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By Sadan Kelebek, Gerhard W. Heinrich, Peter F. Wells, Peter J. Whittaker

"Based on plant mineralogical data & circuit modelling, a pilot plant campaign was conducted at Strathcona Mill to identify options for improving the performance and operating flexibility of the copper-nickel separation circuits. Emphasis was given to options for combined treatment of the two feed streams (primary nickel ore rougher concentrate and copper ore circuit concentrate) to obtain more flexibility in the use of existing column flotation capacities. The pilot test results demonstrated that circuits which incorporated preconcentration prior to column flotation gave the best metallurgical results. Maximum column carrying rates were also related to the preconcentration. Predictions that increased circulating loads from the column scavengers would improve copper-nickel separation performance were confirmed.IntroductionThe copper-nickel separation circuits at Falconbridge's Strathcona Mill produce a high grade (>30% Cu) copper concentrate most of which is treated at the Falconbridge Kidd Creek smelter near Timmins. The remaining nickel concentrate which still contains approximately 20% of the feed copper is combined with other lower grade nickel/copper concentrates and shipped to the Falconbridge nickel smelter. The separation is required because of the limited capacity for copper processing in and downstream from the nickel smelter. Since no nickel is recovered at present in the Kidd smelter, the challenge is to minimize the nickel content of the copper concentrate.At present there are two independent separation circuits treating the rougher concentrates from the copper and nickel circuits as shown in Figure I. The minerals treated are mainly chalcopyrite and pentlandite but there is up to 25% of the copper present as cubanite in the copper ore. The circuits were kept separate because the flotation rates and froth characteristics were very different, probably because of the presence of cubanite and the large difference in the concentrate grades feeding the circuits."

Jan 1, 1995

By Tobias Riedel, Peter Ay, University of Applied Sciences, Bianka Bade, Nicole Pausch, Bernhard Gemende, Florian Logsch, Matthias Leiker, Satyanarayana Narra, Claudia Glaser

This study focusses on the production of micro-structured agglomerates with high porosity from iron (III)-hydroxide-containing materials as adsorbent. A fractal-like structure of the agglomerates was created through a two-stage agglomeration process. Developing porous structures was necessary to increase the adsorption capacity. Various materials were used for the agglomeration experiments: i) filter cake (dry matter content 23 to 27 wt.-%); ii) FerroSorp®DGµ, a product which is already being used commercially for various applications in water cleaning. These materials in the pristine state have different characteristics such as dry matter content and particle size distribution. The desired optimised agglomerates can be achieved with the help of additives or binders, respectively. Additionally, the agglomeration process parameters play a vital role. In order to produce agglomerates with high porosity, the process was carried out in two steps. Firstly, FerroSorp®DGµ (d90 < 30 µm) was agglomerated jointly with the aforementioned filter cake in an Eirich intensive mixer (particle size range 125 µm < d90 < 500 µm). The second agglomeration step post drying was carried out in a pelletisation disc or in an Eirich intensive mixer with the addition of polymeric binders (up to particle sizes ranging from 2 to 4 mm). Agglomerates were characterised both in short and long term loading tests using phosphate as model ion. Contact times were less than 10 min, similar to practical application in percolation columns, or more than 7 days, respectively. The results show that the agglomerates produced in the two-stage pelletisation process have more advantageous properties like better accessibility of the active sites obtained by specificly influencing the particle size distribution during the two-stage pelletisation process.

Jan 1, 2016

By P. Keller, T. E. Lister, G. G. Clark, L. A. Diaz

As electronic technology continues to evolve there is a growing need to develop processes which recover valuable material from antiquated technology. This need follows from the environmental challenges associated with the recovery of raw materials and fast growing generation of electronic waste. Although present in small quantities in electronic devices, the availability of raw materials such as rare earths and precious metals becomes critical for the production of high tech electronic devices and the development of green technologies (i.e. wind turbines, electric motors, and solar panels). Therefore, the proper recycling and processing of electronic waste presents an opportunity to stabilize the market of critical materials, providing a proper disposal and treatment of a hazardous waste stream. This paper describes the development and techno-economic assessment of a comprehensive process for the recovery of value and critical materials from electronic waste. This hydrometallurgical scheme aims to selectively recover several value streams (base metals, precious metals, and rare earths) present in electronic waste. The economic feasibility for the recovery of rare earths from electronic waste is mostly driven by the efficient recovery of precious metals, such as Au and Pd (ca. 80 % of the total recoverable value). Rare earth elements contained in magnets (speakers, vibrators, and hard disk storage) can be recovered as a mixture of rare earth oxides which can later be reduced to the production of new magnets.

