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By A. Ekrem Yüce, Gülay Bulut, Feridun Boylu, Vecihi Gürkan, Istanbul Technical University, H. MustafaTarkan

"Nickel sulphide ores subjected to this investigation were taken from Orhaneli district where widespread chromite mineralisation can be occurred. Significant amounts of magnetite and chromite have been existed in the ore together with sulphide and oxide type nickel minerals. Ore sample is constituted about 70% magnetite, 15% sulphide minerals, 5% chromite and iron oxides as well as 10% gangue minerals.According to the chemical analysis results, the representative ore sample contains 1.32% Ni, 10.8% SiO2, 78.4% Fe2O3, 1.3 g/t Ag and 1.0 g/t Au. Mineralogical examinations showed that pentlandite (Fe, Ni)9S8, violarite (Ni, Fe)3S4, millerite NiS and asbolane (Co, Ni)MnO2.(OH)2.1-2H2O were the main nickel minerals and magnetite, chromite, chromium spinals, pyrite, pyrhotite, chalcopyrite, cubanite, goethite, hematite, limonite, rutile, chlorite, talc and serpentine group minerals as associated minerals.The aim of this investigation was to produce valuable nickel concentrates by using gravity separation, magnetic separation and flotation methods. As a results of laboratory scale studies, alternative flow sheets for the production of nickel sulphide concentrate were developed. The related experimental results are given in Table 1."

Jan 1, 2003

By Rod Hocking, Nenad Djordjevic, Randolph Pax

"Semi Autogenous (SAG) mills exhibit complex interactions between the competing processes occurring inside them. Traditional practice is to measure the operation of the SAG mill at an operating point in terms of its load, power draw, steel load, ore throughput rate, water addition rate and more recently the feed size distribution. However significant departures of operating conditions can occur due to changes, for example, in ore hardness, the quest for higher throughputs and mill internal geometry.Fundamental to SAG mill operation are the internal processes that ensure that a given ore throughput and discharge grind size are obtained and maintained. These processes can be measured using non-contact acoustic emissions from the SAG mills which utilise the information rich noise that is generated by the grinding, flow and charge movement inside a mill.Non-contact acoustic in-mill measurements can be used for the validation of discrete element modelling (DEM). The approach being reported in this paper is to model the individual physical processes and then to simulated the assemblage of a physical mill."

Jan 1, 2003

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

Results of flotation process combined with atmospheric leaching and high pressure leaching for recovery of copper from low-grade silicate-copper ore have been presented in this study. The low-grade copper ore contains a small amount of copper (0.42 mass %), iron (4.6 mass %), and silicon (38.3 mass%) and among many other metals (Al, Ca, K, Mg, Na, Ti and S). The main mineralogical constituents of the copper ore are calcium aluminosilicate (Ca(Al2Si2O8), amesite (Mg2Al2SiO5(OH)4) and gypsum (CaSO4×H2O), and quartz (SiO2) as resulted from the alteration of an igneous rock. This study has shown that about 92 % of copper can be recovered with a maximum copper grade of 4.2 mass% in a froth after 15 minutes of flotation at pH 8.5 when used potassium amyl xanthate as a collector, MIBC and Dow-froth 250 as frothers and calcium hydroxide as a pH adjuster. The images of bubbles in the cell during flotation process were recorded using a high-speed camera. The froth concentrate with average copper grade of 2.0 mass% was further processed by atmospheric leaching with a solution of sulphuric acid and hydrochloric acid, respectively and high pressure leaching with sulphuric acid solution. As a result, leaching rate of copper reached about 60 % in an atmospheric leaching with sulphuric and hydrochloric acid respectively, whereas the rate can reach more than 70 % by autoclave leaching at high pressure under the controlled condition. A process to recover of copper from its low-grade ore can be developed by optimization of flotation combined with acid leaching.

