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By Herman D. Mendoza

"The removal of finely dispersed mercury from the effluent of small-scale gold mining operations and the eventual recycling of the recovered mercury was investigated. The conventional method of amalgamation, or “ball milling”, involves the mixing of a substantial amount of mercury with ore to recover free gold during the grinding stage. During grinding, a portion of the mercury is transformed into minute droplets, which readily float and disperse in the effluent. A kilo of gold ore containing 20-30 g Au/t, when ground with 3 g of mercury, yields 0.005 to 0.050 mg/l of mercury in the effluent, exceeding the concentration of 0.002 mg/l allowed by law for surface water.A rapid recovery/ recycling process was introduced using the surface-chemical properties of finely dispersed mercury in aqueous suspension. It was observed that fine droplets of mercury in an aqueous environment readily attached onto a collector having positive surface potentials for pH 3 to 10. An example of a suitable collector is alumina (Al2O3), which has an iso-electric point of pH 9.8. When an effluent, coming from a settling procedure after an amalgamation-grinding stage, was passed through a column-bed packed with alumina fiber, removal and subsequent collection of finely dispersed mercury occurred under acidic conditions, pH 3-6. Recovery was around 99.5%, yielding an acceptable quality of water safe for disposal into the aquatic environment. The collected droplets of mercury, which are left in the column-bed of alumina fiber, were subsequently treated with a cleansing solution at alkaline conditions pH 10 -12. At these conditions, both mercury and alumina fiber carry negative surface potentials, which results in the eventual detachment of attached mercury from the alumina collector. The effective recovery of attached mercury from the alumina collector by the detachment procedure is only 30-35%. The collected mercury from the cleansing stage can be collected, cleaned and can be used for subsequent amalgamation processes. The attachment and detachment behaviour of finely dispersed mercury onto and from the microscopic collector alumina is shown in Figure 1. The result of attachment and detachment were analyzed using existing interaction energy theories like the Darguin-Landau- Verwey- Overbeek (DLVO) Theory."

Jan 1, 2003

By Güven Önal, Zeki Dogan, Neset Acarkan

"In general, the processing of glass sand may involve comminution, washing, wet classification and attrition scrubbing. Additional processes may include magnetic separation, flotation and leaching in various combinations.Yaliköy sand dune deposit is located near to Istanbul on the European side by the Black Sea coast. The reserve of deposit is around 70 million tons. Iron content of Yaliköy sand dunes is 0.48 % Fe2O3 and not suitable for glass manufacture. The iron content of glass sand should be less than 0.15 % Fe2O3 for window glass manufacture and 0.05 % Fe2O3 for glassware. In this study, it was aimed to produce glass sand with less than 0.15 % Fe2O3 for window glass and with less than 0.05% Fe2O3 for glassware manufacture from Yaliköy sand dunes. A representative sample with a weight of 200 kg was taken from Yaliköy sand dunes deposit through opening small pits.• Mineralogical studies have shown that iron is present in three forms:• Hematite, limonite, clay and iron silicate minerals such as epidote and garnet occur in free form.• Quartz grains are coated by hematite and limonite.• Iron is also present in the cracks of quartz grains.Free iron minerals such as hematite and limonite and silicate minerals were subjected to magnetic separation and flotation for removal. Iron coatings were tried to be removed by scrubbing and leaching. Iron removal in the cracks of quartz grains was attempted by grinding and leaching.Under optimal conditions, sand concentrate with 0.38 % Fe2O3 was obtained by means of high intensity disc type magnetic separation. Grinding and scrubbing were applied to see the effect of size reduction and attrition at room temperature and 90 ºC, by using non-magnetic fraction. Sand concentrate with 0.13% Fe2O3 was obtained at 90 ºC in this test.Sand sample which was previously ground to minus 0.5 mm and scrubbed, was floated by using sulphonates as collector. After flotation tests, sand Concentrate was obtained with 0.11 % Fe2O3.The optimum iron removal was achieved by H2SO4 leaching at 90 ºC depending on acid concentration and leach duration in static, percolation and agitation leaching. Agitation leaching was done by the use of mixer at 600 rpm, 50% solids and leach duration was 15 minutes."

Jan 1, 2003

By G. M. Savov

Copper recovery via Solvent Extraction (SX) process is a well-known technique, which is practiced commercially for approximately 40 years. The main advantages of the SX process are: an absence of using high temperatures (not more than 50oC), complete recycling of barren solutions, very high selectivity at removal and purification of metals, metal conversion to a form, suitable for easy subsequent recovery. Despite the wide acceptance of Solvent Extraction, it does have some shortcomings, namely SX employs large quantities of volatile organic substances. This leads to huge plant ?footprint? of SX circuits. The SX is very sensitive to low temperatures and quantity of total suspended solids in feed solutions and this causes bad phase disengagement times and therefore poor process performance. And most important ? low grade solutions (with copper concentrations below 1 gram per litre) are very unattractive for treatment via Solvent Extraction process, because capital costs for SX plants are not a direct function of copper tenor in the feed. The combined process with Ion Exchange (IX) and Solvent Extraction ? further called IX/SX by IONTECH Engineering was developed. This combination eliminates SX negative characteristics shown above. The key findings of the IX/SX process are; (1) recovery of copper from pregnant leach solutions with metal concentrations, varying on a broad range; (2) can be operate effectively with very low-graded copper pregnant leach solutions; (3) the construction of large ponds for pre-filtration of the pregnant leach solutions is not necessary; (4) working with turbid solutions and solutions containing large quantities of suspended solids; (5) winter operation without troubles; (6) strongly reduces organic losses; (7) technological approach lead to plant sizes reducing. This paper reviews the main stages of the combined IX/SX process, the applications for metal recovery and introduces projects where the process is applied. Key findings of the process are also presented. A comprehensive technical comparison between processes has been carried out. This paper also evaluates the advantages of using combined IX/SX technology as a purification and separation step in copper production process, with the aim of further increasing the overall copper recovery and to ensure that a suitable copper concentration in the advanced electrolyte is achieved. The implementation of a Ion Exchange (IX) process in the classic Solvent Extraction-Electrowinning process(SX/EW) is found to be beneficial for processing of pregnant leach solutions both from spent and new copper heaps and dumps, acid mine drainage and other industrial streams without any preconditioning as well as in terms of achieving improved recoveries and copper cathode quality.