Jan 1, 2016

By G. Heinrich

Size by size recoveries and pentlandite liberation data from image analysis were used to evaluate plant tests of pentlandite flotation and pentlandite/pyrrhotite middlings flotation from pyrrhotite in a circuit treating low grade magnetic intermediates at the Strathcona Mill. Raising the pH with lime gave improved selectivity for liberated pentlandite but there were major losses of middlings unless the circuit was pulled hard indicating the need for finer regrinding. However, size by size recoveries at the normal regrind showed that, while the pentlandite recovery improved in the finer ranges, the pyrrhotite recovery also increased leading to little selectivity in any size range. A much finer regrind gave greatly improved selectivity in all size classes suggesting that 'surface polishing' as well as liberation effects were being seen. Using liberation data from image analysis of the products, surface polishing accounted for over 99 % of the improvement in selectivity. Unreasonable recoveries of more highly liberated particles were observed for liberation analysis by volume fraction but showed the expected trend in recovery when the analysis was based on surface liberation. The surface liberation data showed that the improved pentlandite recoveries after finer regrinding were to more than 90 % caused by improved liberation with the remainder contributed by surface polishing effects. The practical implications of this work are that i) the comparison of particle grades with product grades allows to estimate the effect of regrinding on co-flotation/entrainment, ii) surface liberation models should be used for evaluating recoverabilities by flotation and iii) regrind circuits may be set up to ensure that all particles go through the mill or that only the coarse fraction is reground depending on the surface scrubbing found to be required.

Jan 1, 1993

By Wojciech Naworyta

The development of the Gubin lignite deposit, located in western Poland is the first mining investment of this magnitude in 30 years. Gubin is the second largest prospective lignite deposit in Poland, with over one billion tonnes of indicated reserves, which would be sufficient for a 2700 MW power plant for almost 50 years. The paper presents select problems related to the mining development of Gubin deposit. Attention has been paid to changes in legal and formal requirements for mining investments that were introduced after the change of Polish social and economic system. The significance of spatial, environmental and social conditionings has been emphasized. Typical attributes of the mine planning process have also been presented. Those include the long time horizon of the investment, uncertainty of external conditions and resulting difficulties in the project?s economic evaluation. Focus has been placed on mutual connections between particular elements of the project and time dependence during the process of their development. Iterative characteristics of the design process have been proven, together with resulting necessity of frequent verification of previous assumptions. As a background for presenting selected problems related to future development of Gubin lignite deposit, the methods of design have been discussed ? the variants method, the empirical method and the method of consecutive approximations. The tools and the possibilities of their use for solving specific problems related to mine design of an open pit have been broadly discussed. The chapter in which the authors relate to the specific problems of Gubin deposit development constitutes the main part of the paper. It depicts the assumptions on which the annual production and the total extraction of lignite during life-of-mine period have been based. The process of mine designing has been discussed, taking into consideration environmental, geotechnical, spatial, social and mining conditions. The criteria for choosing the location of the opening cut and the external dump have been mentioned. In the summary, the present state of mine design progress has been presented, together with some problems that will need to be solved during the next stages of the deposit?s development.

Aug 1, 2013

By M. K. Aly

"ABSTRACTThis paper briefly describes the mining method proposed for the iron ore deposits in Wadi Shatti, western Libya. The iron ore produced will be used in the steel plant at Misurata on the Mediterranean coast. GeologyThe Wadi Shatti is located in the central part of western Libya. Flat iron-bearing sedimentary rocks are exposed with remarkable regularity in this valley. It occupies an area 200 km long by approximately 20 km wide trending in an eastnortheast direction between Jabal Fazzan in the north and Zel1af Sand Dunes in the south (Fig. I). Only a part of this valley has been ful1y explored by drilling on a 400- by 200-m pattern in an area of 18 km2 south of the town of Taroot. Farther to the south, no exploration by drilling can be done due to the presence of sand dunes .The iron-bearing bed is about 10to 15m thick, with a dip of 2 to 4 degrees in a southerly direction . The overburden consists mainly of clayey sandstones, 4 to 5 m thick, in the northern part, increasing to 35 m in the southern part. The iron ore bed consists of three different layers, from the floor upward: (1) Lower Reduced Ore (LRO), having a varying thickness of 1 to 7 m; (2) Magnetic Ore (MO), with a thickness varying from 4 to 12 m mainly in the central part of the area; and(3) Upper Reduced Ore (URO), with a bed thickness varying from 0 to 7 m."