Jan 1, 2014

By Atul Kumar Bhatnagar

TATA STEEL is the pioneer in coal beneficiation in India. The first coal beneficiation plant was set up in 1951 at West Bokaro and subsequently in 1953 in Jharia. Over the years more beneficiation capacity was added by setting up new coal washeries in both West Bokaro and Jharia, the coal divisions of Tata Steel. currently the beneficiation capacity available is 8.5 mtp aappx Being a captive supplier supplying the coking coal as per demands of our steel works is challenging in the sense that the intent is to leverage the captive raw material advantage and select the best strategy for maximizing value up to steel making. The beneficiation plants were designed to produce clean coal up to 17% ash. With the demand of increasing productivity at the blast furnace the clean coal ash was lowered. subsequently the processes were audited and efficiency improved. The next challenge was to improve the throughputs in the existing plants to maximize the clean coal production. The upgradation was designed in house and implemented to upgrade the plant from nameplate capacity of 2.1 mtpa to 4.5 mtpa while the yearly production targets were also achieved. The presentation will detail the different phases of improvements, process followed, challenges faced, and the results as achieved in the process. Keywords: coal beneficiation, India, coal washeries, beneficiation plants XXVI INTERNATIONAL

Sep 1, 2012

By Xuliang Yang, Yuemin Zhao, Bo Zhang, Liang Dong, Cheng Sheng, Gongmin Li, Zhenfu Luo, Yaqun He, Chenlong Duan

Coal plays an important role in the primary energy consumption of all over the world. In 2012, coal reached the highest share (29.9%) of global primary energy consumption since 1970. Coal burning mainly contributes to the emission of SO2, NOX, Hg and other impurities that jeopardize the future of environment. . Beneficiation before coal utilization is the most economical and effective technology for removing the hazardous material and gangue. In this paper, a novel process of dry beneficiation based on gas-solid two-phase flow composed of the compound dry cleaning separator (FGX), the air dense medium fluidized bed separator (ADMFB) and the vibrated fluidized bed with autogenous medium (VFB) is proposed. The main equipment and the flowsheet of the united system are described in detail. Separating principle of the FGX, ADMFB and VFB devices are all based on gas-solid two-phase flow. In FGX, gassolid suspension is composed of continuous gas flow and coal powder in feed coal. Large size coal (+6- 80mm) can be pre-separated effectively, and large size gangue can be removed effectively. Moreover, - 6mm particles fed serve as autogenous medium in FGX. The probable E value is 0.13 for -80 mm raw coal. In ADMFB, magnetic powder with -0.3 + 0.045 mm size fraction and coal powder in feed coal are compound medium solids. The separation density can be adjusted and controlled according to properties of raw coal. -50 + 6 mm raw coal can be cleaned efficiently, the probable E value is 0.05. In VFB, -6 + 1 mm coal can be separated. Generally, -80 mm raw coal can be cleaned efficiently in the united system based on gas-solid two-phase flow.

Jan 1, 2014

By Cagri Emer

In conventional mechanical flotation cells, when the size of particles exceed a particular value, that is of the order of 100 µm for metal sulphides, the recovery of particles decline rapidly. It appears that the turbulence in the agitated vessel is sufficiently intense to rupture the forces of attraction between the particles and the bubbles, so the large particles are flung off the bubbles in the turbulent field. It has recently been suggested, that a fluidized bed would be a very favorable environment in which to contact bubbles and the particles. In a fluidized bed, particles are supported by rising flow of fluid, which separates them from each other and allows them to move freely while remaining close together. In this novel method, bubbles injected into the fluidized bed are able to rise through it, colliding with hydrophobic particles, transporting them to the top of the column where they can form a froth layer that overflows into a launder. The flow in a fluidized bed is generally slow and nonturbulent, and preliminary experiments have shown, that it is possible to collect mineral particles as large as 1 mm and coal particles as large as 5.5 mm. If the upper particle size limit can be extended, there could be significant reductions in the overall energy consumption in concentrators. Calculations show for example that if the top size for flotation could be increased from 100 µm to 400 µm, there would be a reduction in grinding energy of the order of 40%. It is therefore, important to study the mechanism of coarse particle capture in fluidized beds. In this work, the factors governing the capture rate of particles, particularly the particle and bubble size, the fluidization velocity and the superficial gas velocity have been investigated, and the results will be presented.

