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The recovery was investigated of rare earth elements from a eudialyte concentrate containing 52wt% SiO2, 15.2 wt% Fe2O3, 11.2 wt% Na2O, 8.6 wt% Al2O3, 2.9 wt% ZrO2, 2.1 wt% CaO and 0.94 wt% total rare earth elements by hydrochloric acid leaching and precipitation as carbonate using Na2CO3. Laboratory studies were initially conducted on eudialyte concentrate from Norra Karr, Sweden. The experimental conditions were fuming in 30% hydrochloric for 60 minutes at 20 degrees C, leaching for 60 minutes at 45 degrees C, neutralisation with CaCO3 to pH 4.1 and precipitation with Na2CO3 at pH 5.19-6.50, with the addition of small amounts of flocculants to accelerate the separation process. The concentrate produced contained 15.1% of rare earth elements. Based on the results, large-scale experiments were successfully carried out. The fuming process should start with the addition of concentrate first, followed by the acid at an appropriate ratio to avoid silica gel formation. During neutralisation with CaCO3, more than 95% of the Zr, Al and Si are transferred to the residue which can be used as a potential source for Zr recovery. More than 95% of the rare earth elements were precipitated during treatment with Na2CO3., The recovery was investigated of rare earth elements from a eudialyte concentrate containing 52wt% SiO2, 15.2 wt% Fe2O3, 11.2 wt% Na2O, 8.6 wt% Al2O3, 2.9 wt% ZrO2, 2.1 wt% CaO and 0.94 wt% total rare earth elements by hydrochloric acid leaching and precipitation as carbonate using Na2CO3. Laboratory studies were initially conducted on eudialyte concentrate from Norra Karr, Sweden. The experimental conditions were fuming in 30% hydrochloric for 60 minutes at 20 degrees C, leaching for 60 minutes at 45 degrees C, neutralisation with CaCO3 to pH 4.1 and precipitation with Na2CO3 at pH 5.19-6.50, with the addition of small amounts of flocculants to accelerate the separation process. The concentrate produced contained 15.1% of rare earth elements. Based on the results, large-scale experiments were successfully carried out. The fuming process should start with the addition of concentrate first, followed by the acid at an appropriate ratio to avoid silica gel formation. During neutralisation with CaCO3, more than 95% of the Zr, Al and Si are transferred to the residue which can be used as a potential source for Zr recovery. More than 95% of the rare earth elements were precipitated during treatment with Na2CO3.

The use is discussed of computational fluid dynamics (CFD) for the design of reactors which employ injection of gases to drive the reaction, such as top submerged lance furnaces, QSL furnaces, submerged plasma furnaces and Peirce-Smith converters. CFD models of reactive gas injection contain three distinct submodules, a multiphase model, a turbulence model and a reaction model, and the interaction between the submodules is key to successful simulation. Multiphase models for gas injection inside an incompressible liquid include the volume of fluid (VOF), Eulerian and discrete particle (DPM) models. Two models are used for the turbulence module, the k-epsilon model based on the Reynolds averaged Navier-Stokes equations, where k is the turbulent kinetic energy and epsilon the turbulent dissipation rate, and large eddy simulations. Reaction modelling is subdivided into two approaches for reactive gas injection based on defining the reaction rate or by equilibrium calculations using a thermodynamic database. The complexity and limitations of CFD modelling are illustrated using two examples, gas injection in a Hismelt reactor and Zn fuming in a classical box furnace and in a top submerged lane furnace. A literature review is presented which shows that the most popular and successful model is a combination of the Eulerian multiphase description, the k-epsilon turbulence model and the definition of the reaction rate for a limited number of reactions. Although resulting in complex models able to capture the complex dynamics between reactions and flows, the approach has drawbacks. The Eulerian multiphase description exhibits difficult convergence for more than two phases, the k-epsilon turbulence model only describes the mean contribution of the turbulence and defining the reaction rates requires extensive a priori knowledge of the flows and processes., The use is discussed of computational fluid dynamics (CFD) for the design of reactors which employ injection of gases to drive the reaction, such as top submerged lance furnaces, QSL furnaces, submerged plasma furnaces and Peirce-Smith converters. CFD models of reactive gas injection contain three distinct submodules, a multiphase model, a turbulence model and a reaction model, and the interaction between the submodules is key to successful simulation. Multiphase models for gas injection inside an incompressible liquid include the volume of fluid (VOF), Eulerian and discrete particle (DPM) models. Two models are used for the turbulence module, the k-epsilon model based on the Reynolds averaged Navier-Stokes equations, where k is the turbulent kinetic energy and epsilon the turbulent dissipation rate, and large eddy simulations. Reaction modelling is subdivided into two approaches for reactive gas injection based on defining the reaction rate or by equilibrium calculations using a thermodynamic database. The complexity and limitations of CFD modelling are illustrated using two examples, gas injection in a Hismelt reactor and Zn fuming in a classical box furnace and in a top submerged lane furnace. A literature review is presented which shows that the most popular and successful model is a combination of the Eulerian multiphase description, the k-epsilon turbulence model and the definition of the reaction rate for a limited number of reactions. Although resulting in complex models able to capture the complex dynamics between reactions and flows, the approach has drawbacks. The Eulerian multiphase description exhibits difficult convergence for more than two phases, the k-epsilon turbulence model only describes the mean contribution of the turbulence and defining the reaction rates requires extensive a priori knowledge of the flows and processes.

Metallurgical processes generate daily production data which can be the largest source of information about the process, but many metallurgical plants have data and process predictability issues such as the data being incomplete, containing errors or being stored in different places or offline. Software solutions to handle process data in combination with predictive flowsheet models are considered with reference to ProOpt International's technology for data collection, storage, and statistical analysis. It is important to be able to assess the expected yields, energy needs and product qualities even before material has arrived at a site and without test work. Thermodynamics can be used to predict what is and what is not possible, and integrated flowsheet modelling tools allow estimation of realistic values for all up- and downstream processes, which in turn determines the subsequent efficiencies and potential bottlenecks. Metal production and recycling methods often require an understanding of material behaviour at high temperature, including how much energy is needed to heat the material or will it bring to the process, the properties of the resulting slag and the metal yield. Framing one or more thermodynamic models in a flowsheet model allows the central process to be linked to all ancillary equipment and activities such as storage, filter systems and water cooling. An example is presented for the off-gas section of a smelter. A combined model and data management approach guarantees that conclusions drawn from data analysis can be implemented in the model so that a continuous improvement loop of model and process can be achieved. A good tool will collect, validate, correct or reject and analyse data, and also deliver the conclusions to the relevant stakeholders. Examples are described including the prediction of the slag melting point and tap temperature in furnaces to reduce fluctuations, optimise procedures and targets and save energy and wear by working at low, Metallurgical processes generate daily production data which can be the largest source of information about the process, but many metallurgical plants have data and process predictability issues such as the data being incomplete, containing errors or being stored in different places or offline. Software solutions to handle process data in combination with predictive flowsheet models are considered with reference to ProOpt International's technology for data collection, storage, and statistical analysis. It is important to be able to assess the expected yields, energy needs and product qualities even before material has arrived at a site and without test work. Thermodynamics can be used to predict what is and what is not possible, and integrated flowsheet modelling tools allow estimation of realistic values for all up- and downstream processes, which in turn determines the subsequent efficiencies and potential bottlenecks. Metal production and recycling methods often require an understanding of material behaviour at high temperature, including how much energy is needed to heat the material or will it bring to the process, the properties of the resulting slag and the metal yield. Framing one or more thermodynamic models in a flowsheet model allows the central process to be linked to all ancillary equipment and activities such as storage, filter systems and water cooling. An example is presented for the off-gas section of a smelter. A combined model and data management approach guarantees that conclusions drawn from data analysis can be implemented in the model so that a continuous improvement loop of model and process can be achieved. A good tool will collect, validate, correct or reject and analyse data, and also deliver the conclusions to the relevant stakeholders. Examples are described including the prediction of the slag melting point and tap temperature in furnaces to reduce fluctuations, optimise procedures and targets and save energy and wear by working at low

Research was carried out into the possibility of reducing the odour nuisance of off-gases. Chromatographic analysis of gas composition and content of odour-generating organic substances detected 320 organic compounds including aldehydes, thiophenes, carboxylic acids and N and S compounds. The gases were then oxidised by contact with oxidising solutions of H2O2, O3 and ClO2 in a scrubber or a sprayed column to neutralise odour, under different pH conditions and oxidant dosages. The efficiency of gas deodourising was determined by analysing the chemical composition of off-gas after oxidation. In addition, the degree of deodourisation was evaluated using the organoleptic odour intensity scale of 1 to 5. The study showed that ozone and chlorine dioxide are the most effective in limiting odour nuisance of gases from pyrometallurgical secondary recovery processes., Research was carried out into the possibility of reducing the odour nuisance of off-gases. Chromatographic analysis of gas composition and content of odour-generating organic substances detected 320 organic compounds including aldehydes, thiophenes, carboxylic acids and N and S compounds. The gases were then oxidised by contact with oxidising solutions of H2O2, O3 and ClO2 in a scrubber or a sprayed column to neutralise odour, under different pH conditions and oxidant dosages. The efficiency of gas deodourising was determined by analysing the chemical composition of off-gas after oxidation. In addition, the degree of deodourisation was evaluated using the organoleptic odour intensity scale of 1 to 5. The study showed that ozone and chlorine dioxide are the most effective in limiting odour nuisance of gases from pyrometallurgical secondary recovery processes.

In order to determine the associated ecological impact and sustainability, new transformative technologies need to be evaluated from the environmental footprint point of view prior commercialisation. In this study, the environmental indicator category chosen for evaluation was global warming potential and the other impact categories, such as human toxicity potential and acidification, were left out of the scope. The global warming potentials (GWP) were determined both for the historical chlorination process and for a state-of-art cyanidation process by modelling with HSC Chemistry 8.0 HSC-Sim module combined with life-cycle analysis (LCA) by GaBi 6.0. This analysis provides a baseline scenario for comparison, which can be used to support the future life-cycle assessment research for development-stage gold processes. The combination of HSC-Sim and GaBi was shown to be an efficient method for investigating the environmental footprint of the historical and current gold processes. The GWP of the cyanidation process was 455 kg CO2 equivalent and the GWP of chlorination was 10 500 kg CO2 equivalent. The main emission source in cyanidation was found to be electricity whereas in chlorination the main emissions originated from the chlorine gas, although it is worth noting that environmental impacts are not defined for all chemical products in the GaBi database. Additionally, some compromises had to be made when the LCA equivalents were applied to the process streams modelled in HSC-Sim. The corresponding or most appropriate LCA equivalents were defined from the HSC/GaBi database and the LCA group was defined for all the inputs and outputs within the scope of the process., In order to determine the associated ecological impact and sustainability, new transformative technologies need to be evaluated from the environmental footprint point of view prior commercialisation. In this study, the environmental indicator category chosen for evaluation was global warming potential and the other impact categories, such as human toxicity potential and acidification, were left out of the scope. The global warming potentials (GWP) were determined both for the historical chlorination process and for a state-of-art cyanidation process by modelling with HSC Chemistry 8.0 HSC-Sim module combined with life-cycle analysis (LCA) by GaBi 6.0. This analysis provides a baseline scenario for comparison, which can be used to support the future life-cycle assessment research for development-stage gold processes. The combination of HSC-Sim and GaBi was shown to be an efficient method for investigating the environmental footprint of the historical and current gold processes. The GWP of the cyanidation process was 455 kg CO2 equivalent and the GWP of chlorination was 10 500 kg CO2 equivalent. The main emission source in cyanidation was found to be electricity whereas in chlorination the main emissions originated from the chlorine gas, although it is worth noting that environmental impacts are not defined for all chemical products in the GaBi database. Additionally, some compromises had to be made when the LCA equivalents were applied to the process streams modelled in HSC-Sim. The corresponding or most appropriate LCA equivalents were defined from the HSC/GaBi database and the LCA group was defined for all the inputs and outputs within the scope of the process.

The dismantling and detachment of components from electronic scrap during pre-treatment is essential but involve higher costs as they are usually carried out manually. Automatic dismantling machines have recently been developed but they are limited in relation to their application to a wide variety of scrap materials. Typical crushers and grinding mills for the comminution of waste electrical and electronic scrap and printed circuit boards (PCBs) include jaw crushers for rough crushing and ball mills to achieve a finer particle size. A new process is described in which the scrap is disintegrated by high-voltage electric pulsing in water. The next pre-treatment step is the separation of each component and/or material using magnetic, gravity and eddy current separation to separate iron scrap, non-ferrous scrap and plastics, respectively. Sorting techniques are used for separation based on properties such as colour, shape, mass and chemical composition. Sensor techniques include infrared, X-ray diffraction and X-ray fluorescence, and a recent development has been the use of a laser-induced breakdown spectroscopy sorter for PCBs consisting of copper foil and epoxy resin with flame retardants. Pyrolysis is required before the recovery of precious metals from PCBs as the plastics may cause problems in copper smelting furnaces and in the direct refining of precious metals as a result of highly corrosive flue gases and the loss of valuable metals in dust. Acidic and alkaline solutions are used in the hydrometallurgical processing of PCBs. Fine grinding of PCB powders is required, and strongly oxidising solutions consisting of nitric or sulphuric acids with hydrogen peroxide as oxidant are used as lixiviants. Hydrometallurgical processes are normally applied to high-grade PCBs containing gold-plated electronic junctions, and cyanide solution is used for the selective leaching of gold and other precious metals. Thermal treatment of PCBs sn conducted in incinerators such as ro, The dismantling and detachment of components from electronic scrap during pre-treatment is essential but involve higher costs as they are usually carried out manually. Automatic dismantling machines have recently been developed but they are limited in relation to their application to a wide variety of scrap materials. Typical crushers and grinding mills for the comminution of waste electrical and electronic scrap and printed circuit boards (PCBs) include jaw crushers for rough crushing and ball mills to achieve a finer particle size. A new process is described in which the scrap is disintegrated by high-voltage electric pulsing in water. The next pre-treatment step is the separation of each component and/or material using magnetic, gravity and eddy current separation to separate iron scrap, non-ferrous scrap and plastics, respectively. Sorting techniques are used for separation based on properties such as colour, shape, mass and chemical composition. Sensor techniques include infrared, X-ray diffraction and X-ray fluorescence, and a recent development has been the use of a laser-induced breakdown spectroscopy sorter for PCBs consisting of copper foil and epoxy resin with flame retardants. Pyrolysis is required before the recovery of precious metals from PCBs as the plastics may cause problems in copper smelting furnaces and in the direct refining of precious metals as a result of highly corrosive flue gases and the loss of valuable metals in dust. Acidic and alkaline solutions are used in the hydrometallurgical processing of PCBs. Fine grinding of PCB powders is required, and strongly oxidising solutions consisting of nitric or sulphuric acids with hydrogen peroxide as oxidant are used as lixiviants. Hydrometallurgical processes are normally applied to high-grade PCBs containing gold-plated electronic junctions, and cyanide solution is used for the selective leaching of gold and other precious metals. Thermal treatment of PCBs sn conducted in incinerators such as ro

