Copper was one of the first metals ever extracted and used by mankind. It is used for its unique properties, like corrosion resistance, good workability, high thermal conductivity and attractive appearance. New mines are opened to maintain a supply of primary feedstock to copper smelters. These new deposits are in many instances found to have a more complex mineralogy with several minor elements. Besides treating primary material, copper smelters also show an increasing interest in treating secondary material, such as copper containing scrap from waste electric and electronic equipment, which also have a complex composition.Waste electric and electronic equipment are first disassembled and upgraded by mechanical processing, generating a product stream called e-scrap, that can be added directly to the smelting processes as cold material or melted in a separate furnace producing a metallic alloy (referred to as black copper) and a slag phase. The black copper can be refined in different ways, whereof one is by using it as a secondary feed material for input to Peirce-Smith converters. Consequently the load of minor elements to the converter can be expected to increase with an increased treatment of e-scrap.This increased complexity of the raw material can potentially lead to smelter plants having to deal with a feedstock containing several minor elements such as; antimony, bismuth, arsenic, gold, silver, etc. in levels that can influence the ability to, in a cost effective way, maintain the final grade of the copper cathode. Process simulations can be an important tool for understanding the impact of process parameters on the product quality and for the purpose of process optimisation. In the present work a dynamic, non-equilibrium model based on thermodynamics over the Peirce-Smith converter has been developed. The non-equilibrium conditions have been simulated by introducing individual but linked segments. The purpose of using segments was to consider different reaction zones which yield different conditions within the converter. The model was validated using plant data and showed good agreement for the major elements. The agreement between plant and calculated data for Pb, and Zn was not as good and more work is required regarding this aspect. The model was used to investigate the influence on the distribution of Bi and Sb during addition of black copper with or without slag. When black copper is added to a blow, the removal of Bi and Sb becomes lower compared to a blow without addition of black copper. Similar result is obtained during addition of black copper with slag. To maintain a total removal of Bi and Sb in similar levels as a blow without black copper, the black copper should be added as early as possible during the converting operation.

In this project the internal material transport processes of wet low-intensity magnetic separation (LIMS) is studied. The aim is to use results from experiments combined with published results to create a base for further research and development. During the initial work, presented here, the focus has been to develop an experimental platform and verify that ultrasound is a viable tool for flow measurement in suspensions with solids concentrations realistic for wet magnetic separation processes. The experiments have been carried out in a purpose built flow cell equipped with ultrasonic transducers and supported by a pump, mixer and necessary data accusation electronics. In the first conference paper, ultrasonic velocity profiling is used to estimate flow velocity profiles, and initial results are presented. When applied to flows of mineral suspensions of high solids concentration, similar to those in wet LIMS, the method is unique in combining:* Non-intrusive measurements.* Operation by a single transducer element.* Relatively good spatial resolution.* Operation in opaque suspensions.* Fast sampling rate.In a second conference paper the signal processing behind the measurement method is investigated more thoroughly and focus is on robust profile estimation. The connection between solids concentration, ultrasonic transducer centre frequency and penetration depth is also investigated. In a publication the measurements are put in relation to the application of interest. It is shown that the methods are generally applicable and can be used in situations where variations in suspension flow velocity through narrow geometries are of interest.The novelty shown is that it is possible to measure flow velocity profiles through suspensions carrying at least 10 vol% solids. For solids concentrations of 5 vol% or less it is possible to get a velocity profile through at least 50 mm of suspension. Results from measurements in suspensions of such high solids concentration combined with such long penetration depth have not been published before. The measurement method has gained attention from industry since it is generally applicable to narrow channel flows reachable from only one access point. / Wet LlMS - Measurements and models