Sep 1, 2012

By Kumar Ashish, Shukla Akhilesh, Kr. Rokkam Kiran, M. Sheeba, S. Sombhatla Sundar

"Purification of zinc leachate plays a predominate role in zinc leaching refineries, to produce impurity free cathode sheets. As a part of such purification sequences, copper cementation with zinc dust is the preliminary step, since the presence of copper will preferentially deposit during zinc electro deposition and has the effect of reducing current efficiency. The cake thus produced is being accumulated as an inventory, due to invariable market and leading to space constraints & hazardous ambience. Hence, a project has been taken up for in-house generation of copper sulphate solution through ammonical leach followed by solvent extraction process and utilizing it as an active promoter for zinc sulphide flotation in Hindustan Zinc’s own captive mines. Cold purification cake majorly contains copper (30-40%) along with zinc (10-20%) and cadmium (2%). Confirmation to the above statement, photomicrographs of purification cake also confirmed the major presence of copper in metallic or oxidized form. In present work ammoniacal carbonate leaching of purification cake followed by solvent extraction using diketone based solvent has been studied under the influence of various parameters viz., temperature, agitation, pulp density, ammonia, CO2 dosages, solvent concentration, and impact of w/s zinc on leaching and solvent extraction. Leaching kinetics determined based on shrinking core model. Chemical reaction at unreacted core was found to be the rate controlling step. The estimated activation energy was found to be 24 KJ/mol. Leached copper has been extracted by solvent extraction with a ß diketone based solvent and is stripped with sulfuric acid as concentrated copper sulfate solution. The above established R&D findings are successfully implemented in the Commercial plant with a treatment capacity of two tons of cold purification cake per batch. Till now 500 MT of Purification cake has been treated, and generated 6500 MT of copper sulphate (55 g/l Cu), with final Cu recoveries between 87-90%."

Jan 1, 2018

By Bruce Mathema, Steven Huynh, Terry Cheng, Gerry Gomez

"The use of PM40N, a specialty precious metals promoter in conjunction with a standard Xanthate based collector was extensively investigated in the flotation of a complex, silver-bearing, lead-zinc ore containing naturally occurring graphitic gangue. The ore, considered difficult to treat with conventional reagents, was tested at plant scale to determine the impact of the promoter on silver recovery, lead grade and also on the amount of zinc misplaced to the final lead concentrate.Initial laboratory-scale rougher flotation testing implied significant improvements in the recovery of silver and lead with a marginal amount of zinc displaced. Subsequent mini pilot flotation tests further revealed that a gain in selectivity and recovery could be achieved by adjusting PM40N promoter dosage and incorporating a cleaner flotation step, further improving rejection of sphalerite. Plant trials resulted in an improvement of 3% in Silver recovery coupled with a 2% increase in recovered lead whilst maintaining lead grade in the lead circuit.To date, the use of PM40N as a promoter has enabled the processing operation to significantly increase its plant economics. The partial replacement of the standard Xanthate based collector has also reduced the overall quantity of Xanthate required in the standard circuit by approximately 10% - which is a positive impact on the environment. It is envisaged that further optimization of plant parameters will favor stable operation and continue to boost revenues generated by subsequent improvements in silver metallurgy."

Jan 1, 2014

By Jiang Wenli, Zhang Yimin

In order to enhance the quality of iron concentrate from magnetic separation, reverse-flotation technology is often applied in an iron ore dressing plant.However, this technology has disadvantages in terms of operating cost,energy consumptionand environmental protection(Zhao, Wu& Wang,2005).Developing innovative equipment to process magnetite ore more-effectively is a way to get rid of these disadvantages of reverse flotation and improve the quality of iron concentrate.This paper describes the development of Complex Flashing-Field Magnetic Cleaner,a new kind of equipment based on the principle of combination of separation in magnetic field and separation in gravity field(Hao & Jiang, 2002).This innovative equipment possesses a tailor-made magnetic system and has the characteristics of a magneticfieldwhich can strengthen the formation of magnetic agglomerations and the dispersion of gangue mineralsto improvethe quality of separation.The automatic control system developed specifically for this equipment can further increase the accuracy of separation and stabilize the process.