Jan 1, 1980

By T. Negeri, M. Boisclair, A. Bouajila

CanmetMINING has been working on process development for separation and concentration of rare earth bearing minerals from various ores. Among others, the flowsheet development for the Montviel Rare Earths/Nb deposit, located at the north end of the Abitibi region in Quebec and owned by GeoMegA Resources Inc. (GeoMegA) is now at an advanced stage. The concentrate produced by a flotation-only process at the early stage of the project accounted for higher than 90% TREO and Nb recovery in 40% mass. This concentrate was submitted to SGS Lakefield for hydrometallurgical processing. Lixiviation of the concentrate using hydrochloric acid yielded approximately 99 % REE. An overall (from ore in ground to final precipitate) REE recovery of 80% was obtained through a series of selective precipitation steps. Recently significant additional developments have been made. CanmetMINING is steadily improving the quality of concentrate produced through a simplified, more efficient flowsheet which consists of flotation, magnetic separation, and a thermal treatment procedure. The paper discusses aspects of the process that allowed obtaining 82% recovery of TREO and Nb in 16% of mass, the best result to-date.

Jan 1, 2014

Presentation Plan Hoist ropes ?Importance ?Mandatory and routine Inspections Continuous monitoring of wire ropes for mine hoisting ?Why a continuous monitoring of mine hoisting ropes.? ?Description of the ??Contiscan?? monitoring system ?Information data gathering Advantages of the ??Contiscan?? monitoring system ?Visual rope inspection on daily basis ?Monthly inspection ?Payload capacity flexibility and potential increase for deep mining operations. ?Periodical inspections by external experts ?Importance of the costs associated with hoisting production shutdowns. Commercialization and availability

Aug 1, 2013

By Louis R. Bernier

Most systematic studies carried out on passive treatment systems incorporating the use of limestone drains, under aerobic or anaerobic conditions, essentially deal with the treatment of effluents from mines in the USA. A limestone drain typically comprises a trench filled with crushed limestone rocks surrounded by impervious materials. In the case of anoxic drains, the trench is generally isolated under a geomembrane liner covered with soil. Contaminated mine effluents flow by gravity towards these limestone drains where the passive treatment consists in bringing the carbonates into solution, thus increasing the pH and alkalinity. Factors to be considered in the design are: chemistry of the contaminated mining effluent ([metals, [Al], [sulfates], dissolved oxygen, acidity, pH), desired residence time, nature of the rock, grain size, density, void ratio and mass of the limestone rock used, alkalinity, life span and hydraulic behaviour of the drain. Certain secondary effects (saturation by sulfates, co-precipitation, adsorption) may either be beneficial (lowering the sulphate and metal load) or harmful (armoring the limestone rock and clogging of the drain). This paper gives a brief status report on the available knowledge on the use of limestone drains for passive treatment of mining effluents heavily loaded with metals. Examples given include the old Lorraine mine (Ni-Cu), Témiscamingue, where limestone and dolomite rock drains were installed in 1998 when a cover with capillary barrier effects was placed over the tailings site.