The demands for more effective utilization of ore bodies and proper risk management in the mining industry have resulted in a new cross discipline called geometallurgy. Geometallurgy connects geological, mineral processing and subsequent downstream processing information together to provide a comprehensive model to be used in production planning and management. A geometallurgical program is the industrial application of geometallurgy. It provides a way to map the variation in the ore body, to handling the data and giving metallurgical forecast on spatial level.Three different approaches are used in geometallurgical programs. These include the traditional way, which uses chemical elements, the proxy method, which applies geometallurgical tests, and the mineralogical approach using mineralogy. The mineralogical approach provides the most comprehensive and versatile way to treat geometallurgical data. Therefore it was selected as a basis for this study. For the mineralogical method, quantitative mineralogical information is needed both on deposit and for the process. The geological model must describe the minerals present, give their chemical composition, report their mass proportions (modal composition) in the ore body and describe the texture. The process model must be capable of using mineralogical information provided by the geological model to forecast the metallurgical performance of different geological volumes (samples, ore blocks, geometallurgical domains or blends prepared for the plant) and periods (from minutes via hourly and daily scale to week, monthly and annual production). A literature survey showed that areas, where more development is needed for using the mineralogical approach, are: 1) quick and inexpensive techniques for reliable modal analysis of the ore samples; 2) textural classification of the ore capable to forecast the liberation distribution of the ore when crushed and ground; 3) unit operation models based on particle properties (at mineral liberation level) and 4) a system capable to handle all this information and transfer it to production model. This study focuses on solving the first and the third problem. A number of methods for obtaining mineral grades were evaluated with a focus on geometallurgical applicability, precision and trueness. The method survey included scanning electron microscopy based automated mineralogy, quantitative X-ray powder diffraction with Rietveld refinement, and element-to-mineral conversion. A new technique called combined method uses both quantitative X-ray diffraction with Rietveld refinement and the element-to-mineral conversion method. The method not only delivers the required turnover for geometallurgy, but also overcomes the shortcomings if X-ray powder diffraction or element-to-mineral conversion when used alone. Furthermore, various methods of obtaining modal mineralogy were compared and a model for evaluating precision and closeness of the methods was developed.Different levels of processing models can be classified in geometallurgy based on in which level the ore, i.e. the feed stream to the processing plant, is defined and what information subsequent streams carry. For mineral processing models the following five levels can be distinguished: particle size only level, elemental level, element by particle size level, mineral level, mineral by particle size level and mineral liberation (particle) level. The most comprehensive level of mineral processing models is the particle-based one which includes all necessary information for modeling unit operations. Within this study, as the first step, a unit operation model is built on particle level for wet low-intensity magnetic separation. The experimental data was gathered through a survey of the KA3 iron ore concentrator plant of Luossavaara-Kiirunavaara AB (LKAB) in Kiruna. The first wet magnetic separator of the process was used as the basis for the model development since the degree of liberation is important at this stage. Corresponding feed, concentrate and tailings streams were mass balanced on a mineral by size and liberation level. The mass balanced data showed that the behavior of individual particles in the magnetic separation is depending on their size and composition. The model, which has a size dependent by-pass parameter and a separation parameter dependent of the magnetic volume of the particle, is capable of forecasting the behavior of particles in magnetic separation. Modeling and simulation show the benefits that particle-based simulation provides compared to lower level process models which take into account only elemental or mineral grades.