Jan 1, 2014

By L. J. Li, Q. Zhang, Z. H. Liu, Q. L. Chen

Phosphogypsum (PG) is an industrial waste from the production of phosphate fertilizer. The utilization of PG is restricted by its containing impurities, and the pretreatment is crucial to eliminating these impurities. The purpose of the research is to recycle pretreatment PG as raw material for preparation of a-calcium sulfate hemihydrate (a-HH) by salt solution method. The effect of pretreatment technologies of lime neutralization and washing on dehydration rate, phase composition and morphology evolution of a-HH were investigated by crystal water content analysis, XRD and SEM. Furthermore, the change of pH in the reaction process was also discussed in detail. The results show that the dehydration rate of conversion PG to a-HH increases and the corresponding induction period, crystal growth period and total time decrease with an increase in CaO dosage from 0.2% to 0.6%. In the presence of 0.4% CaO or 0.6% CaO, the a-HH presents a long hexagon-prism-shaped crystal with a purity of 94%, and the end face of a-HH presents polyhedron. The system pH sharply decreases with the reaction time under different CaO dosage, and the pH variation from the initial alkalinity to acidity. The main reason is that as the co-crystallized phosphorus dissolved and released HPO42-, and with the pH decreasing to 4-5, HPO42- transformed into H2PO4-. The dehydration rate of washing PG is lower than that of original PG. Compared with the original PG, a-HH with the larger width and smaller aspect ratio can be achieved from the pretreatment PG subjected to lime neutralization and washing, meanwhile the crystal surface is more smooth.

Jan 1, 2018

By Reza M. Rahman

The froth phase plays an important role in the flotation and upgrading of the final concentrate. At present, the techniques that have been used in the mineral industry to measure the froth recovery have limitations. The aim of this paper is to investigate the performance of the froth in an industrial flotation cell. A unique device has been developed which is capable of decoupling the froth zone from the pulp zone. It consists of two concentric tubes. The inner tube acts as a dropback collection chamber or catcher. The particles that detach from the froth phase fall directly into the catcher and are collected as froth dropback. The froth recovery measurements were carried out in a rougher bank of a copper concentrator treating sulphide minerals. The dropback device is designed so that it can be immersed into a plant size flotation cell and plant froth recovery measurements can be taken at any given time. During the experiments, the bubbles laden with valuable mineral particles entered the device from the plant flotation cell, subsequently rising to form a froth layer at the top of the device. The particles that detached from the froth zone were collected in the dropback collection chamber whereas the concentrate sample was collected by a launder. By drying, sizing and chemical analysis of the concentrate and dropback samples, the froth recovery, was estimated on the basis of the valuable component. The effect of air rate on froth recovery was also investigated. Keywords: flotation, froth, dropback, froth recovery

Sep 1, 2012

By Ljudmil S. Grigorov, Eric Forssberg, Hanumantha Rao Kota, Robert J. Pugh, Lidia B. Alexandrova

"The effects of mixed collectors on some physicochemical parameters (thin film lifetime, kinetics of expansion of the three-phase contact, equilibrium three-phase contact angle and surface tension) involved in the elementary act of flotation were investigated experimentally.We have used the microscopic thin balance developed by Scheludko et al (1970) to study the stability and interactions within the thin film between the silica surface and air bubble with (i) a long chain C10 amine and (ii) a cationic/anionic surfactant systems consisting of mixtures of a long straight C10 chain amine with long straight chain C8, C10 and C12 sulfonates at 1:1 molar ratio. The experiments were performed at a range of concentrations at pH 6.5. The equilibrium contact angle was measured with a microscope technique using the modified Mingins-Sheludko method (Mingins and Scheludko, 1979). The surface tension was measured by the method of Wilhelmy (Adamson, 1960). It was found that the mixed surfactants caused lower contact angles, lower rates of the thin film rupture and longer film lifetime, compared to the pure amine system."

