find a solution for this important issue, but with little success. Only the jarofix process represents a partial solution to this problem as jarofix is stored nearby the smelters with certain limitations as a non-contaminant residue. However, it does not avoid the need to dispose of an important amount of residues every year, representing an environmental liability. Meanwhile, the number of countries limiting or banning the disposal of iron residues generated by the zinc smelters is growing, so that the solution to this problem is becoming an urgent matter right now. At the same time, the lack of appropriate available land for iron disposal is also a major concern in many cases."

Jan 1, 2016

By R. Kong, X. W. Zhang

"High pressure (up to 50 MPa) isotropic compression tests were performed on marble, gneiss and sandstone specimens, of which represents the typical rocks, to investigate isotropic compression deformation characterization at room temperature. Experimental data with respect to volume change in loading and unloading are presented. An apparent volume change inflection point (sensitive stress) was observed at a lower pressure range for porous sandstone, indicating that the elastic-plastic behaviour arising in the isotropic compression stage, but for tight crystalline rock, the sensitive stress is much higher, not probed prior to 50 MPa. The irrecoverable volumetric strain in the unloading can be seen for sandstone as well. The anisotropic deformation was discussed in terms of the axial and radial strain measurement. It was found that essential features of mechanical anisotropy are the irreversible in deformation range. The deformation feature in the isotropic compression tests is a very complex issue due to the factors concerning the stress history, unloading in sampling and in-situ stress. INTRODUCTIONThe isotropic or hydrostatic compression condition that stress is compressive and has the same magnitude for all direction, which is expressed as s1=s2=s3 in principal stress space. A yield locus of soils subjected to the isotropic and anisotropic compression was widely known, for instance Cam Clay model, which forms a closure cap model. However, the similar isotropic compression yield curves for rocks, especially for hard rock, was not established due to the geo-stress in the rock mass is much less than the socalled isotropic yield stress. And hence, the previous experimental studies on the hard rock have been made tend to on the stress-strain-strength relationship (Handin & Hager, 1957; Edmond & Paterson, 1972; Menéndez et al., 1996). Perhaps the yield curves related to an increasing in the hydrostatic pressure is not forced to construct for hard rocks, undoubtedly, the deformation data in this processing is valuable to assess the bulk module (Fabre & Gustkiewicz, 1997), anisotropy (Niandou et al., 1997), field stress (Hakala et al., 2007). Furthermore, many rocks store much elastic volume strain energy link to the hydrostatic pressure exerted, following a deviator stress applying, to drive the unstable brittle fracture for hard rock. Latter was studied to interpret the incentive of rock burst. However, volume change characteristic in the isotropic compression behaviour have been scarcely studied experimentally to our knowledge, less application related to the fracture in hard rock as well. The mechanical behaviour in the isotropic compression of hard rock is notably affected by the original structure anisotropy, stress history and stress condition before sampling, all of which need reasonably assess. It is also a useful understanding to study isotropic compression that elastic and plastic volume change in the isotropic loading and unloading, and the micro crack evolution throughout it."