Demand for tellurium is projected to continue to increase as cadmium telluride materials are an essential component in photovoltaic cell technology. Currently the majority of supply is produced through the processing of copper anode slimes, but this is expected to decline with decreasing global copper ore quality and exploiting alternate sources of the metal is of great importance. Several gold deposits worldwide contain significant proportions of refractory telluride minerals and two regions, Kankberg, Sweden and Dashuigou and Majiagou, China, reportedly produce tellurium as a by-product of gold operations. The four primary methods for treating refractory gold are roasting, chemical oxidation, biological oxidation and ultra-fine grinding. The basic behaviour of tellurium and hessite, as a representative telluride mineral, was investigated in various pre-treatment methods and the potential for subsequent recovery was evaluated. The telluride species were synthesised from elemental powders of tellurium and silver of 99 % purity. The hessite was prepared by mixing stoichiometric amounts of tellurium and silver powder in a Pyrex tube which was then placed in a furnace at 500 degrees C for 96 h. Roasting simulation was conducted using a muffle furnace preheated to 600-700 degrees C. Hessite samples were leached in alkaline cyanide solution for 6 h to investigate the leach behaviour of the synthetic mineral. Chemical oxidation was achieved by immersing the hessite in a solution containing 2 % NaOH and 4.4 g/l Ca(OCl)2 for 6 h. The resulting solids were ground and subsequently leached in alkaline cyanide to determine whether a greater proportion of silver was cyanide recoverable than in the alloy. Three samples were provided from an operation in Western Australia, two from roaster campaigns; a calcine sample and calcine leach residue which were leached with a dissolved oxygen content of about 2 ppm, 400 ppm NaCN and a pH of 9.5, and the third a leach residue which was gro, Demand for tellurium is projected to continue to increase as cadmium telluride materials are an essential component in photovoltaic cell technology. Currently the majority of supply is produced through the processing of copper anode slimes, but this is expected to decline with decreasing global copper ore quality and exploiting alternate sources of the metal is of great importance. Several gold deposits worldwide contain significant proportions of refractory telluride minerals and two regions, Kankberg, Sweden and Dashuigou and Majiagou, China, reportedly produce tellurium as a by-product of gold operations. The four primary methods for treating refractory gold are roasting, chemical oxidation, biological oxidation and ultra-fine grinding. The basic behaviour of tellurium and hessite, as a representative telluride mineral, was investigated in various pre-treatment methods and the potential for subsequent recovery was evaluated. The telluride species were synthesised from elemental powders of tellurium and silver of 99 % purity. The hessite was prepared by mixing stoichiometric amounts of tellurium and silver powder in a Pyrex tube which was then placed in a furnace at 500 degrees C for 96 h. Roasting simulation was conducted using a muffle furnace preheated to 600-700 degrees C. Hessite samples were leached in alkaline cyanide solution for 6 h to investigate the leach behaviour of the synthetic mineral. Chemical oxidation was achieved by immersing the hessite in a solution containing 2 % NaOH and 4.4 g/l Ca(OCl)2 for 6 h. The resulting solids were ground and subsequently leached in alkaline cyanide to determine whether a greater proportion of silver was cyanide recoverable than in the alloy. Three samples were provided from an operation in Western Australia, two from roaster campaigns; a calcine sample and calcine leach residue which were leached with a dissolved oxygen content of about 2 ppm, 400 ppm NaCN and a pH of 9.5, and the third a leach residue which was gro

Thermodynamic simulations using the Thermocalc software were used to investigate the possible chlorination of pure Ta2O5 with C2Cl4 in comparison with the usual method using Cl2 and graphite. The results showed that C2Cl4 can replace Cl2 in the temperature range 873-1 173 K, resulting in the formation of gaseous TaCl5 and TaOCl3 as the major chlorinated products under continuous flow. The presence of oxygen in the reactor can be important for oxidising non-reacted C2Cl4 to Cl2 and CO or CO2, particularly when the atmosphere becomes progressively concentrated in C2Cl4. Experimental results using high-purity Ta2O5 samples at 1 173 K showed that C2Cl4 diluted in N2 at a C2Cl4:N2 ratio of 0.03 resulted in the preferential formation of TaCl5, in agreement with the simulations. The chlorides formed were absorbed in KOH solutions, and fine solid particles were identified in the solutions, attributed to the decomposition of a portion of the injected C2Cl4 as it travelled between the reactor entrance and the sample., Thermodynamic simulations using the Thermocalc software were used to investigate the possible chlorination of pure Ta2O5 with C2Cl4 in comparison with the usual method using Cl2 and graphite. The results showed that C2Cl4 can replace Cl2 in the temperature range 873-1 173 K, resulting in the formation of gaseous TaCl5 and TaOCl3 as the major chlorinated products under continuous flow. The presence of oxygen in the reactor can be important for oxidising non-reacted C2Cl4 to Cl2 and CO or CO2, particularly when the atmosphere becomes progressively concentrated in C2Cl4. Experimental results using high-purity Ta2O5 samples at 1 173 K showed that C2Cl4 diluted in N2 at a C2Cl4:N2 ratio of 0.03 resulted in the preferential formation of TaCl5, in agreement with the simulations. The chlorides formed were absorbed in KOH solutions, and fine solid particles were identified in the solutions, attributed to the decomposition of a portion of the injected C2Cl4 as it travelled between the reactor entrance and the sample.

The recovery of yttrium was investigated from ceramic lining material used in TiAl casting moulds and containing more than 90% Al2O and an yttrium oxide layer. The samples were crushed, ground and sieved to obtain an yttrium-rich target fraction with a grain size of less than 250 micrometres. The samples contained 18.1 wt% Y2O3, 7.78 wt% SiO2, 0.74 wt% TiO2 and 0.18 wt% Fe2O3 in an Al2O3 matrix. The effects of acid concentration, agitation time, solid/liquid ratio and fraction size on hydrochloric acid leaching efficiency were investigated. Increasing the hydrochloric acid concentration from 1 to 2 mol/l improved the leaching efficiency. The maximum yttrium leaching efficiency of 65% was obtained by leaching a 100 g sample with 1 mol/l HCl solution at 70 degrees C for 2 h. Scale-up of the process using 100 l reactors under the same conditions resulted in a leaching efficiency of 40%. Increasing the leaching time from 2 to 6 h increased the leaching efficiency to up to 90%, indicating that leaching was not complete after 6 h and that an extended leaching time of up to 24 h would have a positive effect on dissolution yield., The recovery of yttrium was investigated from ceramic lining material used in TiAl casting moulds and containing more than 90% Al2O and an yttrium oxide layer. The samples were crushed, ground and sieved to obtain an yttrium-rich target fraction with a grain size of less than 250 micrometres. The samples contained 18.1 wt% Y2O3, 7.78 wt% SiO2, 0.74 wt% TiO2 and 0.18 wt% Fe2O3 in an Al2O3 matrix. The effects of acid concentration, agitation time, solid/liquid ratio and fraction size on hydrochloric acid leaching efficiency were investigated. Increasing the hydrochloric acid concentration from 1 to 2 mol/l improved the leaching efficiency. The maximum yttrium leaching efficiency of 65% was obtained by leaching a 100 g sample with 1 mol/l HCl solution at 70 degrees C for 2 h. Scale-up of the process using 100 l reactors under the same conditions resulted in a leaching efficiency of 40%. Increasing the leaching time from 2 to 6 h increased the leaching efficiency to up to 90%, indicating that leaching was not complete after 6 h and that an extended leaching time of up to 24 h would have a positive effect on dissolution yield.

KCM 2000 Group operates a non-ferrous smelter near Plovdiv, Bulgaria with annual production of 70 000 tpa of Pb and 75 000 tpa of Zn, including their alloys and compounds. A new brown-field Pb smelting plant was commissioned in 2014 designed to produce 75 000 tpa of crude Pb and 55 000 tpa of sulphuric acid. The plant replaced one based on sintering and blast furnace slags; reduction; batteries; secondary recovery; fluxes; and employs the top submerged lance (TSL) Ausmelt smelting technology and the wet sulphuric acid process. The process is organised in cycles with a duration of around 9 h each. Batches are smelted for about 7 h, with subsequent 1.5 h of slag reduction with lump coal, both stages taking place in a single vessel. The slag is then tapped, granulated and stockpiled for Zn recovery. Air, oxygen and natural gas are injected through the lance submerged in the bath. The feed material mixture is prepared in a new preparation plant equipped with day bins for each type of feed, weight belt feeders controlled by a furnace control system and a mixing drum for adjusting moisture and particle size. The raw materials consist mainly of Pb concentrates and secondary materials such as Pb paste from KCM's spent Pb-acid batteries plant, Pb cakes from the Zn plant and recycled dust in the form of cake generated by the TSL furnace. Fluxes added for slag chemistry control include roasted iron pyrites, silica and lime ash. The furnace off-gas is cooled to 650 degrees C in a waste heat boiler. The steam is utilised for plant needs and the surplus is condensed and cleaned for reuse. The SO2 content of the off-gas is about 8 vol% during the smelting stage and close to 0 vol% during the reduction stage. Dust generation is around 5 000 kg/h for both stages. The gas is conditioned in a wet gas cleaning section before entering the sulphuric acid plant. Cooling and dust extraction occur in a series of units, quench tower, radial flow scrubber, packed cooling tower and two wet ele, KCM 2000 Group operates a non-ferrous smelter near Plovdiv, Bulgaria with annual production of 70 000 tpa of Pb and 75 000 tpa of Zn, including their alloys and compounds. A new brown-field Pb smelting plant was commissioned in 2014 designed to produce 75 000 tpa of crude Pb and 55 000 tpa of sulphuric acid. The plant replaced one based on sintering and blast furnace slags; reduction; batteries; secondary recovery; fluxes; and employs the top submerged lance (TSL) Ausmelt smelting technology and the wet sulphuric acid process. The process is organised in cycles with a duration of around 9 h each. Batches are smelted for about 7 h, with subsequent 1.5 h of slag reduction with lump coal, both stages taking place in a single vessel. The slag is then tapped, granulated and stockpiled for Zn recovery. Air, oxygen and natural gas are injected through the lance submerged in the bath. The feed material mixture is prepared in a new preparation plant equipped with day bins for each type of feed, weight belt feeders controlled by a furnace control system and a mixing drum for adjusting moisture and particle size. The raw materials consist mainly of Pb concentrates and secondary materials such as Pb paste from KCM's spent Pb-acid batteries plant, Pb cakes from the Zn plant and recycled dust in the form of cake generated by the TSL furnace. Fluxes added for slag chemistry control include roasted iron pyrites, silica and lime ash. The furnace off-gas is cooled to 650 degrees C in a waste heat boiler. The steam is utilised for plant needs and the surplus is condensed and cleaned for reuse. The SO2 content of the off-gas is about 8 vol% during the smelting stage and close to 0 vol% during the reduction stage. Dust generation is around 5 000 kg/h for both stages. The gas is conditioned in a wet gas cleaning section before entering the sulphuric acid plant. Cooling and dust extraction occur in a series of units, quench tower, radial flow scrubber, packed cooling tower and two wet ele

The FAST Pb electrowinning process was developed by Engitec to convert Pb in Pb-acid battery paste into metal. A demonstration plant was constructed to improve the efficiency of the process, resulting in modifications to the operating parameters and plant equipment to make the process more robust. The feed for electrolysis is the super-desulphurised paste obtained from the latest development of the CX desulphurisation process unit, which is based on a double stage counter-current design using Na2CO3 as desulphurisation reagent. The desulphurised paste is contacted with CO2 to remove Na in the form of NaHCO3 which can be recycled for use as a desulphurisation agent. Advantages of the smelting of the super-desulphurised paste include practically zero SO2 emissions and minimisation of the chemicals required to fix sulphur in the slag. In the FAST Pb process, Pb is deposited from a chloride-based electrolyte in a non-adherent flake form. Under particular conditions, the flakes are removed and transported from the cell by the electrolyte flow which strips the flakes from the cathode. The super-desulphurised paste is leached at temperatures above 80 degrees C. H2O2 is used to reduce the PbO2 which is otherwise not leachable in the Pb depleted solution. The Pb leaching efficiency is more than 98 %. The cell was operated at very high current densities of up to 40 000 A/m2, but for industrial application the upper limit is about 12 000 A/m2 to achieve a reasonable energy cost. The cell voltage is in the range 2.9-3.1 V for current densities up to 10 000 A/m2. The current efficiency does not change very much with the current density and is in the range 90%-92 %. The flakes produced were briquetted at 800 bar to produce briquettes which were fed to a crucible to be converted into a cast ingot., The FAST Pb electrowinning process was developed by Engitec to convert Pb in Pb-acid battery paste into metal. A demonstration plant was constructed to improve the efficiency of the process, resulting in modifications to the operating parameters and plant equipment to make the process more robust. The feed for electrolysis is the super-desulphurised paste obtained from the latest development of the CX desulphurisation process unit, which is based on a double stage counter-current design using Na2CO3 as desulphurisation reagent. The desulphurised paste is contacted with CO2 to remove Na in the form of NaHCO3 which can be recycled for use as a desulphurisation agent. Advantages of the smelting of the super-desulphurised paste include practically zero SO2 emissions and minimisation of the chemicals required to fix sulphur in the slag. In the FAST Pb process, Pb is deposited from a chloride-based electrolyte in a non-adherent flake form. Under particular conditions, the flakes are removed and transported from the cell by the electrolyte flow which strips the flakes from the cathode. The super-desulphurised paste is leached at temperatures above 80 degrees C. H2O2 is used to reduce the PbO2 which is otherwise not leachable in the Pb depleted solution. The Pb leaching efficiency is more than 98 %. The cell was operated at very high current densities of up to 40 000 A/m2, but for industrial application the upper limit is about 12 000 A/m2 to achieve a reasonable energy cost. The cell voltage is in the range 2.9-3.1 V for current densities up to 10 000 A/m2. The current efficiency does not change very much with the current density and is in the range 90%-92 %. The flakes produced were briquetted at 800 bar to produce briquettes which were fed to a crucible to be converted into a cast ingot.

