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By G. Chi

Outline 1.Regional geologic setting 2.Host rocks and structures 3.Mineralization, alteration and paragenesis 4.Geochronology 5.Major and trace elements 6.Fluid inclusions 7.Carbon and oxygen isotopes 8.Conclusions

Apr 1, 2005

By M. W. Grabinsky

Better in situ performance data is required in order to advance the rational geomechanical design of cemented paste backfill systems. A few in situ test programs have been conducted, but these have suffered from too few measurement locations within the stope and have not used enough different types of transducers to fully capture the information of significance. In this paper the important geomechanical design aspects are listed, the underlying physical phenomena are described, and some previous in situ measurement efforts are reviewed. A new approach to in situ measurement campaigns is then described including measurement techniques, locations, and the logistical aspects of mobilizing transducers in an active mining environment. A work-in-progress is used to illustrate the principles and concepts involved.

May 1, 2007

By J. R. Kadambi, J. Furlan, Kucukal E.

"Hydrocyclones are used in various industrial applications, including mineral processing and tailings management, to separate or classify solid particles and/or liquid droplets from multiphase feed. The fluid flow in a hydrocyclone (used in mineral processing) was investigated numerically and experimentally using particle-image velocimetry (PIV) and refractive index matching techniques. Two turbulence models were used in the calculations: the k–? model and the Reynolds stress model. Particle image velocimetry measurements were taken at various spatial locations along the cyclone. Computational results were validated globally using pressure and flow-rate readings at the boundaries, and locally by comparing the PIV velocity-vector maps and profiles.RÉSUMÉ On utilise les hydrocyclones dans diverses applications industrielles, dont la minéralurgie et la gestion des résidus, pour séparer les particules solides et/ou les gouttelettes liquides de l’alimentation polyphasée, ou pour les classifier. Une étude numérique et expérimentale sur l’écoulement des fluides dans un hydrocyclone (dont on se sert en minéralurgie) a été menée à l’aide des techniques de vélocimétrie par images de particules (PIV, de l’anglais particle image velocimetry) et d’appariement des indices de réfraction. Deux modèles de turbulence ont été utilisés dans ces calculs, à savoir le modèle de turbulence k–? et le modèle aux tensions de Reynolds (RSM, de l’anglais Reynolds stress model). La valeur de la PIV a été mesurée à diverses positions spatiales le long du cyclone. Les résultats de ces calculs ont été validés globalement à l’aide de relevés des pressions et du coefficient d’écoulement d’une extrémité à l’autre de l’hydrocyclone, et localement à des coordonnées spatiales spécifiques en comparant les cartes et les profils du vecteur vitesse PIV.INTRODUCTIONA hydrocyclone is a piece of equipment used to separate and/or classify solid particles or liquid droplets from liquid solid or liquid-liquid feed. Versatile in application, hydrocyclones have been widely used in industrial applications including the food, oil, power generation, chemical, and mineral processing industries. The first hydrocyclone was introduced in 1891 by Bretney to purify various liquids (Bretney, 1891). Hydrocyclones have a relatively simple design with no moving parts, which leads to an efficient separation process."

Jan 1, 2017

By Abdelkabir Maqsoud

Different management options and rehabilitation strategies are available for mine wastes that produce acid mine drainage (AMD). One of them consists in building a cover with capillary barrier effects (CCBE) on the tailings to limit the availability of oxygen, an essential component of the chemical reactions leading to the generation of AMD. Prior to cover construction, a monitoring strategy must be established to evaluate the response and performance of the cover system. The key parameters to monitor should include matric suction and volumetric water content in the CCBE, oxygen flux (and/or oxygen concentration) through and below the cover, and the main meteorological parameters for the impoundment. In this paper, different techniques available to measure the targeted parameters are described and their main properties are given. Recommendations (based on existing cases) on the location and the number of monitoring stations and on the measurement frequency for each monitored parameter are also proposed.