Slag formation plays a decisive role in all steelmaking processes. In top-blowing Basic Oxygen Steelmaking (BOS), i.e. in the LD process, an emulsion consisting of liquid slag, dispersed metal droplets and solid particles will, together with process gases, form an expanding foam. Extensive research has defined the parameters that govern the foaming characteristics of BOS slag emulsions. It is a well known fact that certain process conditions will lead to an excessive foam growth, forcing the foam out through the vessel opening (mouth). This process event is commonly known as slopping. Slopping results in loss of valuable metal, equipment damage, lost production time and pollution. A literature survey covering the slopping phenomena has been carried out, as well as a deeper investigation into the causes behind slopping on the BOS vessels, type LD/LBE, at SSAB EMEA Metallurgy Luleå, equipped with an automatic system for slopping registration using image analysis. Good slag formation and foam-growth control in order to avoid slopping is primarily accomplished by taking preventive "static" measures. Improved slopping control has been achieved by developing a new oxygen lance control scheme, featuring adjustment of the distance between the lance tip and the metal bath according to scrap quality and ore additions. If "static" measures cannot be effectuated, in-blow control measures are needed. For such "dynamic" measures to be effective, it is necessary to have a system for slopping prediction. In the early-1980s a system for foam level and slopping control, based on BOS vessel vibration, was temporarily installed and tested on one of the vessels in Luleå. Based on the experiences from these tests it was decided to re-investigate the vessel vibration measurement technique. Trials on industrial scale BOS vessels of type LD/LBE have been carried out. FFT spectrum analysis has been applied in order to find the frequency band with best correlation to an estimated foam height. The results show that there is a correlation between vessel vibration and foam height which can be used for dynamic foam level and slopping control. The vessel vibration results have been tested against perhaps the most common implemented technique for dynamic foam height estimation and slopping control, the audiometric system. Parallel vibration and audio measurements have been carried out on a 130-tonne as well as on 300-tonne BOS vessels. The results show that during stable process conditions there is good agreement between the two methods with regard to foam height estimation and that combining the two methods will provide a powerful slopping prediction and control system.

The success of mineral flotation processes depends on the hydrophobization of the surface for the desired mineral particles whilst keeping, or making, all other minerals hydrophilic. This is achieved by adding several reagents to the flotation pulp to adsorb selectively at the mineral/water interface. The reagents (surface active agents) which selectively adsorbs on minerals to be floated are called collectors. It is realized that many of these commonly used reagents are highly toxic and often potential threats to the environment. Use of many of theseconventional chemicals will therefore be restricted soon and must eventually be stopped. An immediate effort is necessary to develop alternative eco-friendly reagents in order to continue to extract valuable minerals from ores. In addition, with the steady depletion of high grade, relatively easy to process ores, the mineral industry is confronted with a challenging task of finding more efficient techniques so as to exploit low grade, complex and disseminated typeof ore deposits and old tailing dumps. The development of new selective and environmentally acceptable substances containing almost tailor-made reagents is thus inevitable for sustainability. Several known chelating agents have been appropriately modified to make those behave as selective flotation collectors with some degree of success. The problem is that almost all of the chelating groups form complexes with almost all of the transition and many non-transition metals. As a consequence, absolute selectivity does not exist. Besides beingunsatisfactory from a scientific viewpoint, it assumes that the metal ion specificity observed for a functional group in bulk aqueous system would remain valid during surface chelation at the interface, while in actual practice, the specificity based on metal ion is neither valid nor useful where the cations participating in the complexation reactions are the same, for example separation among the calcium minerals.It is clear that a selective reagent should be based on the reagent interactions not merely with the metal ion on the surface, but with the whole surface. It is more appropriate to design reagents having functional groups so spaced that those are compatible with the relative positions of the metal ion sites available on the surface, that is, to design not just metalspecificbut structure-specific reagents. The understandings of molecular interactions involved in the recognition of surfaces by organic molecules in biomineralization process suggest the possibility of reagents specific to the crystal structure. These understandings have been successfully applied to the rational design and synthesis of molecules either for the control of crystal morphology or to inhibitcrystal growth processes through the recognition of specific crystal surfaces. The idea of molecules consisting of two groups having appropriate spacing between them to achieve structural compatibility during interaction with surface exhibit structure-specificity is of direct relevance to the reagents selectivity in flotation processes. The present investigation aims to develop and distinguish mineral specific reagents with two functional groups for use in flotation of calcium containing minerals. For this purpose, a series of dicarboxylate-based surfactants with varying spacing between the carboxylate groups (one, two or three methylene groups) were synthesized. As reference, a surfactant withthe same alkyl chain length but with only one carboxylate group in the polar part was synthesized. The adsorption behavior of these new reagents on pure apatite, calcite and fluorite mineral surfaces was studied using Hallimond tube flotation, ζ-potential and FTIR measurements. The relation between the adsorption behavior of a given surfactant on a specific mineral surface and its molecular structure over a range of concentration and pH values, as well as the region of maximum recovery were established. It was found that one of the reagents, with a specific distance between the carboxylate groups, was much moreselective for a particular mineral surface than the other homologues synthesized. This selective adsorption of a given surfactant to a particular mineral surface relative to other mineral surfaces as evidenced in flotation studies is substantiated by ζ-potential and infrared spectroscopy data. Our investigation revealed that it is possible to design and develop mineral specific reagents in flotation.