A pilot study was conducted into the recovery of Co from the hot filter cake produced at the Zanjan Zn plant in Iran, where a leach-electrolysis process is employed for Zn production. The filter cake sample contained 20.44 %Zn, 0.32% Co, 2.65% Mn, 0.001% Ni, 0.02% Cd, 0.78% Fe, 8.57% Ca and 0.15% Pb. The process involved Zn washing with sulphuric acid at pH 1 at a temperature of 55 degrees C for 6 h, with Zn recovery and Co loss of 91.3% and 7.04%, respectively, reductive-acidic leaching of the Zn washing residue at pH 1 and a temperature of 35 degrees C for 3 h, with average Zn, Co and Mn recoveries of 74.36%, 82.34% and 81.92 %, respectively, Zn solvent extraction using 20% D2EHPA + 5 % TBP diluted to 25 vol% in kerosene at pH 3, organic/aqueous ratio of 1,temperature of 25 degrees C for 5 h with average Zn and Mn recoveries and Co loss of 99.96%, 99.72% and 1.38%, respectively, precipitation of Co as carbonate at pH 3 and temperature of 40 degrees C for 3 h, with average Co recovery of 98.12 %, dissolution of CoCO3 in sulphuric acid at pH 1, temperature of 25 degrees C for 2.5 h, with average Co recovery of 100 %, and electrowinning of Co at pH 3, temperature of 25 degrees C for 12 h with additions 30 g/l of Na2SO4 and 10 g/l boric acid to the electrolytic cell. The average purity of the Co cathode was 99.984% and total Co recovery and mass balance were 77.23% and 98.25%, respectively., A pilot study was conducted into the recovery of Co from the hot filter cake produced at the Zanjan Zn plant in Iran, where a leach-electrolysis process is employed for Zn production. The filter cake sample contained 20.44 %Zn, 0.32% Co, 2.65% Mn, 0.001% Ni, 0.02% Cd, 0.78% Fe, 8.57% Ca and 0.15% Pb. The process involved Zn washing with sulphuric acid at pH 1 at a temperature of 55 degrees C for 6 h, with Zn recovery and Co loss of 91.3% and 7.04%, respectively, reductive-acidic leaching of the Zn washing residue at pH 1 and a temperature of 35 degrees C for 3 h, with average Zn, Co and Mn recoveries of 74.36%, 82.34% and 81.92 %, respectively, Zn solvent extraction using 20% D2EHPA + 5 % TBP diluted to 25 vol% in kerosene at pH 3, organic/aqueous ratio of 1,temperature of 25 degrees C for 5 h with average Zn and Mn recoveries and Co loss of 99.96%, 99.72% and 1.38%, respectively, precipitation of Co as carbonate at pH 3 and temperature of 40 degrees C for 3 h, with average Co recovery of 98.12 %, dissolution of CoCO3 in sulphuric acid at pH 1, temperature of 25 degrees C for 2.5 h, with average Co recovery of 100 %, and electrowinning of Co at pH 3, temperature of 25 degrees C for 12 h with additions 30 g/l of Na2SO4 and 10 g/l boric acid to the electrolytic cell. The average purity of the Co cathode was 99.984% and total Co recovery and mass balance were 77.23% and 98.25%, respectively.

LEDs are made from semiconductor materials which combine Ga, Al and In with As, P and N, and other components contain metals such as Fe, Cu, Ni, Pb, Zn, Ag, Au and rare earth elements (REE). The recovery of critical and rare earth elements from spent LEDs is discussed in relation to future demand for these materials, in particular Ga, In, Ce, Eu, Gd and Y. There are three different approaches: direct re-use of lamp phosphors in new lamps, although this is only applicable to one type of fluorescent lamp; recycling of individual phosphor components, using physicochemical separation methods to pre-treat the phosphor mixture to separate the halophosphate phosphor fraction from the rare earth phosphors before recycling of its REE content; and chemical treatment of the phosphors with recovery of REE from solution by precipitation or solvent extraction. Separation of lamp phosphors by flotation is not easy because the components have similar hydrophobicity. The efficiency of the collectors dodecyl ammonium acetate, sodium dodecyl sulphate and sodium oleate has been tested at different pH values and the effect of sodium metasilicate as dispersant has also been investigated, with the results depending on the type of collector and pH. Two-liquid flotation using immiscible liquids consisting of water as a polar solvent and hexane, heptane or octane as non-polar solvent is more suitable for the separation of fine, less than 10 mm particles, and the wettability of the particles can be adjusted using a surfactant. Pneumatic separation in an air stream gave only moderate results. Separation of phosphor particles has been achieved using heavy medium separation, and the process was accelerated by centrifugation and particle pre-treatment with sodium oleate. Processes for the recovery of REEs from phosphors include leaching with HCl/H2O2, heating with sodium carbonate at 1000 degrees C, pressure leaching with H2SO4/HNO3, and extraction with supercritical CO2 containing tri-n-butyl ph, LEDs are made from semiconductor materials which combine Ga, Al and In with As, P and N, and other components contain metals such as Fe, Cu, Ni, Pb, Zn, Ag, Au and rare earth elements (REE). The recovery of critical and rare earth elements from spent LEDs is discussed in relation to future demand for these materials, in particular Ga, In, Ce, Eu, Gd and Y. There are three different approaches: direct re-use of lamp phosphors in new lamps, although this is only applicable to one type of fluorescent lamp; recycling of individual phosphor components, using physicochemical separation methods to pre-treat the phosphor mixture to separate the halophosphate phosphor fraction from the rare earth phosphors before recycling of its REE content; and chemical treatment of the phosphors with recovery of REE from solution by precipitation or solvent extraction. Separation of lamp phosphors by flotation is not easy because the components have similar hydrophobicity. The efficiency of the collectors dodecyl ammonium acetate, sodium dodecyl sulphate and sodium oleate has been tested at different pH values and the effect of sodium metasilicate as dispersant has also been investigated, with the results depending on the type of collector and pH. Two-liquid flotation using immiscible liquids consisting of water as a polar solvent and hexane, heptane or octane as non-polar solvent is more suitable for the separation of fine, less than 10 mm particles, and the wettability of the particles can be adjusted using a surfactant. Pneumatic separation in an air stream gave only moderate results. Separation of phosphor particles has been achieved using heavy medium separation, and the process was accelerated by centrifugation and particle pre-treatment with sodium oleate. Processes for the recovery of REEs from phosphors include leaching with HCl/H2O2, heating with sodium carbonate at 1000 degrees C, pressure leaching with H2SO4/HNO3, and extraction with supercritical CO2 containing tri-n-butyl ph

Nickelhutte Aue, Germany, specialises in the recycling of waste materials containing non-ferrous metals, including Ni-, Cu- and Co-bearing electroplating sludge, filter dust and ash, Ni, Cu and Co salts and solutions, Ni- and V-containing residues from oil combustion or gasification operations, spent Ni, Co, Ni-Mo, Co-Mo, Cu and V- and W-containing catalysts. Li-ion batteries, CuCl2 corrosive solutions, non-ferrous metal scrap and platinum group metal residues. Low metal-bearing material containing less than 5 wt% Ni, Cu and Co are treated in the smelting department, where valuable components such as Ni, Cu, Co and platinum group metals are concentrated in a matte while lighter elements such as Al, Si Mg and Ca form a slag. The sulphur-based smelting technology allows several types of sulphur-containing Ni, Cu and Co materials to be treated. In the treatment of platinum group metal-containing materials, the precious metals are collected in the matte together with Cu and Ni. More highly concentrated materials are subjected to calcining and roasting to burn off any organics and to remove impurities such as sulphur, halogens and traces of ammonia. Material produced by roasting and smelting is crushed and milled subjected to direct acid leaching or pressure leaching at 6-8 bar. Initial purification involves the precipitation of Fe. A combination of mixer settler and columns enables the handling of extremely polluted feed streams. After purification to remove undesired elements, Cu, Zn and Co are separated from Ni. Various solvent extraction circuits are used to produce cobalt sulphate, nickel carbonate, nickel chloride, nickel sulphate and copper sulphate. V-bearing slag is treated using purification stages including precipitation and ion-exchange to produce almost sulphur-free sodium vanadate solution. Corrosive CuCl2 solution from the production of printed circuit boards is converted into CuSO4 and Cu oxychloride and the extracted hydrochloric acid is used for the fur, Nickelhutte Aue, Germany, specialises in the recycling of waste materials containing non-ferrous metals, including Ni-, Cu- and Co-bearing electroplating sludge, filter dust and ash, Ni, Cu and Co salts and solutions, Ni- and V-containing residues from oil combustion or gasification operations, spent Ni, Co, Ni-Mo, Co-Mo, Cu and V- and W-containing catalysts. Li-ion batteries, CuCl2 corrosive solutions, non-ferrous metal scrap and platinum group metal residues. Low metal-bearing material containing less than 5 wt% Ni, Cu and Co are treated in the smelting department, where valuable components such as Ni, Cu, Co and platinum group metals are concentrated in a matte while lighter elements such as Al, Si Mg and Ca form a slag. The sulphur-based smelting technology allows several types of sulphur-containing Ni, Cu and Co materials to be treated. In the treatment of platinum group metal-containing materials, the precious metals are collected in the matte together with Cu and Ni. More highly concentrated materials are subjected to calcining and roasting to burn off any organics and to remove impurities such as sulphur, halogens and traces of ammonia. Material produced by roasting and smelting is crushed and milled subjected to direct acid leaching or pressure leaching at 6-8 bar. Initial purification involves the precipitation of Fe. A combination of mixer settler and columns enables the handling of extremely polluted feed streams. After purification to remove undesired elements, Cu, Zn and Co are separated from Ni. Various solvent extraction circuits are used to produce cobalt sulphate, nickel carbonate, nickel chloride, nickel sulphate and copper sulphate. V-bearing slag is treated using purification stages including precipitation and ion-exchange to produce almost sulphur-free sodium vanadate solution. Corrosive CuCl2 solution from the production of printed circuit boards is converted into CuSO4 and Cu oxychloride and the extracted hydrochloric acid is used for the fur

Ta is an important minor metal for the production of high-performance capacitors used in printed circuit boards (PCBs), and about 50%-60 % of annual Ta production is utilised for electrical applications. The recovery of Ta from solid tantalum electrolytic capacitors was investigated. Automotive control units from different manufacturers containing on average 400 ppm Ta were used as input materials for the experiments. The tests involved comminution using an impact crusher, selective comminution with a roll mill, density separation with a zigzag classifier, magnetic separation with an adjustable electromagnet, bulk processing using a weak magnetic over-belt separator, separation using strong magnetic fields with an electromagnetic circular separator, eddy current separation, cyanide leaching, cementation with Zn powder to recover precious metals, Cu and Ni or recovery by electrolysis. The results showed that comminution could be used to transfer 75% of the Ta to the 2-6 mm fraction and 96.8% of the waste material was removed and could be sent for conventional recycling. During sorting, Ta was collected in two concentrates with low losses of less than 10 %, and 80 % of the Au could be recovered as by-products. The Ta content of the concentrates was up to 30 % and 60%-70 % of the total Ta could be recovered., Ta is an important minor metal for the production of high-performance capacitors used in printed circuit boards (PCBs), and about 50%-60 % of annual Ta production is utilised for electrical applications. The recovery of Ta from solid tantalum electrolytic capacitors was investigated. Automotive control units from different manufacturers containing on average 400 ppm Ta were used as input materials for the experiments. The tests involved comminution using an impact crusher, selective comminution with a roll mill, density separation with a zigzag classifier, magnetic separation with an adjustable electromagnet, bulk processing using a weak magnetic over-belt separator, separation using strong magnetic fields with an electromagnetic circular separator, eddy current separation, cyanide leaching, cementation with Zn powder to recover precious metals, Cu and Ni or recovery by electrolysis. The results showed that comminution could be used to transfer 75% of the Ta to the 2-6 mm fraction and 96.8% of the waste material was removed and could be sent for conventional recycling. During sorting, Ta was collected in two concentrates with low losses of less than 10 %, and 80 % of the Au could be recovered as by-products. The Ta content of the concentrates was up to 30 % and 60%-70 % of the total Ta could be recovered.

Different residual fractions are produced during the beneficiation of waste from electrical and electronic equipment (WEEE), some of which contain high contents of valuable metals together with large amounts of organics in the form of plastics. Investigations were conducted using a fine fraction of filter dust with 98.5% of the particles smaller than 125 micrometres and containing 15.0-28.0 wt% Al2O3, 4.9-5.5 wt% CaO, 17.0-19.0 wt% SiO2, 1.1 wt% MgO, 1.3 wt% FeO/Fe2O3, 6.8-12.0 wt% Cu, 0.59-2.2 wt% Fe, 2.8-2.9 wt% Zn, 311-531ppm Au and 19.2-20.1 wt% C. To generate a material suitable for pyrometallurgical processing and to control the exothermic reaction due to the high amounts of carbon, the filter dust was mixed with low energy scrap in the form of Cu-containing residues and fluxes and binders, homogenised and pelletised. Thermochemical modelling was carried out using the phase diagram module of FactSage 7.0. Experiments were conducted using different pellet mixtures containing fine Cu dust as additive and molasses and water as binders. Melting was carried out in a resistance furnace heated to 800 degrees C and oxygen was blown into the crucible using a ceramic lance. After combustion, the crucible was held at 1 300-1 400 degrees C for 1 h. The pellets exhibited stable melting behaviour, and in most experiments a metal phase was formed at the bottom of the crucible. The mean Cu content of the metal phases was 91.1wt%, in good agreement with the FactSage modelling results. Addition of fine Cu dust to the mixture resulted in a much higher Cu content in the metal phases, probably due to the formation of metal droplets in the slag. The slag contained only 0.03 wt% C, indicating complete combustion of the plastics in the filter dust., Different residual fractions are produced during the beneficiation of waste from electrical and electronic equipment (WEEE), some of which contain high contents of valuable metals together with large amounts of organics in the form of plastics. Investigations were conducted using a fine fraction of filter dust with 98.5% of the particles smaller than 125 micrometres and containing 15.0-28.0 wt% Al2O3, 4.9-5.5 wt% CaO, 17.0-19.0 wt% SiO2, 1.1 wt% MgO, 1.3 wt% FeO/Fe2O3, 6.8-12.0 wt% Cu, 0.59-2.2 wt% Fe, 2.8-2.9 wt% Zn, 311-531ppm Au and 19.2-20.1 wt% C. To generate a material suitable for pyrometallurgical processing and to control the exothermic reaction due to the high amounts of carbon, the filter dust was mixed with low energy scrap in the form of Cu-containing residues and fluxes and binders, homogenised and pelletised. Thermochemical modelling was carried out using the phase diagram module of FactSage 7.0. Experiments were conducted using different pellet mixtures containing fine Cu dust as additive and molasses and water as binders. Melting was carried out in a resistance furnace heated to 800 degrees C and oxygen was blown into the crucible using a ceramic lance. After combustion, the crucible was held at 1 300-1 400 degrees C for 1 h. The pellets exhibited stable melting behaviour, and in most experiments a metal phase was formed at the bottom of the crucible. The mean Cu content of the metal phases was 91.1wt%, in good agreement with the FactSage modelling results. Addition of fine Cu dust to the mixture resulted in a much higher Cu content in the metal phases, probably due to the formation of metal droplets in the slag. The slag contained only 0.03 wt% C, indicating complete combustion of the plastics in the filter dust.