Jan 1, 2003

By D. Pirouzan

It is customary to use more than one stage of flotation to obtain an acceptable level of separation of valuables. Flotation circuit design is usually accomplished by applying empirical rules which usually results in operations not working at their optimum conditions. Given the capabilities of genetic algorithms (GA) in finding an optimum solution in a very disturbed search spaces, they can be used to obtain the desired flotation circuit configuration. In flotation circuit configuration and optimization problem, a combination of metallurgical parameters such as yield and concentrate ash content could be used as the fitness function for the genetic algorithm. The multi-objective nature of the problem justified to use Pareto method to arrive at a set of solutions. The appropriate configuration based on technical or economical considerations could then be chosen. To obtain the fitness function for any given configuration, it is necessary to model every flotation stage. The proposed method was used to find the optimum circuit configuration for a coal washing plant. The objective was to arrive at the highest yield while producing a concentrate with a certain ash content (< 11%). The feed to the flotation circuit was characterized based on the size fractions and their flotation rate constants. Results showed that with a 95% confidence, the absolute difference between the modeled and measured values were 2.9-5.5% for yield and 0.4-1.1% for concentrate ash content. When the proposed GA-based circuit configuration was implemented in the plant, the yield increased from the original value of 57.6% to 65.7% while producing concentrate ash content (10.9%) within acceptable limits. By adding one stage to the current three-stage circuit, it was predicted that the yield could further increase by 3.8% while keeping the quality of the concentrate within the appropriate level. Keywords: modeling, optimizing, flotation circuit configuration, genetic algorithms

Sep 1, 2012

By S. H. Morar

Froth flotation is a physico-chemical separation process that takes place in two phases; pulp and froth phase. It is widely recognised as a complex process, due to the range of interactions present within the system. Of critical importance is the role of the froth phase within the system. The pulp kinetics governs the maximum potential recovery within a flotation cell, while the froth phase governs the selectivity via the froth?s stability. Higher froth stability improves recovery at the expense of grade, and lower froth stability improves selectivity. It is well known that the two major factors affecting froth stability are attached particles and solution chemistry. This paper looks at a flotation bank at different froth depth and frother concentrations. The air recovery to the concentrate is determined at each condition and relates the effect of the operating conditions on froth stability and transport to the amount of solids that are recovered to the froth and concentrate by true flotation. The results presented show that, for a given process objective, which in this case was a consistent mass pull profile, depending on the operating conditions, solids recovery in the pulp is affected by the air rate required to maintain the target mass pull. To a smaller extent, the changes in the pulp composition affect down-stream pulp recovery. The variation in solids recovered into the froth drives the froth stability at the start of the flotation bank, while the frother drives the froth stability at the end of the bank. The performance in between these two extreme conditions arises from an interaction of these two effects. The consequences of these interactions can give rise to counter-intuitive responses that can confound efforts to optimise the performance of the bank.

Jan 1, 2014

By Venkataramana Runkana, Aditya Pareek, Vishnu Masampally, Nagaravi Kumar Varma Nadimpalli

"Recovery and grade of the ore produced are two important metrics, which define the profitability of a mineral processing plant. Achieving high grade and high recovery at the same time in a mineral processing plant is difficult due to conflicting nature of the two objectives. To explore the operability limits of the plant operations and find the set of operating conditions that can help achieve the best possible recovery and grade, there is a need to solve a multi-objective optimization (MOO) problem. Past research focused on building a plant-wide simulation model or on optimization of individual units or a part of a mineral processing plant like grinding circuit, crushing circuit, etc. However, it is necessary to consider optimization of the complete plant in order to optimize the plant performance. Here, we have formulated the mineral processing plant performance optimization as a MOO problem. We have applied machine-learning algorithms such as random forest to build models to predict the recovery and grade of the ore. A non-dominated sorting-based multi-objective evolutionary algorithm called NSGA-II (Non-dominated Sorting Genetic Algorithm II) was employed to solve the MOO problem. Since the number of parameters that influence recovery and grade are quite large, we have applied machine-learning algorithms for feature selection or identification of key process variables. Pareto optimal conditions were determined for manipulated variables like grinding mill throughput, speed ratio at SAG mill, speed of centrifugal pump, water addition at different locations etc. The results obtained are useful in identifying the operability of a mineral processing plant to achieve the optimum grade and recovery for a given feed grade and the processing circuit. INTRODUCTION Copper is majorly mined or extracted as copper sulfides from large open pit mines in porphyry copper deposits. Among all the available ores, sulfides are the most commercial ores, especially chalcopyrite (Greenwood and Alan, 1997). Copper grade in these ores generally lies between 0.4 to 1.0%. In order to perform smelting operation, the grade of copper has to be improved to around 30% (Biswas et al. 2013). This targeted copper grade is achieved through two stages of mineral processing. Firstly, the copper in the ore has to be liberated by crushing the mineral rock. Secondly, the liberated copper in the crushed ore is then physically separated from other minerals like nickel by froth flotation techniques."