One of the measures of development and economy of a nation is its per capita consumption of steel, and the demand is fulfilled by iron ore. In the context of increasing environmental constraints and ores becoming leaner and leaner, recycling and utilization of ore fines becomes necessary. Pelletization, being one of the major agglomeration techniques is increasingly practiced across the world to produce agglomerates that can be fed into the metallurgical furnaces (say blast furnaces) for subsequent processing. In Europe, Sweden has the richest iron ore deposits, and mining and metals production contributes majorly to its net export. LKAB operates with magnetite ore bodies in the northern Sweden to produce magnetite pellets (26 MTPA) exports about 70 % of its product to the European Steel producers. Therefore, constant efforts are necessary to maintain and improve the quality of magnetite pellets, and it is necessary to enhance the understanding on the reaction kinetics and mechanisms responsible while producing pellets.Magnetite pellets prepared from the fines are indurated (heat hardened) to attain the quality standards in terms of strength and other metallurgical properties. The quality of magnetite pellet is primarily determined by the physico-chemical changes the pellet undergoes as it makes excursion through the gaseous and thermal environment in the induration furnace. Among these physico-chemical processes, the oxidation of magnetite phase and the sintering of oxidized magnetite (hematite) and magnetite (non-oxidized) phases are vital. Rates of these processes not only depend on the thermal and gaseous environment the pellet gets exposed in the induration reactor but are also interdependent on each other. Therefore, a doctorate project is undertaken to systematically understand these processes in isolation to the extent possible and quantify them seeking the physics. With this motivation, the current study is focused on investigating the sintering phenomena involved during induration of magnetite pellet.Experiments with single pellets were designed to understand and quantify the sintering behavior of oxidized magnetite (hematite) and magnetite independently. The kinetics of sintering can be described using power law (Ktn) and Arrhenius (ln⁡(TK^((1/n)) )=ln⁡K' - Q/RT ) equations. In the experiments, a single pellet was exposed to different thermal profiles in a controlled atmosphere, and their in-situ shrinkage was captured continuously by a novel technique using Optical Dilatometer. It was found that the sintering behavior captured by shrinkage of the pellet can be quantified using three isothermal kinetic parameters, namely – activation energy (Q), pre-exponential factor (K’) and time exponent (n). The values of activation energy and time exponent derived suggests that sintering of oxidized magnetite (hematite) is dominated by a single diffusion mechanism, whereas sintering of magnetite showed two distinct mechanisms; one operating at lower temperatures and the other at higher temperatures. The isothermal sintering kinetic equation is also extended to predict the non-isothermal sintering for both oxidized magnetite and magnetite, and validated with the laboratory experiments. This is further useful in predicting the sintering state of pellets during induration in the plant scale operations.