Sep 1, 2012

By Haghi. Hamed, Javad Vazifeh Mehrabani, Mehdi Sedigh

In this investigation, the possibility of physical separation methods such as attrition scrubbing, hydro-cycloning and also reverse flotation were investigated for silica and calcite removal from relatively low grade kaolin. The head sample was provided by the Iran China Clay Company and the major oxides contained in the head sample were Al2O3, SiO2, CaO and Fe2O3 with grades of 15.07%, 71.05%, 4.35% and 0.52% respectively. The liberation degree of calcite and iron impurities from kaolin particles were over 80% for particles finer than 210µm and 53µm respectively. Attrition scrubbing tests were carried out with scrubber cell of the conventional Denver flotation cell model D-12. Laboratory scale hydro cyclone was applied for the beneficiation of kaolin. Attrition scrubbing tests with control screen of 106µm did not achieve the desirable yield for purification of this kaolin sample, but hydro cyclone experiments resulted in better results. Hydro cyclone with 2.54 cm diameter used and the only variable was the cyclone inlet pressure ranging from 68.65kPa to 98.07kPa. The best inlet pressure was found to be 68.65kPa. Based on the optimized condition, Al2O3 and CaO amount in the hydro-cyclone overflow was 23% and 1.47% respectively. However weight recovery of the overflow in each cyclone experiment was less than 30% and multi stages of hydro-cyclones would have to be applied for higher weight recovery and better yield. All of the flotation tests were carried out in the conventional Denver flotation cell model D-12. Reverse floatation tests for calcite removal were carried out under alkali pH based on full factorial design of variables in 2 levels including one center point. The three parameters tested were collector dosage (gr/ton), pH and depressant dosage (gr/ton). The collected data from flotation tests were analyzed and optimized collector and depressant dosage were 1000 gr/ton and optimized pH was 8.5. Under the above mentioned conditions, calcite removal was about 75% and CaO content in the final product was reduced from 4.35% to 1.33%. However calcite removal will increase to more than 83% if the feed particle size in flotation tests is reduced from 210µm to finer than 53µm. Finally after optimization, the best working condition and possible flow sheet were developed.

Jan 1, 2014

By Thi Hien Tran, Ngoc Anh Tran, Thi Thu Hien-Dinh, Duy Anh Dao

"Natural graphite exists mainly in three types of amorphous graphite, flakes graphite and lump graphite. Graphite from Bao Ha mine, Lao Cai province is with 11.8 carbon-graphite %, forms in flakes and dense crystals. The paper presents the comparison among enrichment of Bao Ha’s graphite by two types of froth flotation with 90 g/t kerosene oil as collector, 50 g/t montanol 800 as frother, in natural pH, and at 20 wt.% of solid. The differences among three flow charts indicated that using middling flotation and re-middling flotation for the cleaning tails and the scavenger concentrates, the final concentrate was improved and containing up to 94 carbon-graphite % of flakes graphite, up to 82 carbon-graphite % of fine graphite with 93.48 % total recovery, moreover, carbon-graphite in the final tail is 0.86 % with 6.52 % recovery. However, a refining mill and a classification stage were applied before doing the middling flotation stages.1. INTRODUCTIONGraphite (black lead or plumbago) is used in many industrial applications due to its high electrical and thermal conductivity, lubricity, and chemical inertness. High quality graphite can be used in metallurgy, molding, brake linings, batteries, and fuel cells. Graphite, composed exclusively of the element Carbon, is rather common mineral, but fine crystals are rare. Graphite is soft (hardness: 0.5-1 Mohs scale); has a layered, planar structure; occurs in platy or acicular (occasionally) morphology (Minerals.net; Pierson, 1993).USGS (2017) reported, global consumption of natural graphite continuing increase in the period of 2012-2016. In 2016, China used up 35% while produced 66% of the world’s graphite. Graphite mining are being developed in Madagascar, Mozambique, Namibia and Tanzania, but production reduced in Canada from 2015. Based on the statistic of Statista.com shows from 2011 to 2014 and a forecast to 2020, the global graphite demand is mostly in refractories, batteries, others, lubricant/crucible, foundry and pencil, and that is increase every year.Graphite Vietnam lies on the fault of the Red river from Yen Bai to Lao Cai, the reserves is estimated around 29.3 million ton and mostly in Lao Cai province where has Nam Thi graphite mine and Bao Ha graphite mine (MOIT, 2008). Vietnamese graphite products could reached to ~92% graphite carbon (Cg) (Tran Thi Hien et al., 2008). Studied samples was collected in Bao Ha graphite mine, is located in Lao Cai province. Graphite Bao Ha exists in two types as graphite flakes and amorphous (Hoang Thi Minh Thao et al., 2017)."