Jan 1, 2015

By Z. Zhong, Y. J. Hu, D. Elsworth

"A standard triaxial apparatus was set up to perform sliding tests. Slide-hold-slide (SHS) tests, with hold times from 10 to 3000s and at invariant effective confining pressures of 1MPa, incremented to 3MPa and then 5MPa, were conducted on both dry and wet split limestone samples to study the evolution of fracture friction under holding and sliding processes. Limestone fracture shows time-dependent behaviors in SHS tests, friction drops (creep) gradually during hold periods, and recovers (healing) after re-slide. Frictional healing and creep increment are proportional with hold times, while are inversely proportional with effective confining pressures. Moreover, the wet sample displays heavier healing than the dry one, whereas creep increment of the wet sample is lighter than the dry one. We conclude that the evolution of limestone fracture friction is significantly controlled by mechanisms of fluid-rock interactions.INTRODUCTIONThe evolution of the mechanical properties of fractures and faults exerts a critical influence on the behavior of fault zones and on the earthquake cycle. In addition to natural processes, this behavior has broad implications in induced seismicity and influences oil and gas production, the sequestration of carbon dioxide and the disposal of nuclear wastes.Fault friction is typically measured by slide-hold-slide experiments, where fault gouge is deformed at low velocities (<0.1 mms-1) and over small shear offsets (<30 mm) (Dieterich, 1979, 1981; Mair et al., 1999; Marone, 1998, 2004) or alternatively with rotary-shear devices, that enable higher slip velocities over very large shear offsets (Hirose & Shimamoto, 2005; Mizoguchi et al., 2006). Typical experiments (Dieterich, 1972; Marone, 1998; Olsen et al., 1998; Bos & Spiers, 2002; Tenthorey et al., 2003) show significant creep during the holding period and frictional healing (strength gain) upon reloading. Fault creep and healing are often time-dependent with longer hold times typically resulting in greater magnitudes of creep and in strength gains (Tenthorey et al., 2003; Yasuhara et al., 2005; Carpenter et al., 2012). Frictional healing may arise from growth of grain contacts (potentially induced by pressure solution) (Rabinowicz, 1951; Dieterich & Kilgore, 1994; Yasuhara et al., 2005), strengthening of the contacts (correlates with cementation) (Hickman & Evans, 1992; Beeler et al., 1994; Karner et al., 1997; Olsen et al., 1998; Muhuri et al., 2003; Tenthorey et al., 2003; Li et al., 2011) or via other changes in state including changes in porosity and rearrangement of force chains. Experimental evidence (Dieterich, 1979, 1981; Johnson, 1981; Marone & Scholz, 1988; Marone, 1998; Saffer & Marone, 2003; Moore & Lockner, 2007; Ikari et al., 2007; Ikari et al., 2011; Carpenter et al., 2012) quantifies both the mode and magnitude of second-order changes in the velocity-dependent (including velocity- strengthening and weakening) evolution of frictional strength."

Jan 1, 2015

By E. Ghali, W. Zhang, M. Robichaud, G. Houlachi

"The performances of six mesh and one plate dimensionally stable anodes have been investigated in the zinc electrolyte at 38oC and compared to conventional silver-lead alloy anodes. The electrochemical behaviour of these anodes was studied by cyclic voltammetry, galvanostatic, potentiodynamic, open circuit potential and electrochemical noise measurements. It was observed that at the end of 24 h of galvanostatic polarization at a current density of 50 mA/cm2 in the zinc electrolyte at 38oC, the mesh V had the lowest potential followed by mesh III, mesh VI, plate-V, mesh IV, mesh I, and mesh II electrode. When the potentials for mesh I, mesh III, mesh IV, VI and plate-V, mesh V were compared with that of the Pb-0.7%Ag anode (1920 mV/SHE), they were found to be 6%, 16%, 7% 10%, 17% lower, respectively.After 16 h decay following the 24 h polarization, the mesh III had the lowest corrosion rate (51 mS/cm2) among the seven specimens, followed by the plate-V, mesh V, VI, IV, I and II (221 mS/cm2). For the fresh surface as received, after two hour open circuit potential measurements, mesh I had the most positive (noble) potential (1382 mV/SHE) among the seven electrodes followed by the plate-V electrode, mesh V, mesh IV, mesh VI, mesh III and mesh II (751 mV/SHE). After consecutive series of electrochemical testing, mesh V, mesh III, mesh VI had lower corrosion rates (~ 343± 6 µA/cm2) than that of mesh IV, mesh I, plate-V, and mesh II (~ 644± 143 µA/cm2).INTRODUCTIONThe continued use of lead based anodes in the electrowinning process has been mainly due to their relatively low cost as compared to other materials (Msindo, Sibanda, & Potgieter, 2010). They have good electrical conductivity, stability during electrolysis, very good corrosion and wear resistance (Ivanov, Stefanov, Noncheva, Petrova, Dobrev, Mirkova, Vermeersch, & Damaerel, 2010). However, lead-based anodes give their high energy consumption and low corrosion resistance for electrowinning. To overcome these problems, Dimensionally Stable Anodes (DSAs) has been chosen (Msindo et al., 2010). Comparing to the traditional lead anodes, the DSA anodes have many advantages: (1) no contamination; (2) no loss of anode material, pure electrolyte; (3) lower gas bubble effect through special anode designs; (4) use of higher current densities (Msindo et al., 2010). The success of DSA anodes has been attributed to their chemical stability that allows them to maintain nearly constant dimensions during the life of the anode even when operating in environment with low pH values. Due to their stability, the consumption rates of DSA are as low as 1 mg/amp.y (Alfantazi & Moskalyk, 2003)."