The recovery was investigated of rare earth elements from nickel metal hydride (NiMH) batteries. The most common electrode material contains a mixture of rare earth elements including La, Ce, Pr, Nd and Y. The active cathode material contains Ni(OH)2 and NiO, and the electrode material alloys may contain other metals such as Co and manganese Mn. The battery structure consists of a separator made of synthetic fabric such nylon, nickel-plated steel current collector, aluminium casing, polymer adhesives and printed circuit board. The batteries were subjected to crushing and sieving followed by sulphuric acid leaching in the presence of 0.05 mM sodium dodecylbenzenesulphonate as surfactant. The design of experiments was carried out using the MODDE software with H2SO4 concentrations of 1, 2 and 3 M, temperatures of 50, 70 and 90 degrees C and solid/liquid ratios of 1/7.75, 1/10 and 1/12.25. The results showed that the yields of Ni, Co, Fe, Mn and Cd were in the range 58%-100%, with Co and Ni leaching being maximum at later leaching times with Fe having slower leaching kinetics. Rare earth element concentrations in the leach solutions were 900-1 350 mg/l for Pr, 1 450-1 770 mg/l for Ce and 2 270 to 2 870 mg/l for La. The leach residue contained fibrous material, emphasizing the need for proper mechanical treatment to separate plastics from metals prior to hydrometallurgical processing., The recovery was investigated of rare earth elements from nickel metal hydride (NiMH) batteries. The most common electrode material contains a mixture of rare earth elements including La, Ce, Pr, Nd and Y. The active cathode material contains Ni(OH)2 and NiO, and the electrode material alloys may contain other metals such as Co and manganese Mn. The battery structure consists of a separator made of synthetic fabric such nylon, nickel-plated steel current collector, aluminium casing, polymer adhesives and printed circuit board. The batteries were subjected to crushing and sieving followed by sulphuric acid leaching in the presence of 0.05 mM sodium dodecylbenzenesulphonate as surfactant. The design of experiments was carried out using the MODDE software with H2SO4 concentrations of 1, 2 and 3 M, temperatures of 50, 70 and 90 degrees C and solid/liquid ratios of 1/7.75, 1/10 and 1/12.25. The results showed that the yields of Ni, Co, Fe, Mn and Cd were in the range 58%-100%, with Co and Ni leaching being maximum at later leaching times with Fe having slower leaching kinetics. Rare earth element concentrations in the leach solutions were 900-1 350 mg/l for Pr, 1 450-1 770 mg/l for Ce and 2 270 to 2 870 mg/l for La. The leach residue contained fibrous material, emphasizing the need for proper mechanical treatment to separate plastics from metals prior to hydrometallurgical processing.

The materials handling and process-control systems are described available from Schenck Process UK. The company can supply static and dynamic weighing systems, screening and vibratory feeding equipment, mechanical conveying equipment including drag chain conveyor, bucket elevators, bulk reception units and scrap charging technology, pneumatic conveyors and pneumatic injection equipment. Applications in hydrometallurgical and pyrometallurgical plants are described., The materials handling and process-control systems are described available from Schenck Process UK. The company can supply static and dynamic weighing systems, screening and vibratory feeding equipment, mechanical conveying equipment including drag chain conveyor, bucket elevators, bulk reception units and scrap charging technology, pneumatic conveyors and pneumatic injection equipment. Applications in hydrometallurgical and pyrometallurgical plants are described.

The potential advantages were investigated of employing nozzles for furnace firing, based on computational fluid dynamics modelling. Simulations were conducted of the heat transfer in a rotary anode furnace used to refine blister copper from a direct-to-blister flash furnace, and in a converter type rotary furnace used to reduce converter slags from Cu-Pb-Fe alloy treatment. The heat exchange in the anode furnace was considered to be a steady state process, and in the converter was assumed to be a transient process due to the ongoing reduction process. The results were applied to a rotary anode furnace fired with a burner and two nozzles supplying natural gas fuel and to a converter equipped with different numbers of nozzles. The application of submerged nozzles together with the gas burner in the anode furnace resulted in a reduction in fuel consumption. A higher average bath temperature of 1 219 degrees C was achieved for a system equipped with a burner supplying 100 Nm3/h of fuel and two nozzles each supplying 50 Nm3/h, compared with 1 076 degrees C when fired only with a burner supplying 200 Nm3/h of fuel. The use of the nozzles also improved the uniformity of the temperature field in the bath. The standard deviation for the average bath temperature was 18.2 degrees C compared with 26.3 degrees using only a burner. For the same volume of fuel consumed, the heat utilisation efficiency was higher when larger amounts of gas were delivered through the nozzles. The average bath temperature was 1 219 degrees C when 100 Nm3/h of gas was delivered through the burner and 50 Nm3/h through each nozzle compared with 1 205 degrees C for a system supplying 150 Nm3/h through the burner and 25 Nm3/h through each nozzle. For the converter slag reduction process, the highest average slag temperature of 1 227 degrees C was obtained using a furnace equipped with 4 nozzles with a standard deviation of 4.99, compared with 1 212 degrees C and a standard deviation of 6.36 using 3 nozzl, The potential advantages were investigated of employing nozzles for furnace firing, based on computational fluid dynamics modelling. Simulations were conducted of the heat transfer in a rotary anode furnace used to refine blister copper from a direct-to-blister flash furnace, and in a converter type rotary furnace used to reduce converter slags from Cu-Pb-Fe alloy treatment. The heat exchange in the anode furnace was considered to be a steady state process, and in the converter was assumed to be a transient process due to the ongoing reduction process. The results were applied to a rotary anode furnace fired with a burner and two nozzles supplying natural gas fuel and to a converter equipped with different numbers of nozzles. The application of submerged nozzles together with the gas burner in the anode furnace resulted in a reduction in fuel consumption. A higher average bath temperature of 1 219 degrees C was achieved for a system equipped with a burner supplying 100 Nm3/h of fuel and two nozzles each supplying 50 Nm3/h, compared with 1 076 degrees C when fired only with a burner supplying 200 Nm3/h of fuel. The use of the nozzles also improved the uniformity of the temperature field in the bath. The standard deviation for the average bath temperature was 18.2 degrees C compared with 26.3 degrees using only a burner. For the same volume of fuel consumed, the heat utilisation efficiency was higher when larger amounts of gas were delivered through the nozzles. The average bath temperature was 1 219 degrees C when 100 Nm3/h of gas was delivered through the burner and 50 Nm3/h through each nozzle compared with 1 205 degrees C for a system supplying 150 Nm3/h through the burner and 25 Nm3/h through each nozzle. For the converter slag reduction process, the highest average slag temperature of 1 227 degrees C was obtained using a furnace equipped with 4 nozzles with a standard deviation of 4.99, compared with 1 212 degrees C and a standard deviation of 6.36 using 3 nozzl

The use of water as a furnace cooling medium poses risks as a leakage may result in an immediate explosion and severe damage to the furnace structure. The ILTEC ionic liquid cooling technology developed by Mettop, Austria, eliminates almost all the negative effects of a water-cooled system and provides additional benefits. The ionic liquid IL-B2001 is non-explosive, completely decomposing into the gaseous components at temperature of 450 degrees C and above, non-corrosive, non-flammable, non-toxic has a wide operational temperature range of 50-200 degrees and high electrical conductivity. A fundamental characteristic of ILTEC is the use of a closed-circuit loop for the ionic liquid so that there is no contact between the ionic liquid and water and/or air and the ionic liquid is non-consumable. Examples of application described include tap hole cooling, electric furnace side wall cooling involving the creation of a freeze lining and cooling of off-gas systems., The use of water as a furnace cooling medium poses risks as a leakage may result in an immediate explosion and severe damage to the furnace structure. The ILTEC ionic liquid cooling technology developed by Mettop, Austria, eliminates almost all the negative effects of a water-cooled system and provides additional benefits. The ionic liquid IL-B2001 is non-explosive, completely decomposing into the gaseous components at temperature of 450 degrees C and above, non-corrosive, non-flammable, non-toxic has a wide operational temperature range of 50-200 degrees and high electrical conductivity. A fundamental characteristic of ILTEC is the use of a closed-circuit loop for the ionic liquid so that there is no contact between the ionic liquid and water and/or air and the ionic liquid is non-consumable. Examples of application described include tap hole cooling, electric furnace side wall cooling involving the creation of a freeze lining and cooling of off-gas systems.

Hard metal drill bits used for oil and gas extraction contain tungsten, tungsten compounds such as tungsten carbide or cast tungsten carbide with a complex binder alloy. A process for the recycling of drill bit scrap was investigated in which the binder phase is selectively digested while the tungsten compounds remain in the form of a pure solid residue. Experiments were conducted using the tail ends of transverse rupture strength test bars containing an upper section of infiltrated tungsten carbide similar to a drill head and containing 51.1 wt% W, 23.3 wt% Cu, 12.9 wt% Mn, 6.5 wt% Ni, 3.1 wt% Zn, 3.1 wt% C, 0.01 wt% Ti, 0.03 wt% Ca and 0.01 wt% Cr, and a lower section made of a binder alloy consisting of Cu, Mn, Ni and Zn. The drill bit models were cut into narrow cross-sections, embedded in a polymer resin and polished. Part of the metallic surface was then covered with chemically resistant lacquer to form a reference level for measurement of the penetration depth. Leaching tests were conducted at 50 degrees C for 4 h at ambient pressure using acid and alkaline solutions, either alone or with the addition of oxidising agents such as hydrogen peroxide. Oxygen was also bubbled through selected acids. Selective digestion was not feasible with non-oxidising acids, but the addition of hydrogen peroxide resulted in more complete attack. The addition of potassium permanganate and ferric chloride to sulphuric acid did not significantly enhance digestion. Injection of oxygen only improved digestion for the strongly oxidising nitric acid. Alkaline solutions such as sodium hydroxide and ammonia did not exhibit much attack even with the addition of hydrogen peroxide. Organic acids did not result in digestion in the absence of hydrogen peroxide, and no, slight and noticeable digestion was observed in oxalic, lactic acid and formic acids, respectively, with the addition of the oxidising agent. The strongest digestion was obtained using strongly oxidising nitric acid. The binde, Hard metal drill bits used for oil and gas extraction contain tungsten, tungsten compounds such as tungsten carbide or cast tungsten carbide with a complex binder alloy. A process for the recycling of drill bit scrap was investigated in which the binder phase is selectively digested while the tungsten compounds remain in the form of a pure solid residue. Experiments were conducted using the tail ends of transverse rupture strength test bars containing an upper section of infiltrated tungsten carbide similar to a drill head and containing 51.1 wt% W, 23.3 wt% Cu, 12.9 wt% Mn, 6.5 wt% Ni, 3.1 wt% Zn, 3.1 wt% C, 0.01 wt% Ti, 0.03 wt% Ca and 0.01 wt% Cr, and a lower section made of a binder alloy consisting of Cu, Mn, Ni and Zn. The drill bit models were cut into narrow cross-sections, embedded in a polymer resin and polished. Part of the metallic surface was then covered with chemically resistant lacquer to form a reference level for measurement of the penetration depth. Leaching tests were conducted at 50 degrees C for 4 h at ambient pressure using acid and alkaline solutions, either alone or with the addition of oxidising agents such as hydrogen peroxide. Oxygen was also bubbled through selected acids. Selective digestion was not feasible with non-oxidising acids, but the addition of hydrogen peroxide resulted in more complete attack. The addition of potassium permanganate and ferric chloride to sulphuric acid did not significantly enhance digestion. Injection of oxygen only improved digestion for the strongly oxidising nitric acid. Alkaline solutions such as sodium hydroxide and ammonia did not exhibit much attack even with the addition of hydrogen peroxide. Organic acids did not result in digestion in the absence of hydrogen peroxide, and no, slight and noticeable digestion was observed in oxalic, lactic acid and formic acids, respectively, with the addition of the oxidising agent. The strongest digestion was obtained using strongly oxidising nitric acid. The binde

The use was investigated of a UASB anaerobic bioreactor for the removal of sulphate and Mn from wastewaters such as acid mine drainage. The reactor was fed with an industrial waste mine water containing Mn supplemented with synthetic Postgate C medium using lactate as a carbon and electron source. During the Mn removal experiments, the reactor was fed daily with mine water dosed with nutrients to achieve a composition consisting of 0.5 g/l KH2PO4, 1.0 g/l NH4Cl, 0.06 g/l MgSO4.7 H2O, 0.1 g/l FeSO4.7 H2O, 0.25 g/l yeast extract, 2.96 g/l Na2SO4 and 3.76 g/l lactate to stimulate sulphur-reducing bacteria growth. The Mn content was in the range 82-145 mg/l due to dilution and possibly Mn precipitation by phosphate. Batch precipitation experiments were then conducted on Cu- and Ni-containing solutions which were mixed with the biogenic sulphide solution containing 98 mg/l total sulphide obtained as the effluent from the UASB reactor. The reactor achieved sulphate and chemical oxygen demand removal efficiencies of 80 +/- 8 % and 48 +/- 13 %, respectively, and an alkalinity of 1 246 +/- 283 mg HCO(3-)/l, with 86 % Mn removal as MnCO3 and an average residual Mn concentration of 15 mg/l. The use of the biogenic sulphide at a stoichiometric ratio of 1/1 enabled almost complete Cu precipitation at pH 2 and above, with similar results for Ni at pH 5. Ni precipitation also depended on the composition of the biogenic sulphide-bearing solution, as species such as phosphate may also remove the metal from solution. In continuous experiments, the concentrations of the metals were reduced to below 0.5 mg/l, with CuS and NiS being the compounds formed., The use was investigated of a UASB anaerobic bioreactor for the removal of sulphate and Mn from wastewaters such as acid mine drainage. The reactor was fed with an industrial waste mine water containing Mn supplemented with synthetic Postgate C medium using lactate as a carbon and electron source. During the Mn removal experiments, the reactor was fed daily with mine water dosed with nutrients to achieve a composition consisting of 0.5 g/l KH2PO4, 1.0 g/l NH4Cl, 0.06 g/l MgSO4.7 H2O, 0.1 g/l FeSO4.7 H2O, 0.25 g/l yeast extract, 2.96 g/l Na2SO4 and 3.76 g/l lactate to stimulate sulphur-reducing bacteria growth. The Mn content was in the range 82-145 mg/l due to dilution and possibly Mn precipitation by phosphate. Batch precipitation experiments were then conducted on Cu- and Ni-containing solutions which were mixed with the biogenic sulphide solution containing 98 mg/l total sulphide obtained as the effluent from the UASB reactor. The reactor achieved sulphate and chemical oxygen demand removal efficiencies of 80 +/- 8 % and 48 +/- 13 %, respectively, and an alkalinity of 1 246 +/- 283 mg HCO(3-)/l, with 86 % Mn removal as MnCO3 and an average residual Mn concentration of 15 mg/l. The use of the biogenic sulphide at a stoichiometric ratio of 1/1 enabled almost complete Cu precipitation at pH 2 and above, with similar results for Ni at pH 5. Ni precipitation also depended on the composition of the biogenic sulphide-bearing solution, as species such as phosphate may also remove the metal from solution. In continuous experiments, the concentrations of the metals were reduced to below 0.5 mg/l, with CuS and NiS being the compounds formed.