Jan 1, 2018

By J. K. Wen

The Mojiang nickel deposit is situated in Mojiang county, yunnan province, a tropical area in the southwestern China. The known geological resource (measured and inferred) includes 4 million tons of nickel sulphide ore at an average grade of 0.50% Ni and 0.34% As. Nickel minerals are mainly gersdorffite (NiAsS), millerite(NiS), morenosite (NiSO4?7H2O), and more than 50% nickel minerals is gersdorffite which contains arsenic. Pyrite is the dominant sulfide mineral, 22% in the ore and part of nickel minerals are locked in pyrite. It is difficult to attain the nickel concentrate quality allowed by a smelter and the acceptable nickel recovery through flotation due to the high arsenic in the ore and the nickel minerals locked in pyrite. The shake flask tests and column tests for nickel recovery from the Mojiang nickel sulfide ore had been finished by the end of 2003. The column bioleaching tests were carried out using the columns with diameter of 120 mm and height of 1200 mm, and 70% of nickel dissolution was attained in 300 days. A 10,000t pilot test heap was built in the Mojiang Nickel Mine by June 2003. The ore used in the test heap was supplied from the underground mine and was crushed to -20 mm before being feeding to the test heap, which is 4~5 m high, 30 m wide and 50 m long. The naturally adapted mesophile and moderate thermophile have been inoculated into the leaching pile by June 2003. The operating temperature in the heap ranges from 38? to 55?, varying with season. 30% of nickel dissolution was achieved in 3 months and the typical pregnant leach so-lution (PLS) contains 2~3 g/L Ni and 20~30 g/L Fe. The Ni sulfide which was produced by us-ing Na2S 10 t/t?Ni contains 16%~20% Ni and Ni recovery was 98%. 70% of Ni dissolution is expected in 12 months. The test heap was scaled up to 60,000 t of ore per year by mid of 2005. In this paper metallurgical studies and performance of the pilot test heap will be described. The results of mineralogy, shake flask tests and column tests on bioleaching of the nickel sulfide ore are reviewed. The key test heap operation parameters, the performance of bio-heap leahing-Na2S precipitation, ferrous oxidation-ferric precipitation-Na2CO3 precipitation, bacteria in the pile, water and acid balance, effluent disposal, capital and operation costs are also discussed.

Jan 1, 2014

By John Starkey

"The development of the SAG Power Index (SPI) Test in Canada has created a new and accurate way to measure ore hardness in a SAG environment. Coupled with the standard Bond Ball Mill Work Index Test it is now possible to determine from these two relatively inexpensive laboratory tests, the amount of energy required to grind an ore from the primary crushed size to the desired liberation size. By testing multiple samples to discover where the hard and soft ore zones are located, the representative hardness distribution for the ore body can be determined and grinding equipment selected that is capable to meet production targets for all mining areas of the deposit.Previously, it was deemed prudent to select up to 50 t samples of ore for Semi Autogenous Grinding (SAG) pilot plant tests and a second 5 t sample for metallurgical pilot plant work. The methods outlined in this paper show how the 50 t sample is no longer required for SAG design because it is better to measure the representative hardness distribution of the ore body than the hardness of a “chosen” representative sample. This is demonstrated to be true by the fact that the lower horizons of an ore zone cannot initially be accessed to recover 150mm pieces, and the bulk samples taken, rarely represent the desired design hardness of the ore. Median hardness will not yield a good result in most cases and a clear understanding of the location and extent of the hardest zones is imperative to achieve a truly successful grinding circuit design.This paper discusses two plants (Kubaka and Agnico Eagle) that have been designed using SPI and Bond with no pilot plant test backup. Both have been very successful in that they can grind design tonnage on the hardest ores in the mine. This guarantees that the mine’s profitability is protected in any mining sequence. It also allows the extra power to be converted to higher throughput on softer ores if additional cash flow is needed. The advantages from the owner’s point of view are dramatic. This procedure will be attractive as well when considering the costs for test work, and the samples for testing. Existing diamond drill core is normally used and this allows the inclusion of data from the deepest regions of the mine. In addition, it is also clear that the decision to use a crusher can now be made at the design stage of a project based on measured parameters and the estimated capital costs of alternative designs."