Nov 1, 2002

By M. Jackman

"Froth stability is critically important for an effective separation of fine and ultrafine particles using column flotation. While a desirable froth stability is required to facilitate the drainage of entrained gangue particles for secondary cleaning, excessively stable froths could create problems in separation efficiency and down-stream processing. In this work, we demonstrate that a stable foam was produced using a poly (N-isopropylacrylamide) (PNIPAM) solution at the temperatures below the critical solution temperature. In contrast, very little foam was produced at the temperature above the critical solution temperature (~32 oC) of PNIPAM. We can also tune the foam stability by changing the temperature of wash water. When warm water of temperature of 50 oC was applied, the foam collapsed and diminished. This is probably attributed to the switching of PNIPAM molecules from coiled to a globule conformation as the temperature is increased. Our results indicate that the foamability and foam stability can be tuned by tempartureresponsive PNIPAM. This is the first time that thermo-responsive PNIPAM has been shown to have potential as a modifier of froth stability in column flotation.INTRODUCTIONIn mineral processing, column flotation has become a more prevalent process in the separation of fine mineral particles (typically < 100µm). Traditional flotation cells are sometimes less efficient in selective separation of ultra-fine particles (Kohmuench et al., 2009). The efficiency of flotation columns is dependent on the froth stability at the top of the cell to drain the unwanted gangue particles entrained in the froth. The greater froth height in flotation columns, compared to mechanical flotation cells, gives a longer froth residence time, allowing more of the drainage process to take place before the froth reports to the launderer. Secondary cleaning, by applying a wash water shower to the froth, enhances drainage of gangue particles entrained in the froth, and improves the recovery of the targeted minerals (Finch et al., 2007).While a stable froth is desirable to facilitate drainage of entrained gangue particles, greater amounts of fine particles have been found to contribute to excessively stable froths which could create problems for down-stream processes (Bethell et al., 2012). Operators have tried spraying additional wash water on the froth, and reducing frother dosage to destabilize the froth and to prevent backup at the launderer. The former is sometimes inefficient, and the latter results in unacceptable recovery loss. Currently, the best practice is often the installation of larger froth launderers, adding expensive defoaming agents, and/or using deaeration tanks (Bethell & Luttrell, 2005)."

Jan 1, 2016

By C. S. Touchard, F. Bourgeois, L. Cassayre, Julcour-Lebigue C.

Developing a hydrometallurgical process that can overcome the refractory nature of chalcopyrite (CuFeS2) to leaching still eludes the minerals industry, despite major research efforts spanning decades. The several leaching experiments of this ore in sulfuric acid tend toward the same conclusion that the precipitation of elemental sulfur or iron containing species are the main dissolution inhibitors. Describing both the electrolyte composition and the various solid phases in equilibrium for the Fe-H2SO4-H2O system is essential to overcome this issue and to find an operating window. According to the thermodynamic models and to the thermodynamic data, we demonstrate, in this study, the need to use the Specific ion Interactive Theory (SIT) model to describe complex electrolytic systems through a limited number of parameters. Hence, the binary interactions for our case study have to be defined and estimated. This work develops a methodology to determine these parameters of the SIT model up to 90°C and an acid concentration of 1 mol/kg from published data for H2SO4-H2O and Fe(II)-H2SO4-H2O sub-systems or from new experimental data for the Fe(III)-H2SO4-H2O one. This method will eventually be applied to copper systems to lead to a thermodynamic description of the complete CuFeS2-H2SO4-H2O system. INTRODUCTION A leaching process consists in the dissolution of a solid phase into an aqueous medium in order to extract compounds of interest. It is a key step in the hydrometallurgical processes used in the primary production and recycling of metal components. One of the main leaching issues is the formation of passivating layers on the surface of solid particles, whose consequence is a partial or total hindrance of the dissolution. The leaching of chalcopyrite (CuFeS2) in sulfuric acid is a prime example. No industrial solution for the atmospheric pressure leaching of this ore has been developed to date because of passivation phenomena caused by the formation of elemental sulfur or iron-containing precipitates on the surface of the particles (Li, Kawashima, Li, Chandra, & Gerson, 2013). In the 1970s, the Pourbaix diagrams established for the Cu-Fe-S-H2O system at 25°C for low concentrated solutions (Garrels & Christ, 1965; Peters, 1976) agreed that copper and iron are fully dissolved at high potential and in acidic medium (pH < 4). However, they neglected the many stable solid phases that can form in these conditions. In addition, the Pourbaix diagrams of chalcopyrite in water, calculated with two recent databases (FactPS and Thermoddem 2017) associated to FactSage and PhreeqC softwares respectively, show large discrepancies at 75°C (Figure 1). Indeed, at high potential, FactSage predicts the formation of goethite FeOOH at a pH value of 1, while PhreeqC predicts the formation of jarosite (H3O)[Fe3(SO4)2(OH)6] for pH greater than 2.5.