May 1, 2007

By M. Fan, M. Mankosa, G. Luttrell

"Column and other non-conventional flotation approaches continue to provide benefits where conventional flotation processes have exhibited limited success. This is particularly evident with respect to flotation of ultrafine and coarse particles. Advancements in air sparging technology have provided the ability to produce finer bubbles than previously possible. The generation of finer bubbles coupled with the froth washing capabilities of a flotation column has demonstrated improvements in ultrafine particle recovery while simultaneously producing a high-grade float product. On the other side of the spectrum, the use of these sparging technologies while applying flotation fundamentals to a fluidized-bed reactor has increased the maximum floatable particle size up to and in excess of 3 mm. Benefits of higher recoveries and reduced reagent consumption have resulted from the use of these technologies. An additional benefit includes a reduction in plant grinding requirements. As such, this paper describes the application of Cavitation-Tube sparger technology and fluidized-bed flotation (HydroFloat). A review of data from industrial and pilot-scale applications is also provided.INTRODUCTIONConventional (mechanical) froth flotation is considered the workhorse of the minerals processing industry. This low-cost process makes it possible to selectively recover minerals by exploiting the difference in surface characteristics (e.g., wettability) of the valuable component and the waste material. Unfortunately, the upper particle size limit for flotation rarely exceeds 0.600 mm, and is typically maintained at less than 0.300 mm for most industrial applications. For many commercial base metal plants, the particle size range for optimum recovery is typically between 0.01 and 0.1 mm (Lynch et al., 1981). This finding is not novel. In fact, the effect of particle size on flotation recovery was studied as early as 1931 by Gaudin et al. who showed that coarse (+0.250 mm) and extremely fine particles (-0.010 mm) are more difficult to recover than particles of an intermediate size. The loss of recovery is due to inherent limitations associated with the flotation process, including bubble-particle adhesion and collision, detachment, buoyancy, and transport."

Jan 1, 2010

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama, Ryuji Oinuma

"Recently, the arsenic grade in copper ores has been increasing year by year. Arsenic bearing minerals (enargite (Cu3AsS4) and tennantite (Cu12As4S13)) coexist with typical copper minerals such as chalcopyrite (CuFeS2). The removal of arsenic from copper ores that contain relatively high arsenic content is essential due to its adverse impact on the environment and the resulting economic consequences. In this study, arsenic removal and dissolution behaviour of Cu and As from enargite (Cu3AsS4) were investigated using ferric sulphate (Fe2(SO4)3) solutions containing sodium chloride under high pressure oxidation conditions. The effects of numerous parameters including the concentration of Fe2(SO4)3 and NaCl, temperature, total pressure, leaching time and particle size were studied while other variables such as pulp density and stirring speed were fixed at 100 g/L and 750 rpm. The results showed that more than 60% of the copper was dissolved in a 0.1 M ferric sulphate solution under the high pressure oxidative conditions, whereas the leaching rate of arsenic was less than 20%. SEM observation revealed that elemental sulphur was deposited on the surface of enargite forming a passivation layer and the residue was agglomerated. Those phenomena may cause a decrease in the copper leaching rate. It was found that more than 80% of copper was dissolved in the ferric sulphate solution in the presence of 0.34 M sodium chloride under the conditions optimized, whereas the dissolution of arsenic in the mixture did not exceed 2%. XRD analysis of the solid residues from the leaching experiments used ferric sulphate and mixtures of ferric sulphate and sodium chloride showed arsenic to be present as scorodite (FeAsO4).INTRODUCTION In recent years, there has been an increase of the arsenic (As) content of copper ores due to the gradual depletion of high grade copper ores. In some copper deposits, arsenic occurs in the form of minerals such as enargite and tennantite. The removal of arsenic in ores containing copper is critical to the production of pure copper and the management of the environmental challenges caused by the smelting process. Due to the environmental risk caused by arsenic emission and the deteriorated quality of the copper product, smelting of enargite concentrates is not an effective method for the extraction of copper (Padilla et al., 2010; Safarzadeh et al., 2014; Long et al., 2012). Therefore, a number of studies have been carried out to develop arsenic removal technologies from copper ores using various methods such as hydrometallurgy (Curreli et al., 2009; Tongamp et al. 2010), pyrometallurgy (Dalewski., 1999; Padilla. et al, 2012) and bio-oxidation (Ehrlich, 1964; Sasaki et al., 2010). Several hydrometallurgical leaching processes like sulphuric acid, alkaline sodium sulphide and sodium sulphate and ferric sulphate leaching have been developed to treat enargite ores and concentrates. The alkaline sodium sulphide leaching process has been developed for the removal of arsenic from enargite (Anderson, & Twidwell, 2008; Parada et al., 2014). The mechanism of the dissolution of arsenic from enargite in sodium sulphite solution is"