Jan 1, 2003

By H. Breuer

Mineral processing is characterized by a constant adaptation to changing raw materials and market conditions. It is the link between the mined raw material and a marketable product. As a lot of high grade reserves are exploited, a steady deterioration of raw material quality can be observed. At the same time, the customers´ requirements for product purity and consistent quality increase. Indian iron ores are characteristically friable in nature that results in generation of sizeable quantity of fines (0-10mm) every year during various stages of mining and processing in the country with grade varying from 50-60% Fe(T) and are high in silica/alumina (Al2O3 5-8%). The fines generated cannot be directly used for iron making (blast furnace/DRI units) in the country due its physical and chemical properties; therefore, large portion of it gets exported or dumped. Value addition to iron ore fines through various activities such as beneficiation & palletization is the need of the hour, which will not only be economically beneficial to the mining sector in the long run but also generate value addition & employment opportunities. Incentives may be given for adopting latest technologies for direct use of fines in iron making by agglomeration of fines in the form of Pellets and Sinters, as fines forms considerable part of iron ore resources. Due to selective mining, although the iron content of the ore is high but the alumina(Al2O3) content in iron ore is also significantly high which demands for beneficiation as high alumina will create problem during iron making like higher fuel consumption, poor quality of slag produced, reduced productivity of B.F. all mineral has been engaged in hematite iron ore beneficiation with its gravity and magnetic separators since the mid-nineties. The delivery of jigs started for the upgrading of iron ore for its utilization in a Direct Reduction Plant in Australia. Since then, various other installations with jigs for lumps and fines as well as upstream separators for fines are in operation in Brazil, Australia, India and South Africa. Low grade run of mine and/or dump ores are being processed with alljig®-, allflux®- and gaustec®-WHIMS as the core equipment. The article describes the technology in use, the characteristics of various iron ores and the product qualities achieved. The data presented shows the specific advantages of jig application on iron ore upgrading due to the possible high gravity cuts and the easy and low cost operation. The allflux®- technology provides the big advantage of high capacities and a two-step process, both in a single unit. The gaustec®-wet high intensity magnetic separators offer high capacity in combination with unsurpassed process flexibility. Keywords: iron ore, gravity separation, jig, upstream classifier, wet high intensity magnetic separation (whims), pelletization, sinteringMineral processing is characterized by a constant adaptation to changing raw materials and market conditions. It is the link between the mined raw material and a marketable product. As a lot of high grade reserves are exploited, a steady deterioration of raw material quality can be observed. At the same time, the customers´ requirements for product purity and consistent quality increase. Indian iron ores are characteristically friable in nature that results in generation of sizeable quantity of fines (0-10mm) every year during various stages of mining and processing in the country with grade varying from 50-60% Fe(T) and are high in silica/alumina (Al2O3 5-8%). The fines generated cannot be directly used for iron making (blast furnace/DRI units) in the country due its physical and chemical properties; therefore, large portion of it gets exported or dumped. Value addition to iron ore fines through various activities such as beneficiation & palletization is the need of the hour, which will not only be economically beneficial to the mining sector in the long run but also generate value addition & employment opportunities. Incentives may be given for adopting latest technologies for direct use of fines in iron making by agglomeration of fines in the form of Pellets and Sinters, as fines forms considerable part of iron ore resources. Due to selective mining, although the iron content of the ore is high but the alumina(Al2O3) content in iron ore is also significantly high which demands for beneficiation as high alumina will create problem during iron making like higher fuel consumption, poor quality of slag produced, reduced productivity of B.F. all mineral has been engaged in hematite iron ore beneficiation with its gravity and magnetic separators since the mid-nineties. The delivery of jigs started for the upgrading of iron ore for its utilization in a Direct Reduction Plant in Australia. Since then, various other installations with jigs for lumps and fines as well as upstream separators for fines are in operation in Brazil, Australia, India and South Africa. Low grade run of mine and/or dump ores are being processed with alljig®-, allflux®- and gaustec®-WHIMS as the core equipment. The article describes the technology in use, the characteristics of various iron ores and the product qualities achieved. The data presented shows the specific advantages of jig application on iron ore upgrading due to the possible high gravity cuts and the easy and low cost operation. The allflux®- technology provides the big advantage of high capacities and a two-step process, both in a single unit. The gaustec®-wet high intensity magnetic separators offer high capacity in combination with unsurpassed process flexibility. Keywords: iron ore, gravity separation, jig, upstream classifier, wet high intensity magnetic separation (whims), pelletization, sintering

Sep 1, 2012

By Kazutoshi Haga, Atsushi Shibayama, Altansukh Batnasan

"The aim of this study was to develop a hydromellurgical processing for removal of arsenic from high-arsenic copper sulfide ores and concentrates, providing increase in copper recovery. In this study, three different enargite ore samples containing 28.2–46.7% copper, 1.9–15.0% arsenic and 10.6–27.5% iron were used to determine optimum conditions for the extraction of copper and removal of arsenic. High pressure oxidative leaching process was preformed using various leaching reagents such as sulfuric acid (H2SO4), ferric sulfate (Fe2(SO4)3), hydrochloric acid (HCl), ferric chloride (FeCl3), sodium sulfate (Na2SO4), sodium chloride (NaCl) solutions and distilled water (DW), respectively, under the fixed conditions which were selected to be as follows: NaCl concentration, 0.34 M; concentrations of other leaching reagents, 0.1 M; particle size, -45 ?m; pulp density, 100 g/L; agitation speed, 750 rpm; leaching temperature, 160 ?C; leaching time, 60 min and total pressure (oxygen gas and vapor pressure) into an autoclave, 1 MPa. Results show that under the leaching conditions, more than 90% of copper leaching was achieved with lower than 2% arsenic content as a consequence of the formation of arsenic precipitates in a solution of Fe2(SO4)3–NaCl. It was determined that copper, arsenic and iron concentrations in the pregnant leach solution were 42 g/L, 0.3 g/L and 0.3 g/L, respectively. As a result, it can be concluded that the Fe2(SO4)3 solution in the presence of NaCl or FeCl3 are more efficient leaching media compared to other leaching agents under the high pressure oxidation conditions, because of their high copper recovery and lower arsenic level in the pregnant leach solution.INTRODUCTION Generally, copper is usually produced from copper sulfide ores by a combination of several processes which are flotation and smelting followed by refining [1-3]. Arsenic content in copper sulfide ores tends to increase due to the depletion of high grade copper deposits, hence conventional mineral processing and metallurgical processes cannot effectively treat these ores containing high levels of arsenic to obtain consistently high purity copper concentrates. Because high levels of arsenic in copper ores/flotation concentrates reduces the value of concentrates due to the restriction of smelting process and causes it to be a source of environmental pollution [4-6]. From this point of view, the known way of prevention against high-arsenic copper sulfide concentrates in further metallurgical treatments is to develop economically and environmentally competitive technologies for removal of arsenic impurity from copper sulfide ores before treating the ores/flotation concentrates by the pyrometallurgical routes."