The batteries were disassembled and the steel discs, paper and copper caps removed. The Zn and black paste components were rinsed with warm water, filtered and dried, and the black paste was separated by sieving. Samples of the black paste containing manganese oxides such as pyrolusite were heated in an inert atmosphere at temperatures up to 1 050 degrees C to reduce MnO2 to Mn metal using the carbon present in the black paste as reducing agent. The corroded Zn cans were melted at 500 degrees C and purified by skimming. Molten Mn was then alloyed with the Zn. Samples of the corroded Zn cans and black paste were also leached with sulphuric acid to obtain the respective sulphate salts. The solution was filtered, neutralised with ammonia and treated with hydrazine hydrate to yield ultra-fine particles of Zn and Mn metals. Alternatively, the two metal ions in solution were co-electrodeposited at pH 3.5-4.5 at a current density of 12 mA/cm2 and 20-24 mV DC using two graphite electrodes. The co-deposited metals were washed and re-melted under an inert atmosphere to obtain the Zn-Mn alloy. The recovery efficiencies of the chemical and electrodeposition methods were 82% and 94%, respectively. The maximum Mn content of the alloy was 15.1% using electrodeposition, and alloys with a higher Mn content could only be obtained with the chemical method., The batteries were disassembled and the steel discs, paper and copper caps removed. The Zn and black paste components were rinsed with warm water, filtered and dried, and the black paste was separated by sieving. Samples of the black paste containing manganese oxides such as pyrolusite were heated in an inert atmosphere at temperatures up to 1 050 degrees C to reduce MnO2 to Mn metal using the carbon present in the black paste as reducing agent. The corroded Zn cans were melted at 500 degrees C and purified by skimming. Molten Mn was then alloyed with the Zn. Samples of the corroded Zn cans and black paste were also leached with sulphuric acid to obtain the respective sulphate salts. The solution was filtered, neutralised with ammonia and treated with hydrazine hydrate to yield ultra-fine particles of Zn and Mn metals. Alternatively, the two metal ions in solution were co-electrodeposited at pH 3.5-4.5 at a current density of 12 mA/cm2 and 20-24 mV DC using two graphite electrodes. The co-deposited metals were washed and re-melted under an inert atmosphere to obtain the Zn-Mn alloy. The recovery efficiencies of the chemical and electrodeposition methods were 82% and 94%, respectively. The maximum Mn content of the alloy was 15.1% using electrodeposition, and alloys with a higher Mn content could only be obtained with the chemical method.

In the context of the European Commission's initiative to develop the measurement of environmental footprints that will allow producers to be compared, the pilot organisation environmental footprint (OEF) project on Cu production is currently ongoing and is expected to deliver the final OEF sector rules by the end of 2017. An organisational life-cycle assessment (LCA), compared with a product LCA, considers the organisational activities as a whole: this approach allows the most important environmental issues to be identified, both in the sector and the organisation itself. It is useful for an organisation to rely on a robust measurement tool, such as the OEF, which can be implemented in other environmental management schemes. This helps with prioritising environmental management actions and identifying improvement potential, both within the organisational boundaries (direct activities) and outside (upstream and downstream, indirect activities). The focus, therefore, is mainly on performance tracking and improvement. The OEF sector rules for Cu production cover the pyrometallurgical route. The product portfolio of a copper-producing company may cover many products, all directly linked to Cu production: cathodes, anodes and blister, H2SO4, Fe2SiO4, anode slime, NiSO4, CuSO4, Ag, Au, PGE concentrate, Pb, Pb-Sn alloys, Sn, crude Se and Te, and ZnO. The selected approach is to assess each production route (primary, secondary or integrated multi-metal pyrometallurgy) on the basis of a real organisation in order to take into proper consideration the differences in process configuration and product portfolio. All life-cycle stages from cradle to gate should be included in the OEF boundary but the usage stage and end of life excluded. The results and experience gained in this process were used to understand the most relevant life-cycle stages, processes and environmental impacts, such as influence on climate change or ozone depletion. The next step was to test these rules in, In the context of the European Commission's initiative to develop the measurement of environmental footprints that will allow producers to be compared, the pilot organisation environmental footprint (OEF) project on Cu production is currently ongoing and is expected to deliver the final OEF sector rules by the end of 2017. An organisational life-cycle assessment (LCA), compared with a product LCA, considers the organisational activities as a whole: this approach allows the most important environmental issues to be identified, both in the sector and the organisation itself. It is useful for an organisation to rely on a robust measurement tool, such as the OEF, which can be implemented in other environmental management schemes. This helps with prioritising environmental management actions and identifying improvement potential, both within the organisational boundaries (direct activities) and outside (upstream and downstream, indirect activities). The focus, therefore, is mainly on performance tracking and improvement. The OEF sector rules for Cu production cover the pyrometallurgical route. The product portfolio of a copper-producing company may cover many products, all directly linked to Cu production: cathodes, anodes and blister, H2SO4, Fe2SiO4, anode slime, NiSO4, CuSO4, Ag, Au, PGE concentrate, Pb, Pb-Sn alloys, Sn, crude Se and Te, and ZnO. The selected approach is to assess each production route (primary, secondary or integrated multi-metal pyrometallurgy) on the basis of a real organisation in order to take into proper consideration the differences in process configuration and product portfolio. All life-cycle stages from cradle to gate should be included in the OEF boundary but the usage stage and end of life excluded. The results and experience gained in this process were used to understand the most relevant life-cycle stages, processes and environmental impacts, such as influence on climate change or ozone depletion. The next step was to test these rules in

Scorodite can be used as a final product for the storage of arsenic removed from circuits in copper smelters. It is characteried by very low solubility and high stability over a broad pH range and is the most stable combination of iron and arsenic. The effects were investigated of temperature, reaction time, scorodite seeding, initial arsenic concentration and initial copper concentration on the precipitation of scorodite at atmospheric pressure. The results showed that a solid product containing scorodite with particle size of 15-35 micrometres and low arsenic leaching rate of 1.0 mg/kg) was obtained at temperatures above 95 degrees C, Fe(III)/As(V) molar ratio in the range 1:1 to 1.3 :1, reaction time slightly longer than 6 h, scorodite seeding of more than 40 g/l and As(V) concentration of 5-15 g/l. The Cu concentration in the feed solution had no effect on product quality., Scorodite can be used as a final product for the storage of arsenic removed from circuits in copper smelters. It is characteried by very low solubility and high stability over a broad pH range and is the most stable combination of iron and arsenic. The effects were investigated of temperature, reaction time, scorodite seeding, initial arsenic concentration and initial copper concentration on the precipitation of scorodite at atmospheric pressure. The results showed that a solid product containing scorodite with particle size of 15-35 micrometres and low arsenic leaching rate of 1.0 mg/kg) was obtained at temperatures above 95 degrees C, Fe(III)/As(V) molar ratio in the range 1:1 to 1.3 :1, reaction time slightly longer than 6 h, scorodite seeding of more than 40 g/l and As(V) concentration of 5-15 g/l. The Cu concentration in the feed solution had no effect on product quality.

Outotec's innovative solution-cooling tower design, incorporating horizontal air/vapour outflow technology, has been in use since 2012 for spent electrolyte cooling in a Zn refinery. An important additional benefit provided by the new tower design is a reduction in the emissions contained in the entrained liquid droplets carried by the cooling air/vapour when it exits in the tower. Depending on the solution being cooled, these entrained droplets can include acids, dissolved metal ions, dissolved salts, suspended solids and other compounds. The emissions cause metal losses, local environmental and occupational health and safety issues as well as increasing corrosion to equipment, structures and potentially private property in the vicinity. The main method to control these emissions is the installation of demisters to the tower; in the Outotec tower, two layers of vaned or chevron demisters are used, one coarse and the other fine, with the horizontal flow configuration allowing considerably greater velocities through the demister, reducing entrainment. Detailed testing was performed on an Outotec tower using isokinetic sampling to quantify the emissions. The test work provided promising results and demonstrated reduced emission levels, but some discrepancies in the Zn:S ratio results require further investigation. The emission levels were noted to be considerably lower than in the typically used vertical outflow cooling towers. The findings of the measurements are presented and the methods used for preliminary emission estimation discussed., Outotec's innovative solution-cooling tower design, incorporating horizontal air/vapour outflow technology, has been in use since 2012 for spent electrolyte cooling in a Zn refinery. An important additional benefit provided by the new tower design is a reduction in the emissions contained in the entrained liquid droplets carried by the cooling air/vapour when it exits in the tower. Depending on the solution being cooled, these entrained droplets can include acids, dissolved metal ions, dissolved salts, suspended solids and other compounds. The emissions cause metal losses, local environmental and occupational health and safety issues as well as increasing corrosion to equipment, structures and potentially private property in the vicinity. The main method to control these emissions is the installation of demisters to the tower; in the Outotec tower, two layers of vaned or chevron demisters are used, one coarse and the other fine, with the horizontal flow configuration allowing considerably greater velocities through the demister, reducing entrainment. Detailed testing was performed on an Outotec tower using isokinetic sampling to quantify the emissions. The test work provided promising results and demonstrated reduced emission levels, but some discrepancies in the Zn:S ratio results require further investigation. The emission levels were noted to be considerably lower than in the typically used vertical outflow cooling towers. The findings of the measurements are presented and the methods used for preliminary emission estimation discussed.

The precipitation was investigated of rare earth elements from chloride media using oxalic acid. The chloride solution was obtained by treating flotation concentrate from the processing of phosphate ore from Esfordi, Iran, by alkali digestion, hydrochloric acid leaching and neutralisation of iron. The solution contained 525 ppm Ce, 209 ppm Nd, 136 ppm La, 31 ppm Sm, 63 ppm Y and 63 ppm Pr. Four process variables were studied, oxalic acid concentration, stirring speed, temperature and time. The maximum rare earth element precipitation was obtained using four times the stoichiometric amount of oxalic acid, stirring speed of 700 rpm, temperature of 60 degrees C and reaction time of 60 minutes. The precipitation efficiencies of Ce, Nd, La, Sm, Y and Pr were 99.62%, 99.52%, 97.79%, 96.77%, 98.41% and 90.74%, respectively, and the Ce, Nd, La, Sm, Y and Pr grades in the oxalate concentrate were was 18.8%, 7.48%, 4.87%, 1.11%, 2.26% and 1.93%, respectively., The precipitation was investigated of rare earth elements from chloride media using oxalic acid. The chloride solution was obtained by treating flotation concentrate from the processing of phosphate ore from Esfordi, Iran, by alkali digestion, hydrochloric acid leaching and neutralisation of iron. The solution contained 525 ppm Ce, 209 ppm Nd, 136 ppm La, 31 ppm Sm, 63 ppm Y and 63 ppm Pr. Four process variables were studied, oxalic acid concentration, stirring speed, temperature and time. The maximum rare earth element precipitation was obtained using four times the stoichiometric amount of oxalic acid, stirring speed of 700 rpm, temperature of 60 degrees C and reaction time of 60 minutes. The precipitation efficiencies of Ce, Nd, La, Sm, Y and Pr were 99.62%, 99.52%, 97.79%, 96.77%, 98.41% and 90.74%, respectively, and the Ce, Nd, La, Sm, Y and Pr grades in the oxalate concentrate were was 18.8%, 7.48%, 4.87%, 1.11%, 2.26% and 1.93%, respectively.

The recovery was investigated of rare earth elements from a monazite concentrate containing 14.36 wt% Ce, 8.65 wt% La, 16.25wt% Fe and 7.24 wt% P. Dissolution experiments were conducted using tartaric, citric, oxalic and maleic acids. The effects of pH, residence time and the use of tailings mine site water were also studied. Oxalic acid exhibited the highest dissolution potential and was selected for further study. The results showed that the effectiveness of each acid correlated fairly well with their ability to bind rare earths, but citric acid exhibited greater complex stability with rare earths than oxalic acid indicating that reprecipitation of the oxalate is important for increased mineral dissolution. The system exhibited surprising selectivity for the dissolution of monazite over the iron phase, as oxalic acid has a greater binding strength with iron than with rare earths. This may be due to other interactions with the phosphate, the different stability of the specific phases, a kinetic limitation on iron dissolution or preferential adsorption to the monazite surface. The conversion of monazite to rare earth oxalate appeared to exhibit a rate limitation with time possibly due to the precipitating material forming a passive layer on the surface of the monazite or consumption of the oxalate by dissolving ions. A rare earth hydroxide could be produced from the leach residue, allowing the process to be applied to conventional solvent extraction. Tailings water used to investigate the effect of dissolved solids on the process showed a marked decrease in phosphorus release, suggesting a significant decrease in monazite conversion., The recovery was investigated of rare earth elements from a monazite concentrate containing 14.36 wt% Ce, 8.65 wt% La, 16.25wt% Fe and 7.24 wt% P. Dissolution experiments were conducted using tartaric, citric, oxalic and maleic acids. The effects of pH, residence time and the use of tailings mine site water were also studied. Oxalic acid exhibited the highest dissolution potential and was selected for further study. The results showed that the effectiveness of each acid correlated fairly well with their ability to bind rare earths, but citric acid exhibited greater complex stability with rare earths than oxalic acid indicating that reprecipitation of the oxalate is important for increased mineral dissolution. The system exhibited surprising selectivity for the dissolution of monazite over the iron phase, as oxalic acid has a greater binding strength with iron than with rare earths. This may be due to other interactions with the phosphate, the different stability of the specific phases, a kinetic limitation on iron dissolution or preferential adsorption to the monazite surface. The conversion of monazite to rare earth oxalate appeared to exhibit a rate limitation with time possibly due to the precipitating material forming a passive layer on the surface of the monazite or consumption of the oxalate by dissolving ions. A rare earth hydroxide could be produced from the leach residue, allowing the process to be applied to conventional solvent extraction. Tailings water used to investigate the effect of dissolved solids on the process showed a marked decrease in phosphorus release, suggesting a significant decrease in monazite conversion.