Jan 1, 2019

By A. Malfliet, A. Van den Bulck, D. Willems, M. Guo

The feed stock of copper fire refining processes consists of two major constituents: blister copper and high-Cu secondary scrap. A large fraction of this secondary scrap, originating from urban mining, is quite high in impurities such as Fe, Ni, Pb and Sn. Problems arise in electro-refining when these elements exceed certain limits. As a consequence, transformation of the secondary copper raw materials into high grade copper has become troublesome. However, the behavior of impurities relevant to fire refining conditions, essential for anode refineries, has not been studied in detail. Furthermore, limited research has been performed on the distribution of Ni between copper metal and slag. In this study, equilibration at 1300°C of liquid Cu-1wt.% Ni alloy and CaO-CuOx- FeOy-SiO2 slag in vacuum sealed quartz ampoules was performed for 12 hours, followed by water quenching. Electron Probe X-ray Microanalysis was employed to investigate the composition of the equilibrated alloy and slag. The distribution coefficients of Ni between slag and alloy were determined as a function of oxygen partial pressure and CaO concentration. INTRODUCTION Europe’s initiatives for a circular economy, pushing for resource-efficient ways for a sustainable future, puts Europe at the forefront of the globe. In Europe, 50% of the copper originates from recycling, being an excellent example of Europe’s circular economy (Juncker, 2018; Lacasse, 2018). High-grade copper recyclates are made available which regularly contain impurities like Fe, Ni, Sn and Pb. Aurubis Belgium is responsible for recycling these types of recyclates to anodes, cathodes and copper products. However, after fire refining of copper, the aforementioned impurities remain in the anode copper, leading to problems in electrorefining. During the fire refining of copper, a CaO-“Cu2O”-“Fe2O3”-SiO2 slag is used. Fundamental information of the distribution of impurities between the copper and this slag system is therefore essential to understand and improve the copper fire refining process at Aurubis Belgium.

Jan 1, 2019

By Changmiao Liu, Dong Dong

"Magnesite is a major mineral resource in China, whose output and export are massive and quite important to the world. Flotation performances of eight anionic collectors on magnesite were studied to find the proper ones. Micro-flotation results indicated that collectors of HOl, SDBS and AEP had outstanding flotation performances on magnesite, with the recovery ratio all reaching above 90% in a wide range of alkaline pulp. Solution chemistry calculation, Zeta potential and FTIR measurements were employed to reveal the interaction mechanism between anionic collectors and magnesite surfaces. Solution chemistry analysis indicated that the main active components of HOl, SDBS and AEP were anions and association complex anions in alkaline pulp. Magnesite surface was negatively charged in the alkaline pulp. The active sites on magnesite surface were mainly the components of MgOH-, Mg (OH)2, and MgCO3 when the pulp pH was over 8. Anionic components of HOl, SDBS and AEP reacted with Mg in magnesite surface to make the insoluble anionic salts. And the chemical adsorption was confirmed by the FTIR measurement evidences.INTRODUCTIONAs an important industrial mineral, magnesite has been widely applied in metallurgy, building materials, chemical industry, agro-pasturage and other fields (Hu et al., 2006). It is a major mineral resource in China with abundant reserves. Its output and export are massive and quite important to the world. Liaoning province is rich in magnesite, whose reserves and production both come first in China. So far the proven reserves have reached 2.91 billion tons, accounting for more than 90% of total reserves in China, and more than 25% in the world (Li et al., 2012).Magnesite ores in Liaoning province mainly contain crystalline magnesite, and the gangue minerals of quartz, dolomite, talc, tremolite, chlorite and carbonaceous materials. Some ores also contain a little pyrite, chalcopyrite, hematite and apatite. Most of the ores must be preprocessed to separate the useful parts from the large amount of gangue minerals, so that their value and industrial application can be enlarged and promoted.Flotation is commonly used to process the magnesite ores, and the conventional flotation methods include anionic and cationic flotation. In the former, gangue minerals of silicates were separated by using amine type cationic collectors. After the impurity minerals were removed, high qualified magnesite concentration could be gained. For the cationic flotation method, it had some drawbacks of lower recovery ratio and poor performance in the lower temperature pulp (Li et al., 1995)."