Jan 1, 2018

By B. S. Moroney

Elemental distribution varies widely in the ores, minerals and beneficiated products depending on the constituents. Accurate analysis of each element present in various beneficiated products is utmost required to assess the grade in view of zero waste technology. Hence, the suitable methods and instrumental technique is of prime importance as the distribution of various elements differs in the beneficiated products. Conventional classical technique of analysis is absolute and due to this, it has not become obsolete. Besides, it is much needed for making standards for various relative techniques. Instrumental techniques comprising of X-ray fluorescence, Induced Coupled Plasma Analyzer, Carbon-Sulphur Analyzer, Atomic Absorption Spectrometer etc. are important for quick and accurate analysis. However, the selection of these instruments depend on the nature of material to be analyzed. This paper highlights the importance and use of various methods/techniques and instruments involved in the analysis of various ores, minerals and beneficiated products. Keywords: wet classical, x-ray fluorescence, induced coupled plasma analyzer, atomic absorption spectrometer

Sep 1, 2012

By S. Banisi, A. R. Ghasemi, M. Maleki-Moghaddam

Direct observation of the charge shape and its motion in industrial mills are not possible, it is then customary to use transparent end, small-scale mills (i.e., model mills) to determine the charge trajectory directly by visualization methods. Because of a short length of the model mills, the end-wall effects could introduce a significant bias on the observed charge trajectories and shape. The shorter the mill, the stronger the bias. This could be due to some material bridging between the end walls. In this research, the end-wall effects were investigated by gradually increasing the model mill length and analysing the charge trajectory and shape variation at a given operating conditions (i.e., ball filling, mill speed, liner type). The unbiased mill length was assumed to be beyond the mill length where no significant change observed in the charge trajectory and shape. For all tests, the torque of the charge was also measured by a torque sensor. The special design of the model mill with the diameter of 100 cm made possible to increase the mill length in steps of 3.6 cm up to 21.6 cm. Four types of liners, five steel ball fillings (10, 15, 20, 25, 30% v/v) and three mill speeds (55, 70, 85% of critical speed) were tested. The ball sizes ranged from 4 to 12 mm. The results indicated that when the mill length was below 10.8 cm the end walls prevented the charge from free falling. This resulted in lower impact point angles and lower power draws compared with the case of no end-wall effects. When the same experiments were performed using an iron ore instead of balls with the same size range, the impact of end-wall effects was more pronounced. For an increase of 30% in the speed (from 55 to 85%), the torque due the change in the charge shape decreased by 41% (from 3.7 to 2.2 kgf.m) when mill filling was 30%. At the same operating conditions when steel balls used the torque decreased only by 19% (from 9 to 7.2 kgf.m). This was expected due to irregular shapes of the ore particles which facilitated bridging. Three-dimensional DEM simulations also resulted in the similar conclusions.

Jan 1, 2014

By Arkady Senchenko, Sergey Rybkin, Svetlana Lagutina, Alexander Aksenov

"The application of autoclave leaching technology for sulfide concentrates containing base and precious metals is spreading nowadays. In the process of concentrate leaching silver accumulates in the forms of jarosite compounds, transforms into passive chemical state and is mostly found in solid phase. Effective technology development for silver recovery from solid phase of concentrate autoclave leaching raises the necessity for comprehensive research of silver forms of occurrence and their properties in this product. The possible form of silver presence in solid phase of concentrate from autoclave leaching is argentojarosite. This article is about the study of physical and chemical properties and structure of synthetic argentojarosite - AgFe3(SO4)2(OH)6. A method of argentojarosite synthesis has been developed under conditions simulating the process of autoclave leaching of sulfide concentrate, and includes silver dissolution in the form of Ag2SO4, Fe3+ sulfate hydrolysis, AgFe3(SO4)2(OH)6 build-up. X-ray data of synthesized argentojarosite completely correspond to those of natural mineral. The structural formula of this complex compound has been studied and confirmed. The chemical properties of synthetic argentojarosite, its solubility in acids and alkalis have been investigated. Phase changes when heating argentojarosite with the build-up of iron oxide and silver sulfate stop at the temperature equal to 650 ºC. Based on the results received the technological solutions for solid phase of autoclave leaching of concentrate have been developed with silver recovery of 95-98%."