Jan 1, 2016

By N. P. Hristov, L. V. Kuzev

"A simulation study of the fill factor of a drum mill was conducted using two types of grinding media: balls and tetraballpebs. Each medium was composed of equal number of grinding bodies having equal weight but different shape. The difference was in the density of the two grinding media.Standard spherical grinding bodies (balls), which were used in practice, and equivalent (by weight) grinding bodies in the shape of spheroidal tetrahedrons (tetraballpebs) were studied. The simulation was conducted in a module of a 1.2 m diameter drum mill, whose length exceeded twice the diameter of the corresponding grinding body. The grinding balls had a few percent (3 to 6) higher fill factor than the grinding tetraballpebs in the same shell.Vertical sections of the grinding medium were taken across the studied shell module along its axis. They showed the location of each grinding body in the fill volume. Defects in the packing of the grinding bodies and the contacts between them and with the inner surface of the shell were also show. Differences in the contact points of the balls and the tetraballpebs were noted. For the tetraballpebs, we observed contacts both between the spherical surfaces and with the edges and vertices of the grinding bodies.The topography of the grinding media is shown visually.INTRODUCTIONThe grinding balls have been studied extensively. It is the traditional grinding medium in ore processing industry. The parameter characterising the grinding medium is its density. It is the mass of the grinding medium per unit of volume. The parameter may have a range of values. It depends on:• The physical properties of the material they are made of;• The shape of the grinding body;• The dimensions of the body;• Size distribution of the grinding bodies in the grinding media.In comminution, the fill factor of the mill shell is the process parameter that is representative of the grinding medium. Some publications have indicated the disadvantages of ball grinding. For example, Wu, Xiang, Glen, and Duan (2008) note in the analysis of the mechanism of grinding that the balls are not the best grinding media and disadvantages include among others:• Easy over-grinding resulting from the point contact between the balls;• Attrition surface is the smallest in the grinding media;• Poor selective exposure of minerals resulting from the accidental and often significantly more powerful impact on the feed;• The grinding efficiency drops as the fineness of grind is increased;In the Handbook of mineral processing, Polkin (1950) noted that the shape of the grinding bodies determines the volume of the void spaces therein and therefore the shape affects the grinding efficiency per unit of volume of the grinding medium."

Jan 1, 2016

By Martin G. Wafforn, George J. Greer

"In early 2014, Pan American Silver Corp. commenced a project to construct a second shaft at its La Colorada underground mine located in Zacatecas State in Mexico. This was the first new shaft development project undertaken by Pan American Silver in the history of the company. The existing mine workings permitted the use of a raise bore for the excavation of the shaft opening, utilizing directional drilling in poor ground conditions to advance the pilot hole. This was an interesting project with many challenges and successes, and numerous lessons learned. The project was successfully completed in 2016, with the shaft operating at full mine production capacity within weeks. INTRODUCTION The La Colorada Mine has a long history dating from at least the 1920’s. The mine is located in the northern portion of Zacatecas State in Mexico in the Chalchihuites area, an approximate two hour drive south from the city of Durango. Pan American Silver acquired the La Colorada mining concessions in 1998, with mine production at that time consisting of only a few hundred tonnes per day. Over the course of the following 15 years, with the addition of an oxide processing plant and some expansions to the sulphide processing plant and site infrastructure, production was eventually increased to a maximum of 1,250 tonnes per day, all of it generated from conventional cut-and-fill stopes in the underground mine. An existing shaft was used to hoist the ore to surface, and over the years Pan American Silver had deepened the shaft and had changed out some of the hoisting plant. However, there were historic workings in close proximity to the shaft that made it unsuitable for personnel transport, and shaft maintenance was becoming difficult. The shaft was increasingly considered as a risk to the long term production continuity, and also to further deepening of the mine. In addition, expanding the capacity of this old shaft was not possible. Drilling of the La Colorada mineralization at depth commenced in 2010, with very good exploration success. As a result, by the end of 2013, the mine reserves had grown to a total of 6.5 million tonnes of high grade ore (388 g/t Ag, 0.36 g/t Au, 2.4 % Zn, and 1.3% Pb). Furthermore there were more (and higher grade) sulphide reserves than oxide reserves at depth. Following the completion of an internal study in 2013 with assistance from one of the international shaft contractors, a decision was made to increase the production from the mine with the construction of a new shaft and a new sulphide processing plant, a project with a calculated payback of 2.2 years."