Jan 1, 2016

By Yuhua Wang, Wei Sun, Dongfang Lu, Yuehua Hu, Tao Jiang

"With the continuous depletion of high-grade tungsten ores, the low-grade refractory wolframite has become an important source for the tungsten industry. However, the large-scale utilization of low-grade refractory wolframite is disappointedly scarce, it is due to leanness of valuable minerals, complex mineral characterization and use of toxic collectors.The aim of this study was to develop an environment-friendly processing flowsheet, which was used for beneficiation of low-grade tungsten ore containing 0.45% WO3 from New Brunswick in Canada, and the content of arsenic is 0.68% in this materials. The objective of beneficiation studied was to obtain a concentration containing 60% WO3 with less than 0.5% of arsenic under the condition without any flotation reagents. In this environment-friendly processing flowsheet, magnetic separation technology, classification and gravity separation technology were combined and employed. Firstly, magnetic separation was employed for the first pre-concentration process, over 95% of the feed was rejected as the final tailings, meanwhile, over 85% of arsenic minerals were removed with tailings. Secondly, gravity separation tests were employed on the classified magnetic pre-concentrate in the second pre-concentration process, a second pre-concentrate with WO3 grade of 54.32% and As grade of 5.41% was obtained. Finally, two stage magnetic cleaning were conducted on the second pre-concentrate for upgrading the second pre-concentrate, the mostly arsenic and gangues were removed from the second pre-concentrate. Finally, this novel processing achieved an effective recovery of fine wolframite from low grade tungsten ore with high arsenic. A concentrate with WO3 grade of 64.81% were obtained in new opened processing flowsheet, and As grade was less than of 0.5% in this concentrate. Furthermore, remarkable feature of this processing flowsheet is environment-friendly due to avoid using the toxic flotation reagents during the whole mineral processing.INTRODUCTIONTungsten is an essential metal and is the most widely used in industrial production such as national defense, electro-mechanical products, textile and leather sectors (Srivastava et al., 2000). Wolframite and scheelite are the main mineral sources of tungsten metal. With the continuous depletion of high-grade tungsten ores and the increasing demand for tungsten alloy, the low-grade refractory wolframite has become an important source for the tungsten industry. However, the large-scale utilization of low-grade refractory wolframite is disappointedly scarce, it is due to leanness of valuable minerals, complex mineral characterization and heavy use of toxic collectors."

Jan 1, 2016

"In 1989, improved milling and recoveries resulted in a record production of 7 722 200 Ib Nb2P5 at the Niobec Mine. At the beginning of 1990, proven and probable reserves stood at 11 952 000 tons grading 0.65% Nb205. This year, the world steel market has slowed and as a result, niobium consumption is down. Consequently, production at the mine dropped 14% for the second quarter.IntroductionThe Niobec Mine is located approximately nine miles northwest of Chicoutimi, Quebec, (Fig. I). After ten years of production, the Niobec Mine is the only operating niobium mine in North America and the second largest producer of niobium in the world after Brazil.The Niobec Mine is a joint venture held equally by Teck Corporation and Carnbior. Teck Corporation operates the mine while Cambior is responsible for marketing and selling the niobium pentoxide concentrate (Nb2P5).The deposit was discovered in 1964 following an airborne radiometric survey in search of uranium. However, mining of the deposit did not begin until 1976. The initial production capacity of 1500 tons per day was increased several times to the present capacity of 2450 tons . The deposit is considered as an ample source of pyrochlore concentrate for at least 15 years at the present production rate.Niobium, also known as columbium, is a greyish-white metal used primarily as an alloy in various steels and other metals. The use of niobium in alloys heightens tensile strength and improves corrosion resistance while at the same time reducing the weight of the alloy. Niobium is also used in the fabrication of steel for the aerospace, automobile, and oil and gas pipeline industries. Although substitutes for niobium exist, their use often results in diminished performance and net cost increases."