Jan 1, 2014

By C. Hazotte, M. -O. Simonnot, V. Houzelot, B. Laubie, M. Guilpain, B. Jally

"More than 400 plants in the world are able to accumulate nickel (Ni) in their aerial parts at concentrations higher than 10 g per kg of dry biomass. Then, Ni can be recovered from the biomass of these hyperaccumulator plants. Agromining is a chain consisting in growing such plants and producing metal. It has been developed thanks to the combined know-how of agronomy and hydrometallurgy. Hydrometallurgical processes have been designed; two different pathways were developed: i) acid leaching of ashes after biomass combustion and ii) direct Ni water extraction after biomass grinding. Experiments with 2 species of Ni hyperaccumulators (A. murale from the Balkans and R. bengalensis from Malaysia) have shown that the first process is very robust and transposable. For the second one, the way Ni is stored in the plant has a strong influence on extraction. A. murale transformation was explored in more details from the lab scale to the pilot scale. Results proved that Ni is totally transferred into the aqueous phase after 2 hours at 90 °C. A life cycle assessment has clearly shown the environmental interest of this approach (especially by detoxifying agricultural soils). INTRODUCTION Agromining is an emerging phytotechnology aiming at recovering metals thanks to hyperaccumulating plants (van der Ent et al., 2018). These plants are able to accumulate very high concentrations of metal in their aerial parts. For example, for nickel, the hyperaccumulation threshold is 0.1%, i.e. 10 grams of metal per kilogram of dry biomass. There are some advantages to use this technology instead of conventional mining techniques: •,Treating unconventional metal resources: they can be (i) natural soils with low concentrations but also low fertility because of the metal presence (like serpentine soils for Ni) (Bani et al., 2015), (ii) mine tailings or (iii) industrial wastes (like surface treatment sludge) after soil construction (Rue, 2017);"

Jan 1, 2018

By Raquel Franco de Souza Lima, Isao Matsuoka

"Microflotation studies and abstraction experiments were conducted with synthetic oxide spinels relating to chromite (MgCr2O4, FeCr2O4, FeFe2O4 and FeAl2O4), using sodium oleate as collector. An attempt was done to correlate this data with the speciation of oleate in the bulk solution and at the interface solid-water, to explain the floatability curves. The microflotation tests were conducted using a microflotation cell. The abstraction of sodium oleate was measured at 25°C, with one gram of synthetic spinel and the abstraction time was 30 minutes. After centrifuging, the residual oleate concentration in the supernatant was measured by using the spectrophotometric method described by Mizuniwa et al. (1977), that covers the range from 1.5 to 9 ppm. To cover the range from 0.2 to 1.5 ppm, the residual oleate concentration was measured using the spectrophotometric method described by Gregory (1966). The amount of oleate abstracted was calculated from the difference between initial and residual (equilibrium) concentrations.Figure 1 shows the floatability of FeCr2O4, FeAl2O4, FeFe2O4, MgCr2O4 as a function of pH and sodium oleate concentration of 1 x 10-5 M. The IEP of each sample is indicated in the figure by a vertical line in the horizontal axis. The chromium-bearing spinels FeCr2O4 and MgCr2O4 show the lowest isoelectric points (IEP), around pH 6, while in the case of FeFe2O4 and FeAl2O4, the isoelectric points lie at pH 6.5 and 7.3, respectively. It is seen that the flotation-pH curve for FeFe2O4 has a peak at the pH of the IEP, while flotation peak for FeCr2O4 appears at pH below the IEP and that of MgCr2O4 does at pH above the PZC. On the other hand, FeAl2O4 is floated at pH range between about 4 and 7,3 (IEP). This suggests that the pH value at which the maximum flotation occurs is dependent on the sort of spinel to be floated, though it appears in the vicinity of the respective IEP."

Jan 1, 2003