The extraction and separation were investigated of Cr and other base metals from a reference material, SARM 131, and from samples of UG2 Reef chromitite and Merensky Reef ores from South Africa. The samples contained 28.64% 64.85% and 16.61 % Cr, respectively. The samples were digested using Na2HPO4/NaH2PO4.H2O in the ratio 1:25 at 800 degrees C for about 20 minutes and the resultant green melts were cooled to room temperature and dissolved in deionised water to yield clear green solutions. Isolation of Cr present in the solutions was investigated using basic solutions of NaOH, KOH and NH4OH. The addition of NaOH and KOH and NH4OH at pH 6.5-6.9 to the solutions of SARM 131 yielded newly formed grey-green gelatinous precipitates, which were also similar in colour and impurity matrix to the grey-green gelatinous precipitates of the UG2 chromitite and the Merensky Reef ores. Qualitative analyses of the product isolated from the second precipitation process indicated complete separation of Cr from the initial green solution using the different bases. Separation of Cu and Ni from the grey-green gelatinous precipitates was conducted using concentrated NH4OH. Crystals were isolated from the blue filtrate after the solution was left to stand for 48 h, indicating the formation of a mixed product containing Ni and Cu in the ratio 3:1. The results showed that the use of NH4OH as a selective extractant or leaching agent for Cu and Ni can be cost effective, rapid and efficient, with recoveries of up to 95% for both metals., The extraction and separation were investigated of Cr and other base metals from a reference material, SARM 131, and from samples of UG2 Reef chromitite and Merensky Reef ores from South Africa. The samples contained 28.64% 64.85% and 16.61 % Cr, respectively. The samples were digested using Na2HPO4/NaH2PO4.H2O in the ratio 1:25 at 800 degrees C for about 20 minutes and the resultant green melts were cooled to room temperature and dissolved in deionised water to yield clear green solutions. Isolation of Cr present in the solutions was investigated using basic solutions of NaOH, KOH and NH4OH. The addition of NaOH and KOH and NH4OH at pH 6.5-6.9 to the solutions of SARM 131 yielded newly formed grey-green gelatinous precipitates, which were also similar in colour and impurity matrix to the grey-green gelatinous precipitates of the UG2 chromitite and the Merensky Reef ores. Qualitative analyses of the product isolated from the second precipitation process indicated complete separation of Cr from the initial green solution using the different bases. Separation of Cu and Ni from the grey-green gelatinous precipitates was conducted using concentrated NH4OH. Crystals were isolated from the blue filtrate after the solution was left to stand for 48 h, indicating the formation of a mixed product containing Ni and Cu in the ratio 3:1. The results showed that the use of NH4OH as a selective extractant or leaching agent for Cu and Ni can be cost effective, rapid and efficient, with recoveries of up to 95% for both metals.

The extraction was investigated of In from a solution obtained from the biologicalleaching of sphalerite from the Reiche Zeche mine in Freiberg, Germany. The development of the process was initially based on a model solution containing 10 mg/dm3 In and 10 g/dm3 each of Fe and Zn. The process involves nine steps. In-containing solution is subjected to liquid-liquid extraction with 0.05 mol/dm3 DEHPA to concentrate and separate In from substantial amounts of accompanying elements. If present, Fe must be reduced to Fe(II) before extraction. The behaviour of DEHPA with respect to other metals was also tested using Ni, Cr-, Co- and Mn-containing solutions and the results showed that, unlike Zn and Fe, none of the other metals were extracted under the conditions used. After phase separation, an In-free aqueous phase and an In-rich organic phase remain. The organic phase is washed with dilute sulphuric acid to selectively separate Zn, and the organic phase, which still contains approximately the same amount of Fe in addition to In, is mixed with 3.5 mol/dm3 hydrochloric acid re-extraction solution at an organic/aqueous phase ratio of 2/1 and the metals transferred to the aqueous phase. After phase separation, an In- and Fe-containing re-extract and a regenerated organic phase remain, with the latter reused for extraction in the first step. The chloride-containing metal-rich phase is treated with Cyanex 923 at an organic/aqueous phase ratio of 20, with the metals passing into the organic phase and a metal-free hydrochloric acid solution obtained which can be used for the re-extraction step. The organic phase is washed several times with dilute sulphuric acid to separate more than 90% of the Fe and is then subjected to re-extraction with 3.1 mol/dm3 nitric acid. Downstream diffusion dialysis is used to regenerate the nitric acid for reuse, with a recovery of more 95 %. Indium is then separated from the remaining Fe using pH-dependent precipitation with Na2S or H2S. Following, The extraction was investigated of In from a solution obtained from the biologicalleaching of sphalerite from the Reiche Zeche mine in Freiberg, Germany. The development of the process was initially based on a model solution containing 10 mg/dm3 In and 10 g/dm3 each of Fe and Zn. The process involves nine steps. In-containing solution is subjected to liquid-liquid extraction with 0.05 mol/dm3 DEHPA to concentrate and separate In from substantial amounts of accompanying elements. If present, Fe must be reduced to Fe(II) before extraction. The behaviour of DEHPA with respect to other metals was also tested using Ni, Cr-, Co- and Mn-containing solutions and the results showed that, unlike Zn and Fe, none of the other metals were extracted under the conditions used. After phase separation, an In-free aqueous phase and an In-rich organic phase remain. The organic phase is washed with dilute sulphuric acid to selectively separate Zn, and the organic phase, which still contains approximately the same amount of Fe in addition to In, is mixed with 3.5 mol/dm3 hydrochloric acid re-extraction solution at an organic/aqueous phase ratio of 2/1 and the metals transferred to the aqueous phase. After phase separation, an In- and Fe-containing re-extract and a regenerated organic phase remain, with the latter reused for extraction in the first step. The chloride-containing metal-rich phase is treated with Cyanex 923 at an organic/aqueous phase ratio of 20, with the metals passing into the organic phase and a metal-free hydrochloric acid solution obtained which can be used for the re-extraction step. The organic phase is washed several times with dilute sulphuric acid to separate more than 90% of the Fe and is then subjected to re-extraction with 3.1 mol/dm3 nitric acid. Downstream diffusion dialysis is used to regenerate the nitric acid for reuse, with a recovery of more 95 %. Indium is then separated from the remaining Fe using pH-dependent precipitation with Na2S or H2S. Following

The recovery was investigated of Sc from waste slags generated in the processing of oxidised nickel ores, with reference to operations at the Rezh plant in Sverdlovsk, Russia, where the waste dumps are 50-60 m high, cover an area of tens of square kilometres and amount to a few hundred million tons. Typical slag compositions contain 2.598% Al, 6.416% Na, 8.613% Ca, 0.008% Sc, 0.113% Ti, 17.853% Fe, 0.005% Y, 0.006% Zr, 0.003% Mo, 0.0005% Th and 0.0003% U. Sulphuric acid leaching experiments on slag samples showed that the optimum conditions were a leaching time of 1.5 h, sulphuric acid concentration of 100 g/l, solid/liquid ratio of 1:10 and temperature of 20-40 degrees, which resulted in Sc recovery of more than 90%. Carbonate leaching using Na2CO3, NaHCO3 or propylene carbonate was shown to be ineffective. Sulphuric acid leaching results in the release of large amounts of silicic acid in the leaching system, and hyper-crosslinked polystyrene-based sorbents were used to separate the silicic acid from the sulphate leach solution. The sorption properties of Purolite MN-202 containing no ionic groups and Purolite MN-102 containing tertiary amine groups were investigated. The results showed that weakly basic anion exchangers based on hyper-crosslinked polystyrene are preferred for cleaning the slag leaching solutions., The recovery was investigated of Sc from waste slags generated in the processing of oxidised nickel ores, with reference to operations at the Rezh plant in Sverdlovsk, Russia, where the waste dumps are 50-60 m high, cover an area of tens of square kilometres and amount to a few hundred million tons. Typical slag compositions contain 2.598% Al, 6.416% Na, 8.613% Ca, 0.008% Sc, 0.113% Ti, 17.853% Fe, 0.005% Y, 0.006% Zr, 0.003% Mo, 0.0005% Th and 0.0003% U. Sulphuric acid leaching experiments on slag samples showed that the optimum conditions were a leaching time of 1.5 h, sulphuric acid concentration of 100 g/l, solid/liquid ratio of 1:10 and temperature of 20-40 degrees, which resulted in Sc recovery of more than 90%. Carbonate leaching using Na2CO3, NaHCO3 or propylene carbonate was shown to be ineffective. Sulphuric acid leaching results in the release of large amounts of silicic acid in the leaching system, and hyper-crosslinked polystyrene-based sorbents were used to separate the silicic acid from the sulphate leach solution. The sorption properties of Purolite MN-202 containing no ionic groups and Purolite MN-102 containing tertiary amine groups were investigated. The results showed that weakly basic anion exchangers based on hyper-crosslinked polystyrene are preferred for cleaning the slag leaching solutions.

The Rammelsberg syn-sedimentary-exhalative ore deposit in Germany was exploited for 1 800 years before the mine closed in 1988. The main minerals of economic interest were Cu, Pb and Zn sulphide, and the ore also contained small amounts of scarce metals such as Co, In, Ga and Ag and environmentally relevant elements such as As, Cd and Th. During the last 70 years of operations, flotation was applied to the finely disseminated and inter-grown minerals, producing about 7 000 000 t of tailings which were directed to the Bollrich tailings pond. A research project was initiated to characterise the tailings with the aim of developing an effective reprocessing route. Drill samples were analysed for total metal content, metal correlations, incorporation, alteration and intergrowth. The results showed that the material is a fine-grained silty to clayey mixture of pyrite, chalcopyrite, sphalerite, galena and barite in a matrix mainly consisting of mixed carbonate, chlorite and illite. The common sulphide compound is pyrite, and most of the metals of commercial interest, Cu, Zn, Pb, Co, In and Ag, are concentrated within the sulphide fraction. Co is primarily related to pyrite, followed by chalcopyrite and sphalerite, In to sphalerite and chalcopyrite, and Ag is solely found in galena. If a mixed non-ferrous metal concentrate is produced, it will contain noticeable amounts of As, Cd and Tl. If the non-ferrous metals and trace elements are concentrated in single fractions, As will be related to Cu as chalcopyrite and Cd to Zn as sphalerite. Apart from minute amounts of Ba silicate, Ba is present as sulphate. Ga is enriched in iron oxide, primary magnetite and chlorite, and is not readily accessible for processing. The sediment in the settling pond has no significant layered structure and shows no alteration within the drilling profiles. Depth profiles show an increase in the barite concentration and a slight decrease in the four sulphide compounds with increasing depth. The mat, The Rammelsberg syn-sedimentary-exhalative ore deposit in Germany was exploited for 1 800 years before the mine closed in 1988. The main minerals of economic interest were Cu, Pb and Zn sulphide, and the ore also contained small amounts of scarce metals such as Co, In, Ga and Ag and environmentally relevant elements such as As, Cd and Th. During the last 70 years of operations, flotation was applied to the finely disseminated and inter-grown minerals, producing about 7 000 000 t of tailings which were directed to the Bollrich tailings pond. A research project was initiated to characterise the tailings with the aim of developing an effective reprocessing route. Drill samples were analysed for total metal content, metal correlations, incorporation, alteration and intergrowth. The results showed that the material is a fine-grained silty to clayey mixture of pyrite, chalcopyrite, sphalerite, galena and barite in a matrix mainly consisting of mixed carbonate, chlorite and illite. The common sulphide compound is pyrite, and most of the metals of commercial interest, Cu, Zn, Pb, Co, In and Ag, are concentrated within the sulphide fraction. Co is primarily related to pyrite, followed by chalcopyrite and sphalerite, In to sphalerite and chalcopyrite, and Ag is solely found in galena. If a mixed non-ferrous metal concentrate is produced, it will contain noticeable amounts of As, Cd and Tl. If the non-ferrous metals and trace elements are concentrated in single fractions, As will be related to Cu as chalcopyrite and Cd to Zn as sphalerite. Apart from minute amounts of Ba silicate, Ba is present as sulphate. Ga is enriched in iron oxide, primary magnetite and chlorite, and is not readily accessible for processing. The sediment in the settling pond has no significant layered structure and shows no alteration within the drilling profiles. Depth profiles show an increase in the barite concentration and a slight decrease in the four sulphide compounds with increasing depth. The mat

Research undertaken at the Pyrometallurgy Innovation Centre (PYROSEARCH) at the University of Queensland, Australia, involves integrated experimental and thermodynamic modelling studies of complex multi-component, multi-phase systems. High temperature studies have been conducted on Cu-containing systems involving gas/slag/matte/metal/speiss/solids oxide and sulphide systems, with the major elements Cu-Pb-Fe-S-O-Si, the addition of slagging elements Al, Ca and Mg and selected minor elements including, Zn, As, Sn, Sb, Bi, Ag and Au. The studies provide experimental data and sophisticated thermodynamic databases and predictive modelling to achieve improvements in pyrometallurgical processes. Equilibration, quenching and microanalysis experimental approaches are being continuously improved and have enabled increasingly more complex systems to be characterised. An example is presented of a simplified one-step reactor thermodynamic model of a copper flash smelting furnace. The model is used to illustrate the differences between the predicted process outcomes using simplified process control predictions and the latest thermodynamic databases. The model was used to predict the effects of Al2O3, CaO and MgO impurities and of the matte grade on the liquidus temperatures of the slag. The model was also employed to evaluate the potential for adjustment of the Fe/SiO2 ratio in the slag. The calculations showed that changing the ratio from 1.30 to 1.40 enables an additional 2 t/h of precipitated Cu product as a novel feed material to be added to the feed, resulting in an increase of 9 900 tpa of Cu through the smelter. Alternatively, an additional 2 t/h concentrate to the feed and an additional annual production of 4 400 tpa of Cu for the same percent Cu in the matte can be achieved., Research undertaken at the Pyrometallurgy Innovation Centre (PYROSEARCH) at the University of Queensland, Australia, involves integrated experimental and thermodynamic modelling studies of complex multi-component, multi-phase systems. High temperature studies have been conducted on Cu-containing systems involving gas/slag/matte/metal/speiss/solids oxide and sulphide systems, with the major elements Cu-Pb-Fe-S-O-Si, the addition of slagging elements Al, Ca and Mg and selected minor elements including, Zn, As, Sn, Sb, Bi, Ag and Au. The studies provide experimental data and sophisticated thermodynamic databases and predictive modelling to achieve improvements in pyrometallurgical processes. Equilibration, quenching and microanalysis experimental approaches are being continuously improved and have enabled increasingly more complex systems to be characterised. An example is presented of a simplified one-step reactor thermodynamic model of a copper flash smelting furnace. The model is used to illustrate the differences between the predicted process outcomes using simplified process control predictions and the latest thermodynamic databases. The model was used to predict the effects of Al2O3, CaO and MgO impurities and of the matte grade on the liquidus temperatures of the slag. The model was also employed to evaluate the potential for adjustment of the Fe/SiO2 ratio in the slag. The calculations showed that changing the ratio from 1.30 to 1.40 enables an additional 2 t/h of precipitated Cu product as a novel feed material to be added to the feed, resulting in an increase of 9 900 tpa of Cu through the smelter. Alternatively, an additional 2 t/h concentrate to the feed and an additional annual production of 4 400 tpa of Cu for the same percent Cu in the matte can be achieved.