Jan 1, 2018

It is well known that the crushing and grinding steps in the mineral processing of ores represent a substantial proportion of total energy, cost and carbon emission in mining. For instance a recent study calculated that 36% of the energy consumed by gold and copper producing mines in Australia was attributable to comminution (Ballantyne et al, 2012). Understanding this component of a business is essential in a world seeking to lower production costs, improve earnings and reduce its carbon footprint. The issue commands the attention and support of industry leaders in the mining sector, government, equipment manufacture, engineering design and research organizations. Furthermore, several researchers have often stated that globally, comminution processes consume between 3-4% of the world?s electrical energy production (Daniel and Lewis-Gray, 2011). In 2009-10, Australia?s Federal Government Energy Efficiency Opportunity (EEO) participants in the mining sector used 336.5 PJ, representing approximately 6 percent of Australia's total energy use. It is evident that the current practices of breaking rocks in tumbling mills are relatively inefficient (Fuerstenau and Abouzeid, 2002). Research and improved engineering design has established that a range of improved blasting, crushing and grinding techniques and flow sheets may lower project costs and carbon footprint. These include relatively simple process flow sheet strategies such as gangue rejection or pre-concentration prior to size reduction, a better combination of grinding technologies (both current and new systems), and selection of coarser grind sizes where ore and waste mineralogy allow. Yet, despite this systematic and detailed research, the industry remains somewhat reluctant to implement these strategies. CEEC (Coalition for Eco-Efficient Comminution) was established in 2011 to support knowledge sharing and change in an area of high-energy consumption for the global mineral industry. CEEC is a not-for-profit industry company, funded by grants from world wide industry companies, whose mission is to accelerate knowledge and technology transfer in the field of energy-efficient comminution. The key benefit to the mining sector arising from CEEC?s activities will be an increase in the rate at which more energy-efficient comminution data sets, processes, and technologies are developed, demonstrated, shared and applied. This will contribute directly to the industry?s target of reducing operating costs and carbon emissions. CEEC International Ltd and JKTech Pty Ltd teamed together to facilitate the inaugural CEEC Workshop, in June 2012. The purpose of the workshop was to produce a roadmap to provide guidelines for how (1) operations can achieve best EEC practice in the short to medium term, in collaboration with vendors, engineering companies and other service providers, and (2) what R&D is needed to achieve paradigm change in EEC in the medium to long term, e.g. 75% reduction in comminution energy in 10 years.

Sep 1, 2012

By H. Araneda

Throughout Chile?s history, mining has consistently been a leading industry in the country. The 1990s marked the beginning of a boom in Chile?s mining industry, especially in copper mining, due to both foreign and local direct investment in the sector. Mining?s contribution to the GDP was 15.6% at current prices in 2009, and it accounts for 56% of the country?s exports. Chile?s copper market share reached 35% in 2009. The investment portfolio in mining in Chile for the coming decade has been estimated at 90 billion US dollars. The acute scarcity of human resources currently experienced by the mining sector in Chile will be accentuated in the coming decade as a result of this scenario of investments. The situation is, additionally, similar to the one confronted by other countries involved in the mining industry such as Australia, South Africa, Canada, and in Latin America includes, Peru, Colombia and Brazil. A study carried out by Fundacion Chile, ¨Labor Force in the Chile Mining Industry, Diagnostic and Recommendations, 2011-2020? shows the projected labor shortage for the operation of projects in the development stage, considering only the requirements of the process of extraction, processing, and maintenance. The demand, accumulated incrementally, will be around 115,000 technically skilled workers, professionals, and qualified workers. This figure refers to the operational work force, and does not include the human resources that will be required for the prior stage of the engineering and construction of the projects (brownfield and greenfield). The demand for technically skilled workers and professions for the stages of engineering and construction of the projects that have already received investments is estimated to be 190,000 people for the period 2012-2020. The peak of the demand for engineering is 2012, with a figure that surpasses the 12,000 professionals for all of the engineering specialties. The analysis of the formation of human resources at the tertiary level (where tertiary includes technical/vocational and academic degrees at the undergraduate level) for mining and related majors demonstrates that the programs are scarce with a bias toward theoretical contents and that they are excessively long (40% longer than comparable programs in Australia, Canada, and in OECD countries). The completion rates are low, especially in the technical-vocational programs, which have a completion rate of only around 30%. Finally, the rates of attraction of technically skilled labor and professionals towards careers in mining are insufficient to reduce the identified skill gaps. In the case of university professionals such as geologists, mining engineers and engineers in metallurgy, 80-90% will work in mining. In the case of graduates of relevant majors, that cover skills also utilized by other industries, not only mining?such as maintenance workers, operators of fixed and mobile equipment, etc.?the mining sector appears to be unattractive for reasons such as shift schedules, geographical location of work sites, and other factors. On the basis of the available diagnosis a strategy for the sector has been developed (Workforce Development and Skills Strategy) whose objective is (i) to define a Mining Qualifications Framework to identify the critical roles to be filled (ii) to start, with public resources, a ¨fast-track? training program to assure the availability of at least 28,000 operators (machinery operators, truck drivers, etc.) and maintenance specialists in 2015 (iii) to promote the development of training hubs for mining with learning technologies to train technical labor and professionals on the basis of world class standards (iv) to design and implement a campaign oriented toward the perception of mining as an attractive field of employment opportunity (v) to develop curricular innovations for tertiary education programs, on the basis of international benchmarks. Keynotes: Chile, mining, copper mining, skilled human resources