Jan 1, 1990

By Jiahao Xu, Namsoo P. Kim, Guikuan Yue

Iron (ferrous ions and ferric ions) plays an important role in the heap leaching of copper sulfide minerals and copper electrowinning. In our previous study, a novel oxidation-reduction potential (ORP) equation only based on the variables of temperature and nominal ferric/ferrous ratio was developed to predict the redox potential of the quaternary H2SO4-Fe2(SO4)3-FeSO4-H2O system, and its applicability was also extended to more complex acidic iron sulfate solutions, especially those generated during the acidic leaching and bioleaching of Printed Circuit Boards (PCBs) to recover copper. In this work, the broader range of applicability of this equation to other complicated acidic iron solutions containing copper ions has been extended and determined. Specifically, synthetic iron-containing solutions with copper ions according to the conditions of industrial electrolyte during copper electrowinning and heap leaching have been used to measure the ORP and test the developed equation. It seems that this equation can still be employed to predict the redox potential of the Fe3+/Fe2+ couple for the simulated solutions under conditions related to copper electrowinning and pregnant leaching solution generated during heap leaching over a wide range of solution concentration, nominal Fe3+/Fe2+ ratio and temperature. It has shown that the prediction and measurement of redox potential is highly useful to estimate and understand the iron chemistry of industrial leaching and electrowinning processes. INTRODUCTION It is well-known that iron (ferric or ferrous) ions are widely involved in the hydrometallurgical production of non-ferrous metals such as Cu, Au and Zn, mainly in the form of acidic iron sulfate solutions. As far as Cu is concerned, in a typical leaching-solvent extraction-electrowinning process, ferric ion plays a central role in the heap leaching process and is reduced to ferrous ions in the pregnant leaching solution. Ferric ion is the most important surrogate oxidant and the dissolved ferric/ferrous couple plays an important catalytic role in accelerating the leaching rate with oxygen as an oxidant (Peters et al., 1986; Schlesinger, King, Sole, & Davenport, 2011). In addition, during the traditional electro-winning process, oxygen is evolved at the anode, and copper is reduced at the cathode. A common idea is that iron in the electrolyte in copper electrowinning cell (with total iron of 1–4.4 g/L) plays an important role in reducing the current efficiency. According to the published results (Winand et al., 1975; Dew et al., 1985; Moats et al., 2009; Khouraibchia et al., 2010), thecurrent efficiency decreases linearly with increasing ferric ion concentration and the energy consumption increase was mainly caused by the decrease in current efficiency.

Jan 1, 2019

By I. Farr, P. Quinn, H. Mani

"Difficulties in the flotation of Inca's ultramafic ores at the Thompson Mill are related to the presence of serpentine slimes and some floatable rock minerals. Serpentine slime coatings depress pentlandite recovery while rock flotation decreases concentrate grade. Carboxy methyl cellulose (CMG) has been shown to promote sulphide flotation and to reject floatable rock.To improve ultramafic ore flotation, CMC addition in the rougher stage was studied in a laboratory scale column. Batch tests were conducted in the lab and followed up with continuous testwork at the Thompson mill. At 0.5-1.0 g CMG per kg ore, pentlandite recovery is increased by 5-10% in rougher flotation, pentlandite/rock selectivity is improved and pentlandite/pyrrhotite selectivity is unaffected.IntroductionAt Inca's Manitoba Division, two distinctly different sulphide ores are processed at the Thompson Mill, primarily by froth flotation. The sulfide ores are contained within the Thompson Nickel Belt and currently there are two operating mines, Thompson and Birchtree. Sulphide mineralization in both mines is simple, consisting of pyrrhotite (both monoclinic and hexagonal), pentlandite, and minor chalcopyrite. Both deposits are hosted by metasedimentary and metavolcanic rocks with variable ultramafic rock associations. Ultramafic rocks are mostly serpentinites: magnesium rich rocks composed predominately of serpentine. There is little serpentinite at the Thompson mine, while mineralization at Birchtree is closely associated with significant ultramafic rocks.Excellent recoveries of pentlandite (Pn) and chalcopyrite (Cp) are achieved at the Thompson Mill when processing ore from the Thompson mine. In treating the ultramafic associated Birchtree ore, nickel recoveries are some 6 to 8% below that achieved with Thompson ore. Furthermore, a typical concentrate from Birchtree ore contains more rock and is higher in MgO content. The presence of rock in the concentrate is related to free floatable rock species as well as rock locked with sulfides. The reason for the difference in flotation responses of these two ores has been the subject of on-going re-search."