Hydrogen presents itself as a variable alternative to the conventional energy resources, as the combustion of hydrogen releases large amounts of energy and the by-product is water vapour, a harmless gas. Aluminium dross is a waste product of aluminium melting/remelting units. The metal-water reaction between metallic content of aluminium dross and water is studied in this paper. NaOH was added to make the water a little alkaline, enhancing the rate of reaction. The presence of Al2O3 over metallic Al causes the reduction of overall hydrogen generation. Also various methods are discussed to enhance the production of hydrogen gas in order to design an economic and sustainable process to treat aluminium dross., Hydrogen presents itself as a variable alternative to the conventional energy resources, as the combustion of hydrogen releases large amounts of energy and the by-product is water vapour, a harmless gas. Aluminium dross is a waste product of aluminium melting/remelting units. The metal-water reaction between metallic content of aluminium dross and water is studied in this paper. NaOH was added to make the water a little alkaline, enhancing the rate of reaction. The presence of Al2O3 over metallic Al causes the reduction of overall hydrogen generation. Also various methods are discussed to enhance the production of hydrogen gas in order to design an economic and sustainable process to treat aluminium dross.

The pyrometallurgical Waelz process is the most often applied technology for recycling of dusty steel mill residues globally, especially for EAF dust. A significant amount of the globally generated dust is lost by sending it to landfill, which means an annual loss of more than 500 000 t of Zn. Global Steel Dust is engaged in the development, installation and operation of Waelz plants for treatment of EAF dust, especially in developing countries. Its first new plant has been commissioned in May 2016 in the Kingdom of Saudi Arabia. This paper explains the state-of-the-art plant layout of this greenfield plant, especially the raw material preparation with intensive mixer and the measures to maintain extremely low diffuse dust emission and very low atmospheric emission. The Waelz process as it is applied by GSD is listed in the BREF-notes of the European Community and combines it with the US Best Demonstrated Available Technology. The main focus is on the experiences and challenges during the engineering and construction until the start-up of the plant under the specific conditions of KSA., The pyrometallurgical Waelz process is the most often applied technology for recycling of dusty steel mill residues globally, especially for EAF dust. A significant amount of the globally generated dust is lost by sending it to landfill, which means an annual loss of more than 500 000 t of Zn. Global Steel Dust is engaged in the development, installation and operation of Waelz plants for treatment of EAF dust, especially in developing countries. Its first new plant has been commissioned in May 2016 in the Kingdom of Saudi Arabia. This paper explains the state-of-the-art plant layout of this greenfield plant, especially the raw material preparation with intensive mixer and the measures to maintain extremely low diffuse dust emission and very low atmospheric emission. The Waelz process as it is applied by GSD is listed in the BREF-notes of the European Community and combines it with the US Best Demonstrated Available Technology. The main focus is on the experiences and challenges during the engineering and construction until the start-up of the plant under the specific conditions of KSA.

The Waelz plant for the recycling of steel mill, especially EAF, dust is the first plant of that kind in the Middle East/North Africa (MENA) region. This paper starts with a report of the commissioning of the plant in May/June 2016. It also shows the challenges to find and train suitable operating personnel under the specific conditions of the Kingdom of Saudi Arabia. The results and experiences of the first year of operation are presented: comparing the achieved vs. planned capacity, evaluating the quality of the products Waelz zinc oxide and Waelz iron product, explaining the way and results of controlling the diffuse dust and atmospheric emission, summarising the differences in the raw materials., The Waelz plant for the recycling of steel mill, especially EAF, dust is the first plant of that kind in the Middle East/North Africa (MENA) region. This paper starts with a report of the commissioning of the plant in May/June 2016. It also shows the challenges to find and train suitable operating personnel under the specific conditions of the Kingdom of Saudi Arabia. The results and experiences of the first year of operation are presented: comparing the achieved vs. planned capacity, evaluating the quality of the products Waelz zinc oxide and Waelz iron product, explaining the way and results of controlling the diffuse dust and atmospheric emission, summarising the differences in the raw materials.

In production of stainless steel 10-70 kg of dust per ton of steel is produced. The dusts are usually treated by pyrometallurgical methods such as plasma processing or Waelz kiln. Hydrometallurgical methods have been developed especially for the recovery of zinc from Electric Arc Furnace (EAF) dusts, focussing on unalloyed steel EAF dust. A new concept has been developed to selectively leach zinc from stainless steel bag house dusts enabling recycling the remaining iron, nickel and chromium compounds back to melting. Zinc can be dissolved using either strong alkaline or strong acid solutions. Factors affecting the leaching efficiency have been determined. Mineralogical factors limiting the zinc dissolution have been determined and pretreatment methods have been developed. Using the optimal conditions the zinc level in the treated dust has been low enough allowing recycling of the dust to melting. (Authors.), In production of stainless steel 10-70 kg of dust per ton of steel is produced. The dusts are usually treated by pyrometallurgical methods such as plasma processing or Waelz kiln. Hydrometallurgical methods have been developed especially for the recovery of zinc from Electric Arc Furnace (EAF) dusts, focussing on unalloyed steel EAF dust. A new concept has been developed to selectively leach zinc from stainless steel bag house dusts enabling recycling the remaining iron, nickel and chromium compounds back to melting. Zinc can be dissolved using either strong alkaline or strong acid solutions. Factors affecting the leaching efficiency have been determined. Mineralogical factors limiting the zinc dissolution have been determined and pretreatment methods have been developed. Using the optimal conditions the zinc level in the treated dust has been low enough allowing recycling of the dust to melting. (Authors.)

The increase of raw material base of the PJSC Chelyabinsk zinc plant (CZP) is realised by means of processing increasing volumes of Zn-containing residues of Russia's metallurgical production. The last 10 years showed the following tendencies: it is necessary to correct separate design decisions, including the charge pelletising unit, recycling material collection, gas cooling and Waelz oxides collection; and it is necessary to improve the Waelz process by means of steam usage during blowing, increase of the furnace mouth area, lime feed in the discharge end, and a chlorinator. PJSC CZP has developed a project and is busy implementing the treatment of oxidised zinc-containing by-products in large-size tube kilns. The material is subjected to Waelz processing in a kiln. The off gas heat is used for steam production in a boiler. Waelz oxide is subjected to calcinations in another kiln. Calcined Waelz oxide is leached with subsequent production of zinc-containing solutions used for zinc and indium recovery. To reduce indium losses in slag, a chlorinator is added at the discharge end of the Waelz kiln., The increase of raw material base of the PJSC Chelyabinsk zinc plant (CZP) is realised by means of processing increasing volumes of Zn-containing residues of Russia's metallurgical production. The last 10 years showed the following tendencies: it is necessary to correct separate design decisions, including the charge pelletising unit, recycling material collection, gas cooling and Waelz oxides collection; and it is necessary to improve the Waelz process by means of steam usage during blowing, increase of the furnace mouth area, lime feed in the discharge end, and a chlorinator. PJSC CZP has developed a project and is busy implementing the treatment of oxidised zinc-containing by-products in large-size tube kilns. The material is subjected to Waelz processing in a kiln. The off gas heat is used for steam production in a boiler. Waelz oxide is subjected to calcinations in another kiln. Calcined Waelz oxide is leached with subsequent production of zinc-containing solutions used for zinc and indium recovery. To reduce indium losses in slag, a chlorinator is added at the discharge end of the Waelz kiln.

A novel aluminothermic smelting reduction (ASR) process was investigated for cleaning ferrous and nonferrous slags; this process is not only able to recover valuable iron, but also eliminates hazardous elements from the end-of-life slag products. The effect of adding Al to the reduction of iron oxide and subsequent iron recovery from waste steel slag and copper slag at 1 773 K was investigated using an induction furnace. The content of FeO and Al2O3 in the molten slag varied dramatically over a 5 min. period, followed by nearly constant values. Iron recovery exhibited a maximum value at a specific Al/FeO molar ratio. To determine the iron recovery, a novel triangular material balance diagram, which represents the balance among the reduced iron ingot, the reduced iron droplets dispersed in the slag phase and the residual (unreduced) iron, was proposed. Solid compounds, such as spinel and olivine, were precipitated in the slag during the ASR process. This was confirmed by both XRD analysis and a thermochemical computation method. Finally, the elimination rate of hazardous elements, such as As, Bi, Pb, and Sb, from end-of-life copper slag by the ASR process was determined. (Authors.), A novel aluminothermic smelting reduction (ASR) process was investigated for cleaning ferrous and nonferrous slags; this process is not only able to recover valuable iron, but also eliminates hazardous elements from the end-of-life slag products. The effect of adding Al to the reduction of iron oxide and subsequent iron recovery from waste steel slag and copper slag at 1 773 K was investigated using an induction furnace. The content of FeO and Al2O3 in the molten slag varied dramatically over a 5 min. period, followed by nearly constant values. Iron recovery exhibited a maximum value at a specific Al/FeO molar ratio. To determine the iron recovery, a novel triangular material balance diagram, which represents the balance among the reduced iron ingot, the reduced iron droplets dispersed in the slag phase and the residual (unreduced) iron, was proposed. Solid compounds, such as spinel and olivine, were precipitated in the slag during the ASR process. This was confirmed by both XRD analysis and a thermochemical computation method. Finally, the elimination rate of hazardous elements, such as As, Bi, Pb, and Sb, from end-of-life copper slag by the ASR process was determined. (Authors.)

Metallo specialises in the smelting and refining of complex non-ferrous scrap. Many different raw materials enter the process to be refined to pure copper, tin and lead. Nickel, zinc and precious metals are partially valorised. In this paper, the Metallo process is presented with particular emphasis on the complexity of incoming raw materials. Metallo's newest investment, the Slag Treatment Unit (STU), targets better valorisation of zinc and the production of a clean mineral product called Koranel. The core of the STU is a fuming reactor, using submerged plasma technology. The STU is designed to treat fayalitic copper slag and recover metals (copper bullion) and oxides (zinc and lead dust). The most important product is Koranel, a mineral product with very low levels of metal and excellent environmental performance. Some potential applications for Koranel are discussed. (Authors.), Metallo specialises in the smelting and refining of complex non-ferrous scrap. Many different raw materials enter the process to be refined to pure copper, tin and lead. Nickel, zinc and precious metals are partially valorised. In this paper, the Metallo process is presented with particular emphasis on the complexity of incoming raw materials. Metallo's newest investment, the Slag Treatment Unit (STU), targets better valorisation of zinc and the production of a clean mineral product called Koranel. The core of the STU is a fuming reactor, using submerged plasma technology. The STU is designed to treat fayalitic copper slag and recover metals (copper bullion) and oxides (zinc and lead dust). The most important product is Koranel, a mineral product with very low levels of metal and excellent environmental performance. Some potential applications for Koranel are discussed. (Authors.)

The RLE (roasting, leaching and electrolysis) zinc production process used by Votorantim Metals Holding smelter at Juiz de Fora (VMH JF) was the most common employed at Zn companies worldwide until recently. At VMH JF, there are two waste dams in the unit, the Fish Dam that is already out of operation, and the Stones Dam that is still in operation. The Stones Dam constantly receives waste from zinc hydrometallurgy and the iron residue (jarosite) corresponding to almost 85% of that amount. This material has significant amounts of metals such as Zn, Pb, Ag and In and their recoveries combined with the possibility of producing an inert waste (slag) and the end of the Zn industry need for dams could make it economically viable to install a plant for this purpose. This study assesses in pilot scale the efficiency of the pyrometallurgical route on a TSL furnace to process the material contained in the waste dams of the plant, and if this process could be a technical and economical solution to finish or reduce significantly the waste generated in the Votorantim zinc smelters., The RLE (roasting, leaching and electrolysis) zinc production process used by Votorantim Metals Holding smelter at Juiz de Fora (VMH JF) was the most common employed at Zn companies worldwide until recently. At VMH JF, there are two waste dams in the unit, the Fish Dam that is already out of operation, and the Stones Dam that is still in operation. The Stones Dam constantly receives waste from zinc hydrometallurgy and the iron residue (jarosite) corresponding to almost 85% of that amount. This material has significant amounts of metals such as Zn, Pb, Ag and In and their recoveries combined with the possibility of producing an inert waste (slag) and the end of the Zn industry need for dams could make it economically viable to install a plant for this purpose. This study assesses in pilot scale the efficiency of the pyrometallurgical route on a TSL furnace to process the material contained in the waste dams of the plant, and if this process could be a technical and economical solution to finish or reduce significantly the waste generated in the Votorantim zinc smelters.

The renewable resources industry is currently the fastest developing one of the emerging industries of the whole world. In China, the huge market is waiting for development, since the renewable resources industry has just been established. The lead recycling industry has developed rapidly along with a substantial increase in lead demand and shortage of lead ore resources. Lead waste has been fully utilised to reduce lead pollution and waste of resources, which makes lead into a virtuous cycle of 'production, consumption and recycling'. There are several pyrometallurgical processes for lead recycling which usually cause a variety of pollutants, such as lead dust, SO2, NO2, dioxin, etc. Flue gas containing SO2 must be treated in order to meet the emissions standard and the recovery of sulphur resources. Several processes can be used to treat the flue gas according to scale of secondary lead smelting, gas volume and SO2 concentration, including wet gas sulphuric acid process (H2SO4), sodium alkaline process (anhydrous sodium sulphite) and solvent process (liquid SO2). The principal, process characteristics and by-product utilisation of each process is introduced. Summaries and analysis of each process are made from several aspects, such as investment estimation, techno-economic indexes, operation and maintenance. Conclusions are made according to the different smelting conditions and finally, the suitable gas treatment is advised., The renewable resources industry is currently the fastest developing one of the emerging industries of the whole world. In China, the huge market is waiting for development, since the renewable resources industry has just been established. The lead recycling industry has developed rapidly along with a substantial increase in lead demand and shortage of lead ore resources. Lead waste has been fully utilised to reduce lead pollution and waste of resources, which makes lead into a virtuous cycle of 'production, consumption and recycling'. There are several pyrometallurgical processes for lead recycling which usually cause a variety of pollutants, such as lead dust, SO2, NO2, dioxin, etc. Flue gas containing SO2 must be treated in order to meet the emissions standard and the recovery of sulphur resources. Several processes can be used to treat the flue gas according to scale of secondary lead smelting, gas volume and SO2 concentration, including wet gas sulphuric acid process (H2SO4), sodium alkaline process (anhydrous sodium sulphite) and solvent process (liquid SO2). The principal, process characteristics and by-product utilisation of each process is introduced. Summaries and analysis of each process are made from several aspects, such as investment estimation, techno-economic indexes, operation and maintenance. Conclusions are made according to the different smelting conditions and finally, the suitable gas treatment is advised.

In 2015 the German Resource Institute (GERRI) was founded as a virtual institute and a national network by five leading German research institutions of the raw materials sector. The founding partners of GERRI offer broad knowledge and experience in the following fields of expertise: exploration and mining, processing, hydrometallurgy, pyrometallurgy, materials, waste management, machinery, modelling and special analytics. The main goal of GERRI is to strengthen transdisciplinary research along the entire value chain of metalliferous minerals and raw materials. GERRI's unique feature is a competence mapping, which bundles the available knowledge and experience within all competence fields of this network. GERRI is embedded in the 'Research for Sustainable Development (FONA)' programme of the German Federal Ministry of Education and Research (BMBF). After expiry of the five-year funding period, GERRI is envisioned to persist independently as a network and will be financed, among others, by industrial and research funds., In 2015 the German Resource Institute (GERRI) was founded as a virtual institute and a national network by five leading German research institutions of the raw materials sector. The founding partners of GERRI offer broad knowledge and experience in the following fields of expertise: exploration and mining, processing, hydrometallurgy, pyrometallurgy, materials, waste management, machinery, modelling and special analytics. The main goal of GERRI is to strengthen transdisciplinary research along the entire value chain of metalliferous minerals and raw materials. GERRI's unique feature is a competence mapping, which bundles the available knowledge and experience within all competence fields of this network. GERRI is embedded in the 'Research for Sustainable Development (FONA)' programme of the German Federal Ministry of Education and Research (BMBF). After expiry of the five-year funding period, GERRI is envisioned to persist independently as a network and will be financed, among others, by industrial and research funds.