Jan 1, 2019

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

"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 Jeremy D. Glennon

A number of exciting challenges, opportunities and unanswered questions remain concerning the chemistry, geochemistry, medical, industrial and analytical applications of gold and its compounds.1 Among the challenges presented is the need for extraction methods for gold, at process and indeed analytical scales, that avoid the use of organic solvents or hazardous reagents. Extensive environmental contamination has occurred in the Brazilian Amazon from the mercury-gold amalgamation process. Major problems with the cyanidation process are the relatively slow dissolution kinetics and the health and environmental problems associated with it. With innovation and technical changes, a cleaner way for solubilising and extracting this versatile precious metal can be found. This cleaner future may lie in supercritical fluid science. Interestingly, geochemical studies of gold solubility and speciation in hydrothermal solutions have highlighted the importance of supercritical conditions and sulphur species in the deposition of gold in natural environments. Gold solubility in supercritical brines is far higher than previously supposed while interest in the high content of carbon dioxide in the mineralizing fluids is growing. 2,3 In recent years, also, there has been a tremendous amount of attention paid in analytical chemistry and in industry, to the area of supercritical fluid extraction (SFE) of organic and metallic species. 4 This interest has been fuelled by the introduction of more stringent environmental control and the subsequent need for cleaner solvents and processes to replace established ones. Carbon dioxide is the substance of choice for SFE because of its moderate critical parameters (Tc = 31oC, Pc = 73atm), inertness and availability in pure form. It is an excellent extraction medium for non-polar and moderately polar species, and with careful ligand design, it can be used for analytical and process scale metal extraction from solution and solid phases. In general, the dissolution of a metal in supercritical CO2 (sc-CO2) requires that the charge on the metal ion be neutralised on complexation by the ligand or through a form of ion pairing (analogous to the situation where metal ions are extracted into organic solvents, such as methyl isobutyl ketone or toluene). In this way, then, many chelating agents such as dithiocarbamates and hydroxamic acids have been utilised to extract metal ions such as Cu and Fe into sc- CO2, the process greatly enhanced by the use of fluorinated ligand derivatives. 5-7 In this review paper, we show that supercritical conditions and complexation with sulphur donor atoms, important in the natural deposition of gold, can form the basis of a clean, eco-friendly chemistry for gold extraction. Analytical scale extractions of gold from a variety of matrices can be conveniently used to assess the efficiency of any new proposed ligand.