Jan 1, 2016

By F. Hassani, R. Mikaeil, B. Ferdowsi

Two analytical solutions for estimating the ovaling deformation and forces in circular tunnels are widely used in engineering practice, due to soil-structure interaction under seismic loading. This paper discusses the discrepancy between the two analytical and finite difference methods. A comparison of the two analytical methods shows that the calculated forces and displacements are identical for the condition of full-slip between the tunnel lining and ground. However, the calculated lining thrusts differ by an order of magnitude when considering no-slip between the tunnel lining and the ground. The analytical solutions are compared to numerical analyses of the no-slip condition using the finite difference method to validate which of the two methods provide the correct method. To achieve the goal, numerical modelling, field data of the Kuhin railway tunnel and Manjil earthquake characteristics are used. Numerical analysis results agree with one of the analytical solutions that provides a higher estimate of the thrust on the tunnel lining, thus highlighting the limitation of the other analytical solution.

Jan 1, 2008

By Tiziana Apuani

The study refers to an area of about 70 km2, extended in the Italian Central Alps along the the San Giacomo Valley where different civil and mining works are present The regional geological setting is related to the Pennidic Nappe arrangement, characterized by the emplacement of sub-horizontal gneissic bodies resulting from the Mesoalpine isoclinalic folding of crystalline basements, emplaced throw East, and separated by matasedimentary cover units. The investigated geological rocks, belong to the ?Tambò?Unit, overlapped by its meta-sedimentary cover and by the ?Suretta? Unit. The tectonic contact gently dips to the E?NE. The valley, furrowed by the Liro Torrent, follows a N?S striking tectonic lineament, almost parallel to the nappe contact. A number of more then 70 sites, distributed in the studied area, were chosen and detailed structural and geomechanical field surveys were performed to characterize the rock masses, its intact rock and discontinuities. The procedure identifies the number of joint sets, their representative orientation, the set spacing, the type of movement, the amount of dilatation, the degree of alteration, the roughness coefficient, the presence and nature of infill. Rock mass quality indexes, such as the Rock Mass Rating (Bieniawski, 1989) or the Geological Strength Index (Hoek et al. 2002) have been evaluated in each surveyed site. The work is a contribute in assessing how the geometrical and geomechanical properties of the rock masses can be regarded as regionalized variables. The spatial distribution of these properties is a function of the geological and structural phenomena at which the rock masses have been subjected. The presented geostatistical analysis regards the spatial variation of the orientation of the most representative joint set.

May 1, 2009

By A. Eslami

When producing high grade mattes in smelter plants, copper losses in the slag are inevitable. Hence, the recovery of copper from slag has been the focus of many research projects, copper flotation being a widely used process. In the Khatoonabad converter, 200 t/d of slag are presently produced with a grade of 2.3% Cu. Assuming a 85% recovery by flotation, about 3.9 t/d copper could be produced, which is of prime importance economically. In the present work, a sample of this slag was crushed and ground in the laboratory to liberate the copper-bearing minerals. Following mineralogical studies, optimum laboratory flotation parameters were obtained, resulting in 90.27 % recovery and 9.42% of copper grade for the rougher stage. At pilot scale, a pulp of 28% solids was floated at ph 10.5 using 40 git of Z6, 25 git of R407 and 20 git of MIBC and A65. A recovery of 83.52% was obtained for a concentrate grade of 28.36% Cu. In the rougher cells, the feed consisted of particles 94% finer than 74 microns whereas in the cleaning stage, particles were 95% finer than 44 microns. Based on the obtained results, the Sarcheshmeh Khatoonabad converter slag is presently being fed to the Chargonbad flotation plant, where a concentrate of 32% copper at 86% recovery is being produced.