Sep 1, 2012

By A. Kurkov

Flotation process is based on the formation of primary collector-based supramolecular associates at the mineral surface and within the pulp. Supra molecular chemistry allows studying within a common framework, all kinds of molecular associates, from dimers to large assemblies, resulting from spontaneous association of a large number of interacting components of the flotation system, where the collective action of non covalent bonds can have an important role. Self-organization of supra molecular assemblies in the flotation process can be controlled by modifying the chemical structure of the collector, the nature and arrangement of functional group in it, by introducing secondary agents (SA) into the process, to adjust the interaction between the collector and the surface and to control the hydrophobicity of the adsorption bed. This offers broad opportunities to build supramolecular assemblies to improve the efficiency of flotation. Besides, the more complex is the system, the weaker are the interactions, that determine its behavior. This powerful factor still is not employed deliberately to control the flotation process. Organophosphorus compounds (OPC) - dialkyl phosphoric acids whose radicals contain ethoxylated groups ? are considered as nanocompounds, that can be used in building supramolecular assemblies. The presence in an OPC molecule of functional groups (P=O, P (O) O-, P (O) OH), ethoxylated groups, and aromatic rings is the basis for the formation of stable supramolecular associates. Gross formula: C54H94KO16P. Molecular weight is 1069.39. Phosphoryl anion has a regular tetrahedral structure with the angle of 109.5°. Lengths of -P-O- bonds in a PN molecule are 1.7345Å; length of the hydrocarbonic radical C42-C72 = 33.942Å. In an alkaline environment, PN is in the form of the anion, in an acidic environment ? in the form of the polycation due to protonation of the oxygen in the EO group. Chem Office software suite was used to generate three-dimensional images of an OPC molecule in acidic and alkaline environments, to reveal the OPC geometry (bond lengths, valence angles, aliphatic radical length), to identify the collector's binding sites with the mineral surface, to find the most stable and energy-efficient molecular conformations of an OPC molecule and its complexes with minerals under various conditions. Mechanisms of interaction between OPCs and the mineral surfaces are different. In ores whose minerals are dominated by ionic bonds (fluorite), the interaction mainly involves the anionic form of the collector; in ores, whose minerals are dominated by covalent bonds - the oxygen of the phosphoryl group (donor-acceptor interaction) and the oxygen of the ethoxylated groups. This difference is responsible for the formation of spatial associates with different surface layer structures. Introduction of supramolecular chemistry concepts into the analysis of flotation phenomena and mechanisms can become an effective tool for better understanding the flotation process and offers new means of consistently improving its effectiveness.

Sep 1, 2012

By V. Ross

"The Upper Group 2 (UG2) reef in the Bushveld Complex in South Africa hosts an estimated 40% of the world’s Platinum Group Metal (PGM) resources. The processing of this ore via flotation poses two particular challenges: limiting the levels of chromite in the concentrate to reduce downstream smelting problems, and enhancing the recovery of ultrafine (typically -10 micron), liberated PGMs. Conventional ways of suppressing the entrained chromite, such as froth washing, is not effective in PGM operations, which are dominated by mechanical cells. The recovery of fine particles is hampered by low collision efficiencies with air bubbles, as well as oxidised surface layers especially in tailings operations. This paper documents the results of laboratory tests on a run-of-mine PGM UG2 ore, investigating the application of a pneumo-mechanical column (PMC) and its effect on PGM flotation and its selectivity with regard to chromite. A key feature of the device is the stage-wise release of aerated slurry from an agitated downcomer, thus creating conditions where the superficial velocity profiles of air and slurry change axially within the separation zone. This aids segregation of the silicates and chromite-rich fractions in the separation zone, concentrating the latter towards the bottom of the column and minimising its entrainment into the concentrate. Comparison of the performance of the PMC to that of mechanical laboratory cells under equivalent operating conditions showed improved PGM flotation kinetics and recovery. This was accompanied by a decrease in gangue recovery compared to the mechanical cells by more than 40%, thus doubling the final rougher concentrate grade from 50 to 100 g/t. Significantly, whilst the overall Cr2O3 recovery was reduced by more than 60%, the selectivity (PGM recovery vs Cr2O3 grade in the concentrate) was doubled.INTRODUCTIONThe Bushveld complex in the northern part of South Africa hosts the world’s largest reserves of Platinum Group Metals (PGMs), representing approximately 40% of the total PGE. It is characterised by three main PGM bearing reefs, namely Merensky, Upper Group 2 (UG2) and the Platreef. The Merensky and Platreef are dominated by silicates in the form of pyroxenites, and carry economically recoverable amounts of copper and nickel, in the form of chalcopyrite and pentlandite. The UG2 on the other hand is a chromitite ore with chromite as (ideally, FeO.Cr2O3) the most abundant gangue, often constituting more than 70% by mass with a PGM grade of around 4.5 g/t. Other key gangue minerals are pyroxene, feldspar and talc. The PGEs occur in a wide variety of minerals, with platinum and palladium sulphides typically accounting for about half of the total. Base metal sulphides such as chalcopyrite, pentlandite and pyrrhotite occur in only very minor amounts, accounting mostly for less than 0.1% by mass."