Oct 1, 2003

By D. E. M. Plaumann

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

Jan 1, 1988

By W. B. Evans, D. P. Taylor

"ANFO has two major disadvantages as an explosive low density and no water resistance. ANFO can be blended with more dense, water-resistant explosives to form ""Heavy ANFO"". The use of emulsion or slurry explosives as a matrix is an attempt to correct deficiencies on the basic ANFO while maintaining cost effectiveness.This paper will deal with four main topics.1. Product description, general properties, strengths and weaknesses.2. Technology of ""Heavy ANFO"" , the emulsion matrix concept and the range of products.3. Relative performance and economics in various rock types.4: Current status in Canada and worldwide, and the expectations for future use.IntroductionANFO has been in use for roughly three decades and remains the explosive of choice where conditions are suitable. ANFO has not become the universal explosive - despite its undoubted superiority in cost effectiveness - because it has serious flaws as an explosive. The most serious of these flaws are the low density and absence of water resistance in ANFO. Low density implies a low concentration of explosive energy in a borehole. The amount of energy available can be increased by. adding an energetic material, generally aluminum, or by drilling more holes. Both solutions add considerably to the over-all costs. The absence of water resistance in ANFO is inherent and can only be overcome by changing the fundamental nature of the product. The history of research in the explosives industry since the advent of ANFO is mainly the story of the various attempts to overcome one or both of ANFO&apos;s shortcomings.Many technologies and products have been developed to compete against ANFO, among them slurries, emulsion s and the so-called Dense ANFOs (ANFO with an extra nitroalkane liquid sensitizer). None of them was cost-competitive with ANFO so their market penetration has been limited to wet holes or specialized applications. The exception to this result is Heavy ANFO. The aim here was to have the best of both worlds, the high density and water resistance of emulsions or slurries plus the low cost of ANFO. The concept is very simple. ANFO is about 50 % air, roughly 30% in the prills and 700,10 between the prills . The air in the prills provides the necessary sensitivity but the air between them is waste volume. If one fills this volume with a high density, waterproof product, the resultant mix will be denser and more water resistant than ANFO and with very little added cost. This basic concept was invented by B. Clay in the 1970s and is covered by Canadian Patent No. 1 115 959. The water resistance and/or density of the mix can be tailored by changing the emulsion level. Because Heavy ANFO is relatively inexpensive while providing good in-hole densities and water resistance, it can compete successfully against all other bulk products. It competes against ANFO and Al / ANFO by allowing pattern expansion and reducing over-all costs while it competes against emulsions and slurries on price."

Jan 1, 1987

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 Esau Arinaitwe

The most extensively used methods and approaches for characterizing and selecting frothers for plant application are based on measurements in clean water in the absence of solids and without consideration of physicochemical factors that are known to influence the properties and behavior of practical flotation froths, and hence metallurgical performance. These factors include aquatic chemistry, collectors, modifiers, particle shape and size, and pH. Understanding interactions between these factors and the role and relevance that they play in characterizing frothers and froth zone properties will greatly aid in the frother selection and evaluation process. Such an understanding will also greatly enhance the ability to create frother solutions that are sufficiently robust to plant variability (such as mineralogical, comminution, operational, physical and chemical). This paper reports results of a “systems approach” study designed to probe ways in which physicochemical factors a) influence frothing and bubble size in 2- and 3-phase systems, and b) confound correlation of fundamental measurements with flotation outcome. The concepts and insights gleaned from fundamental frother studies, together with empirical laboratory flotation tests and a thorough understanding of plant needs and challenges were used to develop and implement customized frother formulation at a Cu-Mo flotation plant. Plant metallurgical data and qualitative assessment of froth zone properties showed that a frother formulation developed using the approach described here provided marked improvements in froth characteristics (mineralization, mobility, stability, persistence). These improvements resulted in better circuit control and significant performance advantages, such as improved coarse particle recovery, without detrimental downstream effects. INTRODUCTION Successful recovery of value minerals of sufficient grade via the flotation process requires selective manipulation of the value minerals to render them hydrophobic using collectors and modifiers. The hydrophobized minerals then attach to air bubbles and the mineral-bubble aggregates are selectively transported through the froth zone into the launder. Frothers play a critical role in froth formation and more importantly in achieving and maintaining desirable froth characteristics (for example, stability, mobility, persistence) that are essential in facilitating the selective transport and recovery of value minerals. Fundamentally, frothers contribute to the transient stability of thin liquid films and so aid in reducing bubble coalescence rates in the pulp and froth zones, thus maintaining the small bubbles required for efficient bubble-particle attachment. Invariably, the action of frothers at the liquid/gas interface in the pulp and froth zones impacts bubble stability, water and solids transport to the froth, which in turn impacts froth mobility, froth carrying capacity, and the concentrate grade. Consequently, selection and evaluation of frothers for flotation plants requires careful consideration of the frother type, though the choice is limited to only two main classes of compounds in terms of fundamental chemistry (short chain alcohols and polyglycols). Even with the selection of what is thought to be the best frother for a given application, practical knowledge of the froth zone and experience in controlling its properties is essential for maximizing metallurgical performance.