Jan 1, 2007

By Marie-France Caille, Carl Caumartin

"Audrey Resources' Mobrun Mine is located in Dufresnoy Township some 30 km to the northeast of Rouyn-Noranda. The Mobrun tine deposits consist of two major massive sulphide lenses; the newly discovered 1100 lens located 450 m south of the Main zone, and one ntellite lens, the 850-870-930 complex.The 1100 lens is hosted in a large pyroclastic unit bounded to the mouth by rhyolite and andesite flows and by a thick rhyolite unit to the art h. stratigraphically belon the three lenses. It consists of approximately 15 Mt of massive sulphide, of which 10.4 Mt are of economic ade. The Main lens lies comformably above thick rhyolite flows, ecciated and pyritized by the sulphide horizon, and is overlain by ttercalated ash tuffs and rhyolites which host the over-lying satellite ns. The Main lens consists of 3,3 Mt of massive sulphide, of which 523 000 tgrade0.80% Cu, 2.44%Zn,30.19 g/t Ag, and2.2 g/t Au he richer satellite lens (180 000 t) is part of the same volcanic stem and resulted from a resurgence of hydrothermal activity. ructure played a role in the final geometry of the deposits but did not eatly affect the volcanogenic and stratabound nature of the lens. Total production since the start of operations in 1985 amounts to 58 000 t and is restricted to the Main lens. Production from the 100 will come onstream in 1992 as ore reserves from the Main and e satellite lenses are exhausted.RESUMELa Mine Mobrun de Ressources Audrey inc. est située dans le mton Dufresnoy, a quelque 30 km au nord-est de Rouyn-Noranda. lle comprend deux lentilles de sulfures massifs d’importance majeure ~i sont les lentilles 1100 et Principale, et une lentille satellite qui est complexe 850-870-930.La lentille 1100, la plus ancienne, est contenue dans une épaisse 1ite pyroclastique bordée, au sud, par des coulées de rhyolite et d’andésites, au nord, par une épaisse unité rhyolitique. Cette lentille contient environ 15 Mt de sulfures massifs dont 10,4 Mt à teneur économique. La lentille Principale repose sur d’épaisses coulées rhyolitiques brechifiees et pyritisees à proximité de l’horizon de sulfures; elle est recouverte de tufs a cendres et de rhyolites interlitees qui contiennent la lentille satellite. La lentille Principale est constitute de 3,3 M t de sulfures massifs dont 1 523 000 t sont économiques avec une teneur moyenne de 0,80% Cu, 2,44% Zn, 30,19 g/t Ag, et 2,2 glt Au. Faisant partie de ce second épisode volcanique, la riche lentille satellite (180 000 t) temoigne d’une recrudescence de l’activité hydrothermale. La structure joue un rôle important dans la géométrie finale des dépôts mais ellen' affecte pas, de fa~on significative, la nature volcan génique des lentilles. Depuis le début des opérations, en 1985, la production totale est d’environ 568 000 t, provenant essentiellement de la lentille Principale. La mise en production de la lentille 1100 est prévue pour 1992, alors que les réserves des lentilles Principale et Satellite seront probablement épuisés."

Jan 1, 1990

By D D. Deacon

A brief description is given of the requirements which No. 4 Shaft has been designed to meet and of the methods by which two-stage winding of rock has been accommodated in a single shaft sunk to a depth of 7,421 ft. This is followed by a consideration of the factors which led to the decision to sink the shaft as rapidly as possible. Introduction T HE No. 4 Shaft of Hartebeest-fontein Gold Mining Company, Ltd., has been sited to serve the deep southern area of the mine, some 1,500 claims in extent, between the elevations of 6,250 and 7,300 ft. below surface. From this area, some two-thirds of the mine production will ultimately be obtained. Above the 6,-250-ft. elevation, an existing shaft hoisting system, situated adjacent to the reduction plant, is of adequate capacity to cater for all ore hoisted to surface for milling. No. 4 Shaft will therefore be required to provide downcast ventilation of 1,000,000 c.f.m., together with surface hoisting facilities for men, material and development rock and underground hoisting