Jan 1, 2018

By S. V. Petrov

Oxidised ores possess a significant share among raw materials which are treated by traditional plant technology. As usual, the chemical composition of these ores is rather simple. They can be treated by standard cyanidation. In this case, the resulting wastes have an alkaline reaction. Their toxicity (in case of ROM ores treatment) is mainly due to the presence cyanides and thicyanades. The tails are dexotified before disposal. However, chemical reagents used cannot provide a complete removal of toxic matters. In this case, this disposal poses a real problem because these tails contain a significant amount of CN and SCN and their behavior during a long disposal should be studied. A sample with a simple composition taken from cyanidation tails was studied. The sample presents itself the quartz low-grade sulfide gold-bearing ore. Sulfide sulfur content is 0.15 per cent,and heavy metals content is 0.023 per cent. Method of testing is described further. The samples were tested under moderate and subarctic climates based on the difference in temperature, moisture and the degree of oxygen saturation of air. The pH of water extract, which was prepared using distilled water, was also determined besides measuring the concentration of cyanides and thiocyanates. This demonstrates the stability of sulfide-bearing minerals in the disposed tails. The tests time was 820 days. It was found that the alkalinity of the samples tends to lower under oxygen-containing conditions in proportion to temperature increasing. Under conditions without oxygen, the pH lowering is minimal. Mineralogical and chemical properties of the ore material give positive alkaline balance during testing. The solid sample contains enough alkaline agents to compensate dangerous to the environment reactions of sulfides oxidation.pH value maintaining under alkaline conditions witnesses to the possibility of a safe disposal of this tails type. The largest amount of cyanide is removed under acid-containing conditions at positive temperatures. The higher the temperature is, the more intensive destruction is. At the moment, of finishing the tests at 20ºC temperature, the residual concentration of CN- was 0.06-0.10 mg/L, at 4ºC, the residual concentration of CN- was 0.2-0.3 mg/L and at -18ºC was 0.4-0.6 mg/L. Under conditions with the lack of oxygen, cyanides can be removed within the first 200 days. Then, this process fully stops. A stabilized cyanides concentration is 3.0-4.0 mg/L. The presence of sulfides increases the content of cyanides at the initial stage of testing. Then, the concentration of SCN at the temperatures above 0ºC decreases rapidly under oxygen-containing conditions. At the rise of temperature, a narrowing of peak at the curve and an increase of the destruction process are observed. At the end of testing, the concentration of thiocyanates in the moisture of wastes was 0.2-0.3 mg/L at the temperature of 20ºC and 0.5-0.6 mg/L at the temperature of 4ºC, respectively. The concentration of thiocyanates contained in a frozen sample and media which have the lack of oxygen stabilizes at a certain level owing to sulfides and does not practically change further. It was found that detoxification of oxidized cyanidation tailings with a simple composition which were disposed as filtration cakes can be possible. In this case, the concentration of cyanide in the moisture of disposed wastes can have a close value or equal to actual concentration during cyanidation. In order to meet the requirements of high ecological safety standards and successful detoxification process of toxic compounds in the tails under the action of natural factors, the conditions of air aeration with oxygen for the whole volume of tails are required. Keywords: cyanides, thiocyanates, storage, detoxification, flotation cakes

Sep 1, 2012

By J. J. Lui

Surfactant plays a critical role in many industrial processes such as petroleum, mining, environment and food processing, and daily life. For ionic surfactant systems, adsorption of surfactant at oil-water interface imposes a net charge to the interface. The nature and magnitude of the charges at the oil-water interface depend on charge characteristics of surfactant molecules, concentration of surfactant and solution pH. In our previous study, zeta potential measurement of hexadecane drop-lets in industrial process water was shown to be a simple, yet powerful method for identifying various types of ionic surfactants present in the aqueous system. In this paper, the site-binding model was applied to prediction of zeta potential of oil droplets in aqueous solutions containing a mixture of cationic and anionic surfactant molecules. The total number of interfacial sites for various types of surfactant involved was determined by applying the Gibbs adsorption equation to the measured interfacial tension-surfactant concentration profiles. The distribution of various types of surfactant molecules at the interfaces was evaluated by Langmuir isotherm constants. With the ionization constant of surfactant taken from the literature, the predicted zeta potentials of oil droplets in surfactant solutions from site binding model were shown in a reasonable agreement with the measured zeta potential values.