Meeting the increasing global nonferrous metals demand has been sustained by increased attention to recycling and lean occurrences. In this work, detailed characterisation to identify valuable mineralisation and first investigation toward recommending an extraction route were carried out on a rock formation in a part of the Anka schist belt (SW Nigeria). Using XRD, optical microscopy, SEM-EDS, image analysis, and wet assaying, copper was found in the ores as chalcocite (Cu2S), covellite (CuS), cuprite (Cu2O) and malachite (CuCO3.Cu(OH)2), co-occurring with haematite and barite within a feldspartic quartzite host rock. This mineralogy presents a selective leaching task, to leach the copper sulphides and oxides from other oxides, which will seek to maximise copper recovery with the least iron recovery. Acid and alkali oxidative leaching were investigated using three different lixiviants. Alkali oxidative leaching using NH3 solution with H2O2 rejected the iron most and gave the cleanest solution, but ammonium persulphate extracted the copper most, at copper leach solution assay of 942 mg/l. The lixiviants behaviour was linked to a specific mineralogical feature of the occurrence. Other probable alternatives for investigations in exploiting this occurrence were mentioned., Meeting the increasing global nonferrous metals demand has been sustained by increased attention to recycling and lean occurrences. In this work, detailed characterisation to identify valuable mineralisation and first investigation toward recommending an extraction route were carried out on a rock formation in a part of the Anka schist belt (SW Nigeria). Using XRD, optical microscopy, SEM-EDS, image analysis, and wet assaying, copper was found in the ores as chalcocite (Cu2S), covellite (CuS), cuprite (Cu2O) and malachite (CuCO3.Cu(OH)2), co-occurring with haematite and barite within a feldspartic quartzite host rock. This mineralogy presents a selective leaching task, to leach the copper sulphides and oxides from other oxides, which will seek to maximise copper recovery with the least iron recovery. Acid and alkali oxidative leaching were investigated using three different lixiviants. Alkali oxidative leaching using NH3 solution with H2O2 rejected the iron most and gave the cleanest solution, but ammonium persulphate extracted the copper most, at copper leach solution assay of 942 mg/l. The lixiviants behaviour was linked to a specific mineralogical feature of the occurrence. Other probable alternatives for investigations in exploiting this occurrence were mentioned.

Manganese is frequently found as a naturally associated element with lead-zinc ores. During the processing of Zn concentrates with high Mn content, the selective sulphate leaching of Zn from concentrate is not possible and eventually ends up with extraction of Mn along with other minor impurities in to the zinc leachate. The Mn and Mg levels in zinc hydrorefineries of Hindustan Zinc have increased above acceptable limits. Due to limited capacity of continuously operated Mg Removal (MR) bleeding section, alternative routes for reducing Mn in zinc sulphate solution have been explored. The focus is mainly on using selective oxidants to oxidise and precipitate Mn from Zn leach solution. KMnO4 has been selectively preferred for Mn reduction by taking account of reagent consumption as well as operational lenience. Consumption of MnO2 for oxidising ferrous to ferric ions in the leaching circuit was identified as another source of Mn increase. Sodium persulphate (SPS), a powerful oxidising agent, was explored to replace MnO2 slurry in leaching; SPS has additional advantages in oxidising ferrous to ferric ions along with sodium sulphate production, a prime reagent used for iron precipitation in the form of jarosite., Manganese is frequently found as a naturally associated element with lead-zinc ores. During the processing of Zn concentrates with high Mn content, the selective sulphate leaching of Zn from concentrate is not possible and eventually ends up with extraction of Mn along with other minor impurities in to the zinc leachate. The Mn and Mg levels in zinc hydrorefineries of Hindustan Zinc have increased above acceptable limits. Due to limited capacity of continuously operated Mg Removal (MR) bleeding section, alternative routes for reducing Mn in zinc sulphate solution have been explored. The focus is mainly on using selective oxidants to oxidise and precipitate Mn from Zn leach solution. KMnO4 has been selectively preferred for Mn reduction by taking account of reagent consumption as well as operational lenience. Consumption of MnO2 for oxidising ferrous to ferric ions in the leaching circuit was identified as another source of Mn increase. Sodium persulphate (SPS), a powerful oxidising agent, was explored to replace MnO2 slurry in leaching; SPS has additional advantages in oxidising ferrous to ferric ions along with sodium sulphate production, a prime reagent used for iron precipitation in the form of jarosite.

This study experimentally investigates the distribution behaviour of indium, gallium and tin between metallic copper and FeOx-SiO2-Al2O3 slag. Aluminium is a general impurity metal in copper scrap and its presence will modify the slag properties and thus affect the behaviour of the minor elements in the process. To investigate the influence of Al in WEEE (waste electric and electronic equipment) smelting, the experiments were executed in alumina crucibles. The experimental conditions were chosen to simulate both reducing and oxidising process conditions of black copper smelting within the oxygen potential range of 10-5 - 10-10 atm at 1 300 degrees C. The employed experimental technique was equilibration-quenching followed by electroprobe microanalysis (EPMA). The distribution coefficient of tin between copper and slag increased from 0.1 to 100 as a function of decreasing oxygen partial pressure, and indium from 0.5 to 50. Gallium was mainly distributed into aluminous spinel with distribution coefficient of 2-3. Gallium concentrations in copper were mainly below the detection limit of EPMA and thus distribution coefficients between copper and slag were defined reliably only in the lower oxygen partial pressure range (10-9 - 10-10) as 0.1-0.5., This study experimentally investigates the distribution behaviour of indium, gallium and tin between metallic copper and FeOx-SiO2-Al2O3 slag. Aluminium is a general impurity metal in copper scrap and its presence will modify the slag properties and thus affect the behaviour of the minor elements in the process. To investigate the influence of Al in WEEE (waste electric and electronic equipment) smelting, the experiments were executed in alumina crucibles. The experimental conditions were chosen to simulate both reducing and oxidising process conditions of black copper smelting within the oxygen potential range of 10-5 - 10-10 atm at 1 300 degrees C. The employed experimental technique was equilibration-quenching followed by electroprobe microanalysis (EPMA). The distribution coefficient of tin between copper and slag increased from 0.1 to 100 as a function of decreasing oxygen partial pressure, and indium from 0.5 to 50. Gallium was mainly distributed into aluminous spinel with distribution coefficient of 2-3. Gallium concentrations in copper were mainly below the detection limit of EPMA and thus distribution coefficients between copper and slag were defined reliably only in the lower oxygen partial pressure range (10-9 - 10-10) as 0.1-0.5.

This article describes the strategy of Nyrstar's Metal Processing Segment towards knowledge development across the different smelters at different sides of the world. Two facts were contributing to this necessity: the global transformation of the Nyrstar smelters and the fact that the two smelters were operating almost exclusively on the MMG's Century zinc concentrate, with the latter to disappear end 2015. For these two smelters, projects had been implemented in 2015 enabling the treatment of concentrates containing increased cadmium and iron volumes, following the introduction of a more complex feedbook after the closure of the Century mine. To prepare for the end of the Century concentrate, this approach of knowledge sharing was used to develop extensive test work and process modifications at the Nyrstar Hobart (Australia) and Nyrstar Budel (The Netherlands) zinc smelters. The development of a technical network group for learning, research, testing and linked industrial test campaigns has proven to be of great value for developing and understanding the control mechanisms of fluidised bed roasting, especially separating and linking the physical/chemical properties of the roaster feed to the physical and chemical aspects of roasting process control., This article describes the strategy of Nyrstar's Metal Processing Segment towards knowledge development across the different smelters at different sides of the world. Two facts were contributing to this necessity: the global transformation of the Nyrstar smelters and the fact that the two smelters were operating almost exclusively on the MMG's Century zinc concentrate, with the latter to disappear end 2015. For these two smelters, projects had been implemented in 2015 enabling the treatment of concentrates containing increased cadmium and iron volumes, following the introduction of a more complex feedbook after the closure of the Century mine. To prepare for the end of the Century concentrate, this approach of knowledge sharing was used to develop extensive test work and process modifications at the Nyrstar Hobart (Australia) and Nyrstar Budel (The Netherlands) zinc smelters. The development of a technical network group for learning, research, testing and linked industrial test campaigns has proven to be of great value for developing and understanding the control mechanisms of fluidised bed roasting, especially separating and linking the physical/chemical properties of the roaster feed to the physical and chemical aspects of roasting process control.

The extraction of precious metals as pure metal or alloys from primary raw materials and recycling materials is an important issue for ecological perspective. In traditional fire assay, the precious metals are separated and accumulated by using lead flux. Nickel sulphide and copper flux are also used. The combination of fire assay and modern analytical techniques, such as ICP-OES, spark-OES, XRF and Glow Discharge Mass Spectrometry (GD-MS), has proved the standard process in industry for obtaining gold, silver and platinum group metals. The efficiency of recovery of precious metals (PM) is nearly 100%. This combination is worthwhile to recycle different kinds of recycling materials such as e-scraps, automotive and chemical catalysts, production waste from the electronic and galvanic industries as well as for the analysis of PM from primary raw materials (ores, concentrates). It is essential to analyse the content of precious metals for selling, buying or recycling in order to ensure the quality and value of the materials. Experience with a good background of understanding analysis in the company IMG produces accurate and precise assay results., The extraction of precious metals as pure metal or alloys from primary raw materials and recycling materials is an important issue for ecological perspective. In traditional fire assay, the precious metals are separated and accumulated by using lead flux. Nickel sulphide and copper flux are also used. The combination of fire assay and modern analytical techniques, such as ICP-OES, spark-OES, XRF and Glow Discharge Mass Spectrometry (GD-MS), has proved the standard process in industry for obtaining gold, silver and platinum group metals. The efficiency of recovery of precious metals (PM) is nearly 100%. This combination is worthwhile to recycle different kinds of recycling materials such as e-scraps, automotive and chemical catalysts, production waste from the electronic and galvanic industries as well as for the analysis of PM from primary raw materials (ores, concentrates). It is essential to analyse the content of precious metals for selling, buying or recycling in order to ensure the quality and value of the materials. Experience with a good background of understanding analysis in the company IMG produces accurate and precise assay results.

The purity of Zn leaching solution is very critical for the electrowinning and stripping of Zn plate. Investigations were carried out by Beijing General Research Institute of Mining and Metallurgy, Yunnan Chihong company and Zijin Mining company. In general, Cu and Cd are removed with cementation of Zn powder until Cu and Cd content in solution is less than 0.2 and 0.18 mg/l. In the second stage, Ni and Co are removed by Zn powder and Sb salt, until their content in solution is less than 0.1 mg/l after 3 h at 85 degrees C. Copper slag came from the Cu/Cd residue after H2SO4 leaching to dissolve Cd. It contains 72% Cu and a little Zn, Cd and V. At 60 degrees C chloride is reduced from 614 to 51.4 mg/l. The efficiency is ca. 92%. In some Zn smelters tannin was used to precipitate Ge and some zinc oxide ore also used SX-EW to produce Zn metal. Organic compounds should be removed from the solution. The principal is that the wetting angle and surface tension is different between the aqueous and organic phase. It causes oil particles to agglomerate and float on the surface of the solution which can then be separated from the aqueous phase. Continuous pilot tests were done and the results indicated that the organic content can be reduced from 50 to 2 mg/l., The purity of Zn leaching solution is very critical for the electrowinning and stripping of Zn plate. Investigations were carried out by Beijing General Research Institute of Mining and Metallurgy, Yunnan Chihong company and Zijin Mining company. In general, Cu and Cd are removed with cementation of Zn powder until Cu and Cd content in solution is less than 0.2 and 0.18 mg/l. In the second stage, Ni and Co are removed by Zn powder and Sb salt, until their content in solution is less than 0.1 mg/l after 3 h at 85 degrees C. Copper slag came from the Cu/Cd residue after H2SO4 leaching to dissolve Cd. It contains 72% Cu and a little Zn, Cd and V. At 60 degrees C chloride is reduced from 614 to 51.4 mg/l. The efficiency is ca. 92%. In some Zn smelters tannin was used to precipitate Ge and some zinc oxide ore also used SX-EW to produce Zn metal. Organic compounds should be removed from the solution. The principal is that the wetting angle and surface tension is different between the aqueous and organic phase. It causes oil particles to agglomerate and float on the surface of the solution which can then be separated from the aqueous phase. Continuous pilot tests were done and the results indicated that the organic content can be reduced from 50 to 2 mg/l.

This paper shows the latest technical developments, some already commercialised, some still at development stage, performed by Tecnicas Reunidas to respond to different zinc refinery needs. These developments are linked to ZINCEX Solvent Extraction technology implemented in seven industrial plants since 1976 and which, to date, has enabled the production of Zn from complex raw materials (oxidic ores and Waelz oxides) at very low cost. The developments described in this paper are: design of a Zn refinery based on Roasting-Leaching-Solvent Extraction-Electrowinning, especially suitable for Zn concentrates with high content of impurities like Mn and Mg; industrial implementation of ECOLEAD technology for the recovery of Pb and Ag from a Zn refinery fed exclusively with secondary raw materials (Waelz oxides); significant investment cost reduction at same operating cost with the new generation of ZINCEX technology; recovery of rare earths and critical metals from ZINCEX refinery streams; and adaptation of the ZINCEX technology to small sized plants, especially suitable for metals recycling facilities and recovery from waste streams., This paper shows the latest technical developments, some already commercialised, some still at development stage, performed by Tecnicas Reunidas to respond to different zinc refinery needs. These developments are linked to ZINCEX Solvent Extraction technology implemented in seven industrial plants since 1976 and which, to date, has enabled the production of Zn from complex raw materials (oxidic ores and Waelz oxides) at very low cost. The developments described in this paper are: design of a Zn refinery based on Roasting-Leaching-Solvent Extraction-Electrowinning, especially suitable for Zn concentrates with high content of impurities like Mn and Mg; industrial implementation of ECOLEAD technology for the recovery of Pb and Ag from a Zn refinery fed exclusively with secondary raw materials (Waelz oxides); significant investment cost reduction at same operating cost with the new generation of ZINCEX technology; recovery of rare earths and critical metals from ZINCEX refinery streams; and adaptation of the ZINCEX technology to small sized plants, especially suitable for metals recycling facilities and recovery from waste streams.