Jan 1, 2019

By M. Coleman, M. E. Reed

The control of feed delivery quality to the flash furnace remains an unsolved problem. Many operations around the world still have limited control of feed quality to their flash furnace. Schenck Process and WorleyParsons continue to work towards the ‘holy grail’ of simple quality feed delivering both radial feed distribution and pulseless feed delivery. Ten years ago they joined forces to offer a system based on then Clyde Process RotoFeed system which could deliver ground level installation, online mixing of materials, pulseless feeding direct to the burner, with good radial distribution. Unfortunately, this proved to be too large a step for the Flash Smelting community and met with resistance from the process owners. Since then the community has developed a range of options some of which have been implemented over this period. Despite these improvements there are still areas of concern including safety issues associated with concentrate leaks and spills, uncontrolled variations in feed rate including massive flooding events, intermittent distribution improvement and set up of the burner continues to be trial and error and hard to repeat the success in each burner campaign. These continue to impact flash furnace performance globally on a daily basis. Following the community reticence to pursue the complete solution a number of interim steps are now available to pave the way to the ultimate feed system. The mixing of material has been installed and is delivering line blended materials on a daily basis. This paper will investigate the developments over the last 10 years and discuss the pros and cons of this incremental step process for overall furnace control. INTRODUCTION Schenck Process and WorleyParsons have continued to interact and observe changes in the flash smelting industry and to consider application of the Clyde RotoFeed now incorporated in the Schenck Process ProStreamTM range for delivery of feed directly to the concentrate or matte burner. Schenck Process pneumatic injection from the Clyde Process range, ProStreamTM, utilises two types of volumetric feeders to accurately control the feed of material into an injection pipeline. These are referred to as the RotoFeed and RotoScrew technologies and the choice of feeder type is dependent on the characteristics of the material to be handled. The RotoFeed is generally used for finer, granular and powdered materials which cannot be accurately handled with a screw such as pulverised coal and copper concentrate, whilst the RotoScrew is particularly suited to coarser lump materials, particularly abrasive of cohesive materials like zinc powder, electronic scrap and metal ashes. Both units can operate at high pressures and many systems work at 3 to 10 bar with the occasional system working at 20 or 30bar for carbon to gasifiers. For the Flash furnace the ProStreamTM units can be used to deliver each of the furnace feed materials into a single line, mixing the materials homogeneously and live to the burner.

Jan 1, 2019

By M. Pötsch, A. Kluckner, P. Söllner, W. Schubert

"The in situ block size is a key parameter in the geomechanical characterization of rock masses. It describes the fracturing of the rock mass and thus is a measure for the degradation of the rock mass strength. Several classification systems use the in situ block size. For instance, the Geological Strength Index (GSI) requires the “blockiness” and joint surface condition factor as key input parameters. The “blockiness” has recently been related to the in situ block volume (Cai et al., 2004). Due to the limited information about the internals of a rock mass, it is not possible to determine the in situ block size directly. Currently, the in situ block size is determined by calculations using oversimplified models, vague estimations or, due to the lack of relevant information, it is neglected. As remote measurement systems have become available for rock mass characterization, a more comprehensive record of discontinuities can take place. Measurements can be performed on exposed rock outcrops and in particular on tunnel faces and walls, delivering the location, orientation, spacing and persistence of discontinuities at an arbitrary number and locations. An estimation of the in situ block size at the same level of sophistication is still not available. This paper aims at examining the relationship between visible and measureable information at the rock surface, and the in situ block size. Three-dimensional block model simulations were performed using the block model engine of the distinct element code 3DEC. Initial investigations focussed on the minimally required observation area in order to obtain reliable block size distributions. Based on a representative volume element, a reference distribution of the block area at the observation area for a specific discontinuity system was determined. Subsequently, the size of the observation area was decreased stepwise and each new distribution was compared to the reference distribution. For all analyzed systems, it turned out that the mean block area must be smaller than 1% of the observation area to achieve a reliable block size distribution. Further simulations focussed on discontinuity systems with three non-persistent sets. The results were compared to the formula proposed by Cai et al. (2004). A transformation factor T was introduced which replaces the persistence terms in Cai et al.’s proposed formula. The factor describes the correlation between block volumes generated by persistent and corresponding non-persistent discontinuity sets. The simulations included the variation of the discontinuity set spacing, persistence, and orientation. The mean and additionally the 25%-, 50%- and 75%-quantiles of the block size distribution were analyzed. Examining all values for the transformation factor the distribution can be described by a power function with a negative exponent. The function depends on the persistence only. Finally, the results of random simulations were compared to the analytical formula using the transformation factor. The predicted mean and selected quantiles show a good agreement with the simulation results. It gives a comprehensive picture about the block size distribution in discontinuity systems with three non-persistent sets."