Jan 1, 1965

By A. L. Mackie

Arsenic is a highly toxic element and a known human carcinogen. It is often a contaminant in mine water discharges, particularly from gold mining and roasting operations, and has been found at concentrations as high as 4 g/L. Increasingly stringent controls on the concentration of arsenic allowed to be discharged into the environment cannot always be met by current treatment processes and thus enhanced technologies are required to achieve low treated effluent concentrations at reasonable costs. A novel two-step physicochemical treatment process was developed and tested at the bench-scale using synthetic mine water solutions (SMWs) in order to reduce arsenic concentrations in treated effluent to exceed current discharge regulations. The process includes chemical coagulation with ferric sulphate and ballasted flocculation for enhanced solid-liquid separation. The SMW contained 59 ± 2 mg/L arsenic with a slightly alkaline pH. The effect of arsenic speciation on the process was tested by using either arsenite, As(III), or arsenate, As(V), in the SMW solutions. Tests were performed at cold temperature (3 ± 1 °C) to ensure the process? efficacy year-round in northern climates. Arsenic in As(V) SMW solutions was removed to a final total concentration of 0.0054 ± 0.002 mg/L, almost half the current recommended allowable drinking water concentration of 0.010 mg/L. As(III) was removed to a concentration below the Metal Mining Effluent Regulations? (MMER) current limit of 0.50 mg/L, to 0.32 ± 0.06 mg/L, with double the coagulant dose used in Stage 1 for As(V) removal. Oxidant addition at a concentration of 5 mg KMnO4/L in Stage 2 resulted in final total arsenic concentrations of 0.017 ± 0.002 mg/L and pH adjustment with lime in Stage 2 resulted in 0.0415 ± 0.0007 mg As/L in treated As(III) SMW. Both of these modifications to the process reduced arsenic concentrations to below proposed new MMER limits (i.e., 0.10 mg/L). This two-stage treatment process was shown to reduce arsenic concentrations to well below current treatment guidelines while reducing or eliminating chemical oxidant demand.

Aug 1, 2013

By Michael Cullen

The Quinsam Coal Mine is in the process of extracting a panel beneath the mine access road. To prevent damage to the road, and adjacent powerline, it was necessary to establish a ?protection zone? within which the extent of mining would be limited. The protection zone is analogous to the setback limit. For economic reasons it was desirable to make the protective zone as small as possible, and with the greatest extraction ratio possible. The size of the protective zone was established by determining the angle of critical deformation using the numerical model FLAC, coupled with field data on subsidence from the Mine. The angle of critical deformation (defined as the angle, measured from the vertical, at which intolerable damage is sustained by a structure) was determined to be between 0 and 5 degrees. The allowable extraction ratio in the protection zone was established using the ARMPS method of pillar design coupled with field data on pillar stability from the Mine. Excavations spans to prevent caving were determined by historical experience and field data on span stability from the Mine, coupled with numerical modeling. The minimum pillar size was determined to be 7.5m. The maximum pillar height was established at 2.5m. The maximum allowable excavation span was determined to be 8.5m. With these criteria, extraction ratios between 54 and 73% (measured in plan view) were determined to be appropriate for the protection zone.

May 1, 2001

By Joseph Poplawski, Christian O’Keefe, Robert Maron

"Unique and difficult minerals processing measurement situations will be described and the application of passive array-based technology for performing these measurements will be illustrated. Examples of these situations include volume flow measurements in froth lines and flotation feed lines with entrained air, slurry lines with magnetite and other magnetic ore, slurry lines with abrasive or corrosive materials, high pressure lines, and slurry and non-slurry lines exhibiting scale buildup. Other situations include measurement of entrained air levels in pipes and in flotation columns. These measurements are achieved through CiDRA’s patented passive sonar array-based technology. This technology performs two independent measurements – flow rate and entrained air amounts. Two hardware implementations of this technology have resulted in a non-intrusive flow and entrained air meter which easily is installed on the outside of the pipe, and a submersible version for entrained air or gas holdup measurements which is used in flotation processes. Recent developments in extending this technology to solve other unique measurement problems such as valve movement confirmation will be covered.INTRODUCTIONThe minerals processing industry faces many unique and challenging process control conditions and environments. In terms of flow measurements, many of these situations are not being properly served by traditional flowmeters such as ultrasonic meters, magmeters, turbine meters, orifice plate meters, vortex flow meters, Coriolis flow meters, and venturi meters. A new class of flowmeters has been developed that solves these unique situations and flow measurement problems. This new class of flowmeter technology utilizes sonar-based processing algorithms and an array of passive sensors to measure not only flow, but also fluid composition. It does so accurately, reliably and without making contact with the fluid. These measurements are performed on practically any type of fluid within virtually any type of pipe.PRINCIPLE OF OPERATIONSonar array-based flowmeters are ideal for tracking and measuring the mean velocities of disturbances traveling in the axial direction of a pipe. These disturbances generally will convect with the flow, propagate in the pipe walls, or propagate in the fluid or slurry. First let us focus on the disturbances that convect with the flow. The disturbances that convect with the flow can be density variations, temperature variations or turbulent eddies. The overwhelming majority of industrial flows will have turbulent eddies convecting with the flow, thus providing an excellent means of measuring the flow rate as described below."