Jan 1, 1998

By Christopher R. Davis

Resource modelling is the basis for any economic appraisal of a mining project and includes a number of steps from data acquisition and validation to resource reporting and classification. At each step in the modelling process it is necessary to define the specific objectives, the methodology proposed to achieve those objectives and to establish a set of checks and validation tools to assess the effectiveness of the proposed methodology. The data used has to be both reliable and relevant for the purpose of the modelling. QA/QC for database validation requires verification of sample location; including correct collar location, topographic modelling and down-hole surveying. Validation is also performed on assaying information such as checks for sampling bias and accuracy of core splits, proper use of pulp duplicates and standards, handling of calculated values, explicit/implicit absent data and data entry errors. The second step involves identifying the mineralized domains resulting from the mineralization process. This step requires a geometric interpretation of all relevant geological features that have influenced the spatial distribution of the mineralization including dykes, faults and alteration. In addition, a geologic reference system rather than a standard Cartesian system is established to measure distances between samples. Once the mineralized domains are established and the database has been validated, statistical and spatial description of the mineralized samples are examined in order to establish the histogram and the variogram of all variables of interest and their correlation. Results are also used to calibrate the global resource estimate within the mineralized domain and to establish the search strategy for grade interpolation. The mineral envelope is filled with polyhedral cells representing mining blocks. The average grade of these blocks is estimated using kriging techniques in order to eliminate global bias and to minimize the local error in the block grade estimate. Visual checks are made to validate the spatial trends in grade distribution while global bias and smoothing effect are assessed from the declustered sample statistics. When required, smoothing correction may be applied using ?change of support? techniques. Simulations are used to perform sensitivity analysis and the simulation models are used if severe smoothing of spatial variability makes the interpolated model inappropriate for mine planning. The keys to successful resource estimation are an integrated team of geologists, geophysicists, geostatisticians and mine engineers using the best available technology and reviews by independent auditors. The resource team is also responsible for performing a QA/QC program in order to measure the effectiveness of the methods in meeting the specific objectives at each step of the modelling process.

May 1, 2001

By P. Hansman

"Place Dome Canada&apos;s Detour Lake Mine is a low-grade gold producer located in northeastern Ontario. Production is targeted for 2850 tonnes of ore per day. As reserves and mining areas continued to be found below and far to the west of the shaft (Fig. 1), it became necessary to review and upgrade the materials handling facilities. This paper deals with the selection, design and construction of an underground crusher and conveyor system capable of handling up to 450 tonnes per hour from the lowest existing level in the mine.BackgroundThe Detour Lake mine site is accessed via secondary Highway 652 which runs northeast out of Cochrane, Ontario. The road consists of 200 Ian of hard top followed by 100 Ian of well maintained gravel to the mine site. A 1500 m airstrip is maintained at the mine site for emergency use. Prior to 1991, the workforce was flown to and from the mine site. Crews were now bused to the site from Timmins. Remnants of a winter road exist to the southeast which at one time ended in La Sarre, Quebec. This was the route used to haul bulk items during the construction period and early production years.A town site has not been established at the site for the current operations. Instead, a rotational scheme is being utilized whereby most personnel will work seven days on site followed by seven days off site. For some employees, however, the work week is termed as a four/three schedule where they work 42 hours between Monday at 10:00 a.m. and Thursday at 4:00 p.m. with the following three days off. All surface employees and maintenance employees underground work 12-hour shifts. All underground employees work a lOV2-hour shift. Supervisors work a similar schedule as their crews, however, other staff may overlap their rotation during the week.A five-compartment shaft was sunk in 1984 from surface to the 610 m elevation. Shaft stations were established at the 120 m, 260 m, 360 m, 460 m, and 560 m levels. The cage compartment consists of a single deck, 4O-man cage in balance with a counterweight. The cage hoist is an Ingersoll-Rand 10H and is regulated to travel at 4.1 m/s. Two 9 t skips operate in balance at a travel speed of 7.1 mls. The skip hoist is a 3.05 m drum Nordberg. Both hoists are automated to function without a hoistman during normal skipping and cage service runs."

Jan 1, 1995

By Jonathan Peck, Jim Gray

"With the implementation of more advanced monitoring, control and locating systems in surface mines, there is an ever increasing need for their mine-wide integration toward achieving over-all optimization. AQUILA Mining Systems Ltd. is developing the basis for a comprehensive real-time monitoring, control and information system called the Total Mining System or TMS. This system is being designed to address both current and future needs of open pit mines toward improving productivity and minimizing costs, through better two-way information flow between each component in the operation. At a basic level, feedback from equipment and process monitoring will enable decision making to be based on quantitative, on-demand information, thus permitting proactive rather than reactive planning. Dynamic planning would be a valuable tool as market conditions fluctuate, equipment fleets age and variations in the grade, quality, geometry and hardness of the mined material all contribute to either changing production targets or levels of achievable performance. Such a capability could be implemented within the next two years.At a higher level, and within five years, autonomous operation and/or real-time supervisory control of mobile and stationary equipment in open pit mines will be achieved. A discussion is provided in regard to the development and implementation of an autonomous haultruck and blasthole drill, to demonstrate the need for such a system.This paper outlines a preliminary program for the design, development and implementation of this system, primarily in relation to high-level communications and its role in the future automation of open pit mobile equipment."

Jan 1, 1995