Jan 1, 2014

By Henrique Kahn, Daniel Uliana, Carina Ulsen, Guilherme Nery

"Gold has been a material in high demand since the dawn of humanity. After hundreds of years of exploitation, the increasingly search for gold occurs in lower grade deposits with greater extraction difficulty. Therefore, for maximum efficiency, deep knowledge of the characteristics of these minerals and their associations is crucial, which is acquired through an accurate and rigorous routine of process mineralogy and batch extraction tests. The industrial gold process extraction occurs by gravity concentration, sulphides flotation and cyanidation. For gravity concentration, it is fundamental to establish the proportion of free or coarse gold grains and its association to heavy minerals; regarding flotation, the most relevant data are the proportion of gold associated to sulphides and their liberation by the surface area; for cyanidation, it is crucial to evaluate gold accessibility and the kinetics of leaching. Thus, the characterization of gold associations and liberation is vital to establish the plant design, to optimize operating plants and to reduce gold losses. Our aim is the characterization of gold occurrence in low-grade Brazilian deposits with complex mineralogical assemblages to support processing flowchart definition and to predict metallurgical recoveries. Gold distribution in light and heavy minerals was evaluated by heavy liquid separation and the occurrence of gold in free or locked particles was carried out on heavy products by SEM-based automated image analysis using the MLA system. All the products of heavy liquid separation were cyanidated in head samples with different milling conditions to evaluate the reaction kinetics and maximum recovery."

Jan 1, 2014

By X. W. Lv

The intelligent recognizing and processing system for sinter ore mineragraphy make a great convenience of recognition and measurement. Specially, it?s critical important for a successful intelligent system to extract the minerals? feature accurately and sufficiently. Reflection index, as the crucial feature of mineral, is calculated by the theoretical calculating model of index based on the principle of imaging color models. In gray histogram of minerals? micrograph, the peaks? positions represent the information about mineral categories and contents. The relationships of different minerals? peak-positions in curve of gray histogram were got by using the peak-find model of gray histogram and calculating model of reflection index. These two models which can be regard as the method of feature extraction for minerals? photomicrograph also lay a good foundation for the artificial intelligent system of recognition and processing in sinter ore.

Jan 1, 2014

By Y. F. Guo

Laboratory scale experiments were conducted to produce titanium-rich material from Panzhihua ilmenite. The results showed preoxidation with sodium salt additive especially for Na2B4O7 signif- icantly advance the reduction of iron in Panzhihua ilmenite. Acceleration mechanism of preoxidation with Na2B4O7 has been studied here. The maximum metallization of iron 92% can be obtained under conditions of preoxidation at 850? for 30 min and subsequent reduction at 1100? for 150 min. The rusting separation technique to reduced ilmenite has been further studied, obtaining titanium-rich product with TiO2 over 82% and the recovery of TiO2 above 97%.

Jan 1, 2014

By A. N. Zagorodnyaya, Ye. G. Bochevskaya, E. A. Sargelova, Z. S. Abisheva, A. S. Sharipova

"In Kazakhstan, there is a unique sulphide copper deposit, in the ores containing the natural isotope of osmium-187 (more than 90 %). In the pyrometallurgical processing of concentrates for the target metal, osmium is mainly sublimated, with wet cleaning of metallurgical gases (dust and gas mixture) 40 % of it together with rhenium passes into washing solutions (wash acid). On the basis of this acid, the production of ammonium perrhenate was organized at the enterprise of the RSE ""Zhezkazganredmet"" and for the first time in the world practice - the isotope of osmium-187. However, the degree of its recovery from solutions does not exceed 8 % of the amount supplied with the charge or 25 % of the amount that was received for extraction. But the bulk of it (~ 75 %) remains in the raffinate, discharged to sewage treatment plants (wastewater). The studies presented in this work are aimed at solving the problem of osmium isotope-187 extracting from wastewater, which are a potential source of increasing metal production in Kazakhstan. The aim of this work is the development and scientific justification of osmium deposition regimes from wastewater using such reagents as thiosulfate and sodium sulphide. The influence of various physicochemical factors on the precipitation of the osmium was studied and optimal conditions for its maximum transition to sediment were chosen. An original method for obtaining osmium concentrate in the presence of sodium thiosulfate with the addition of potassium iodide from a new type of wastewater - man-made ammonia solutions has been developed. These solutions were combined mother liquors of solid-phase stripping operations and dissolution of pulp of crude ammonium perrhenate. On the basis of the obtained results, a settling technology for extraction of the osmium from man-made ammonia rhenium-containing solutions was introduced at RSE ""Zhezkazganredmet"". It should be noted that the degree of extraction of osmium, depending on its concentration in solutions entering on the settlement, and taking into account the addition of sulfuric acid, will vary from 60 to 90 %. The recovery of osmium during the test period was 58 %."

Jan 1, 2018

The energy dispersion X-ray fluorescence (EDXRF) analysis technique is a rapid and accurate method for describing sample?s element component and percentage composition. On the base of profound research for the national and international development and research condition, the author researched the EDXRF analyze technique and the EDXRF analyzer?s principle in depth, the difficulties and the key points for research and development of EDXRF analyzer; gave a scheme for re-search of bench EDXRF analyzer; researched some algorithms used in energy spectrum analysis; developed the energy spectrum analysis software based on the Visual C++ 6.0; together with the energy spectrum analysis software, the author designed and processed the structural of the bench XRF analyzer and integrated the import X-ray tube, thermoelectrically cooled Si-PIN photodiode X-ray detector and pulse amplitude analyzer as a model machine. The result shows: The model machine can be used in multi-element analysis at the same time for different types of dry mineral samples, with fast analysis speed, accurate model and high stability.

Jan 1, 2014