Jan 1, 2008

By H. Utsunomiya, J. Su, K. S. Tamimi, S. Yue, M. Sanjari, L. Kestens, A. S. Kabir, R. Petrov

The recrystallization and texture evolution of AZ31 sheets obtained through two stages was investigated at different stages of post-rolling annealing. The specimens were rolled at the temperature of 100? and at a fast speed of 1000m/min and low speed of 15m/min with a reduction of 30% in one pass. The results show thatAZ31B sheets rolled at 15 m/min and 100 °C has fractured for reductions of more than 30% per pass. However, by increasing the rolling speed to 1000 m/min the rollability was improved significantly and the material can be rolled to reductions of more than 70% per pass. The high speed and heavily deformed materials are fully recrystallized even at 100 °C.

Jan 1, 2015

By J. B. Yianatos, F. A. Contreras

Mineral carrying rates across the pulp-froth interface in the range of 1.4-3.2 tph/m2 and bubble surface coverages between 7 and 22 %, have been determined in large rougher flotation cells operated under normal conditions. To evaluate these variables, measurements of bubble size distribution, bubble load and size distribution of particles collected by true flotation, were performed at different concentrators. A model for estimating the bubble surface coverage at the pulp-froth interface level as a function of bubble load, Sauter mean bubble diameter and particle size distribution of the mineral collected by true flotation, was developed. Bubble loadings in the range 24-70 kg/m3, bubble sizes between 1.5 and 3.2 mm and superficial gas rates between 1.2 and 2.0 emfs, were measured in copper rougher flotation cells under normal operating conditions. Estimated bubble surface area fluxes were in the range 31 to 62 cm2/s/cm2- Results of carrying rates and bubble surface coverage, at the pulp-froth interface, were significantly lower than the maximum carrying rate for the collection zone, indicating that in this case is difficult to achieve maximum carrying rate at the pulp-froth interface level. The main constraints regarding carrying capacity limitations are then related to the froth transport characteristics (i.e. cell design, froth residence time, froth stability, froth recovery). Advances in Mineral Processing Science and Technology Proceedings of the 48th Annual Conference of Metallurgists ofCIM Sudbury, Ontario, Canada Edited by

Jan 1, 2009

By Markus Timusk

This paper describes a research project that focuses on developing a diagnostic routine to assist in decision support for the condition based maintenance of mobile mining equipment subject to unsteady duty. The objective is to develop a general method for initial identification of mechanical faults in machinery operating under non-steady state conditions. The platform to be used for the experimental development and validation of the system is the swing machinery in an electric mining shovel operating in an open pit mining environment. The strategic approach that is proposed for the diagnostic routine will involve using novelty detection augmented by temporal classification.