Jan 1, 2015

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 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 K Byrne, W. Hawker

The average grade of copper concentrate is steadily declining, forcing the industry to turn to lower grade and metallurgically challenging orebodies. When faced with a low grade, challenging ore, mining companies are often forced to choose between high recovery at low grade or low recovery at high grade. Hydrometallurgical processing of copper concentrates using metathesis reactions, which replace iron with copper in copper-iron sulphide minerals and increase the copper to sulphur ratio in copper sulphide minerals, provide an opportunity to upgrade copper concentrates and potentially allow high copper recovery for a high grade concentrate. Factors such as concentrate mineralogy, reaction time and temperature, and the concentrations of copper, chloride and acid in solution are reported to affect the reaction extent and mineral phases formed by metathesis. In this paper, a copper concentrate has been studied over a wide range of conditions to investigate the influence of these key parameters on the extent of copper metathesis reactions and product phases formed. Solution and solid phases are assayed with the main mineral phases identified by Quantitative X-Ray Diffraction. INTRODUCTION The minerals present in copper deposits determine the prospective processing options either: pyrometallurgical or hydrometallurgical. For copper oxide ores, copper is typically present as oxides, carbonate, chlorides, silicates and sulphates. These minerals are readily solubilised in acid and are usually treated through hydrometallurgical processes. The typical process route for copper oxides is to leach copper into aqueous solution, purify and upgrade the solution phase through solvent extraction and then recover copper as metal by electrowinning (Davenport, King, Schlesinger, & Biswas, 2002). Copper sulphide deposits may contain a multitude of copper bearing minerals, including chalcopyrite, bornite, enargite, covellite, digenite and chalcocite. The copper ore grade and treatment options are affected by the mineral type and distribution. Sulphide ores are typically treated through comminution and mineral flotation to form concentrates, followed by pyrometallurgical smelting. Hydrometallurgy can also be used to process sulphides depending on the mineralogy and impurity levels (Davenport et al., 2002). The trend of decreasing copper grade with increasing impurity levels are common across the mining industry and pose future challenges to copper processing. This trend has prompted research into hydrometallurgical options for upgrading copper concentrates. One promising hydrometallurgical option is using metathesis reactions. Metathesis reactions, sometimes also called double replacement or displacement reactions, are chemical processes involving the exchange of bonds between two chemical compounds or species. In this case, the metathesis reactions involve copper in solution exchanging with cations in the solid sulphide concentrate. The driving force for copper sulphide metathesis is the relative affinity of the exchanging cation elements for sulphide. Based on sulphide precipitation diagrams, the expected order of affinity for sulphide should be Hg2+ > Ag+ > Cu+ > Cu2+ > Bi3+ > Cd2+ > Pb2+ > Sn2+ > Zn2+ > Co2+ > Ni2+ > Fe2+ > Mn2+ (Monhemius, 1977). Hence, copper metathesis reactions are capable of increasing copper content and removing impurities such as iron, nickel, zinc and lead from copper concentrates.

Jan 1, 2019