May 1, 2004

By V. S. Melamud, A. G. Bulaev

"Sulfide ore processing is characterized by significant losses of metals, for example, with flotation tailings. Therefore, flotation tailings can present an important source of metals. The treatment of flotation tailings is a technological issue due to low content of valuable metals (e.g. copper and zinc) and high iron content. The goal of the present work was to investigate extraction of base metals and gold from old flotation tailings containing 17% iron, 0.26% copper, 0.22% zinc, and 0.7 g/t gold. Chemical leaching in columns with 0.5 to 10% sulfuric acid solutions and bioleaching in air-lift percolators were used for the extraction of base metals. 1% H2SO4 provided high extraction of base metals and best selectivity. Recovery of copper and zinc reached 35% and 40%. Increasing H2SO4 concentration did not lead to significant increase in base metals extraction, but led to increased iron concentration in the pregnant solutions. The 1% H2SO4 leaching solid residues were used for further experiments. Acid leaching and bioleaching did not result in significant increase in base metals recovery. Leaching of the residue with 10% H2SO4 provided extraction of 4% copper and 20% zinc, and bioleaching for 100 days provided extraction of 16% copper and 12% zinc. The recovery of gold by cyanidation from the flotation tailings, the 1% H2SO4 leaching solid residue and residue of two-step acid leaching were 50, 49, and 65%, respectively. The results suggest that acid leaching can be effective approach for treatment of old flotation tailings.INTRODUCTIONAbout 80% of base metals (copper, nickel, zinc) are produced through pyrometallurgical treatment of sulfide ores. Mining and metallurgical treatments of sulfide ores are presently characterized by significant losses of non-ferrous and precious metals into the various wastes. For example, flotation of sulfidic ores yields formation sulfide flotation tailings. These wastes pose a problem not just because of their sheer volume, but some of them may affect local ecosystems. Nowadays, mining and metallurgical wastes can be considered as technogenical source of base metals."

Jan 1, 2016

By J. Holtzhausen

"An analysis of operating costs for underground mining operations in the current global economic climate considers the financial burden for a ventilation system to be in vicinity of 30–50%. These operating costs are often overlooked by management teams in the greater scale of things when looking to save money in the procurement phase. This oversight to operational costs can be as little as the effective delivery of ventilating air to the work zone. One of the biggest opportunities for substantial savings and improvements can be realized in the selection of a quality vent duct that is suitable and fit for purpose. The key component to the selection process for quality products is a solid selection criterion that is followed during the tender process. This allows the management team to filter out unsuitable suppliers and products. The case study identifies some of the key components that were considered in the procurement and selection process for lay-flat ducting products. These included the characteristics and strength of the materials. The selection criteria also considered a ducted performance comparison that offers the best aerodynamic performance with some of the following components. • Airflow at Duct Face (m³/s or cfm), Fan Pressure (Pa), Air Power (kW or hp), Assumed Fan Efficiency, Annual Running Costs, K – Factor or Lambda, Leakage Factor, Joining Method and available Duct Lengths with weights. • Critical components considered were the current Testing standards available for: o Bursting Strength AS 4878.5, Mass AS 2001.2.13, Tensile Strength AS 4878.6, Tensile Strength AS 2001.2.3.1, Tear Strength AS 2001.2.10, Tear Strength AS 4878.7 • Other key considerations were: o Warranty Claim, Damage claims, Local storage / operations, community engagement and relationships, current testing data, previous issues, after sales service with technical assistance, etc. Hidden costs associated with product failures and operational stoppages are often not quantified during the tender process. Buyer beware. INTRODUCTION The largest constraint underground mining operations currently face are their operating efficiencies and electrical costs. The overall cost of an operations ventilation system and its components requires more focus to identify what opportunities exist to maximise efficiency and effectiveness. This tender process may not always be clearly understood and it should be broken down into fewer components to achieve quantifiable results."

Jan 1, 2019

By H. Jung

The microstructures and grain sizes of nanocrystalline Co and Co-1.1 wt % P alloys were characterized by X-ray diffraction. The corrosion properties of as-deposited and annealed Co-l.l wt % P alloys and pure nanocrystalline Co were studied in a deaerated 0.1M H2SO4 solution using electrochemical measurements and XPS surface analysis. Potentiodynamic polarization testing and impedance measurements revealed that an addition of P to pure nanocrystalline Co improved the anodic-corrosion resistance, however, this enhanced corrosion resistance was compromised by the heat treatment process at 350 and 800°C. The corroded surface of annealed nanocrystalline Co-1.1 wt % P at 350°C exhibited a more porous P-enriched layer compared with the as-deposited case. A sample annealed at 800°C corroded uniformly showing second-phase precipitates. The enhanced corrosion resistance of as-deposited nanocrystalline Co-1.1 wt % P was attributed to the elemental P on the surface that acted to hinder the anodic dissolution kinetics of surface Co following an initial selective dissolution of Co. However, at a higher anodic overpotential region, the superior corrosion resistance of Co-l.l wt% P alloys did not last because of the formation of a non-protective surface film consisting of mainly hypophosphite and phosphate as well as elemental P.

Jan 1, 2005