Removal of impurity elements in copper metallurgy is one of the major problems encountered today since pure copper ore reserves are becoming exhausted and the resources of unexploited ores often contain relatively high amounts of antimony, arsenic, mercury and bismuth, which need to be eliminated. The present work aims at pre-treating a tetrahedrite rich complex copper sulphide concentrate by selective dissolution of the impurities, therefore, upgrading it for pyrometallurgical processing. Characterisation of the complex concentrate was performed and the result shows that antimony and part of arsenic were present as tetrahedrite and bournonite. Dissolution kinetics of tetrahedrite in aqueous alkaline sodium sulphide solutions was investigated. It was found that the rate of dissolving tetrahedrite by the lixiviant increases with increase in reaction temperature, sodium sulphide concentration, sodium hydroxide concentration, and with decrease in mineral particle size. The kinetic study indicates that the rate of leaching tetrahedrite in the lixiviant under the selected conditions is chemically controlled through the particle surface reaction. The activation energies of the process were estimated as 81 kJ/mol and 75 kJ/mol, respectively, for antimony and arsenic dissolution from tetrahedrite. The estimated activation energies were within the range reported for a chemically controlled reaction process. Besides, the alkaline sulphide lixiviant proves selective and effective to dissolve these impurity elements from the concentrate with good recoveries. Further investigations on the factors influencing the leaching efficiency of the lixiviant were studied. Analysis of the leach residue indicates that copper content of the tetrahedrite has transformed into copper sulphides with the average chemical formula Cu1.64S. The grade and economic value of the concentrate were improved greatly after sulphide treatment, and therefore, suitable as a feedstock for smelting. The impurities in the concentrate were found to have reduced to a level satisfactory for smelting operation.Furthermore, modelling and optimisation of alkaline sulphide leaching of a complex copper concentrate containing 1.69% Sb and 0.14% Sn were conducted. Response surface methodology, in combination with central composite face-centred design (RSM-CCF), was used to optimise the operating parameters. The leaching temperature, sulphide ion concentration and solid concentration were chosen as the variables, and the response parameters were antimony and tin recoveries, and the time required to achieve 90% Sb dissolution. It was seen that the leaching process was strongly dependent on the reaction temperature as well as the sulphide ion concentration without any significant dependence on the solid concentration. Additionally, a mathematical model was constructed to characterise the leaching behaviour within the experimental range studied. The results from the model allow identification of the most favourable leaching conditions. The model was validated experimentally, and the results show that the model is reliable and accurate in predicting the leaching process.

In the present study, magnetite pellets with large amounts of the additives olivine, calcite and quartzite were isothermally reduced in a tubular furnace to study the interaction between iron oxides and the additives. A first attempt at using exaggerated amounts of additives was made in order to enable analyses of phases that do not otherwise occur in sufficient amounts for Xray diffraction and EDS-analyses. The reduction was thermodynamically set to yield either magnetite or wüstite at three different temperatures, 900, 1000 and 1150°C. For olivine, reduction tests were also performed at 500, 600, 700 and 800°C. The mineralogical phases that had formed were studied after oxidation as well as after reduction. The results showed that it was possible to identify, by X-ray diffraction, the main phases formed by the additives in all samples, after oxidation as well as reduction.The quartzite particles were shown to have remained quite intact after the oxidation treatment, except for small particles in the presence of impurities that formed melts. During reduction the quartzite particles reacted with iron so that fayalitic melts were formed already at 1000°C. After reduction at 1150°C all quartzite had transformed into a fayalitic melt so that most of the small pores had disappeared through sintering or had been filled by fayalite.In the sample with calcium oxide the additive particles had reacted during the oxidation treatment and formed calcium ferrites and calcium silicates. Upon reduction, the ferrites that formed during oxidation reduce, so that a porous calciowüstite becomes the primary phase already at 900°C. Calcium silicates that were formed during oxidation also remain in the sample as silicates during reduction.The results showed that the olivine after oxidation had reacted along the particle boundary and turned into magnesioferrite crystals and pyroxene/vitreous silica. Magnesium is liberated when the olivine particle breaks down, and finally ends up as islands of magnesioferrite surrounded by hematite in the original magnetite particles. In the pellet core the magnesium has diffused relatively long distances so that the magnesioferrite islands are not just found close to-, but also further away from the olivine particles. Upon reduction, the hematite converts to magnetite already at 500°C and in the tests carried out at 500-700°C, cracks were observed along the hematitemagnesioferrite boundary. At 800°C, temperature is enough to allow slow diffusion of magnesium from the magnesioferrite to the surrounding magnetite or wüstite, and at 900°C the cracks around the magnesioferrite phase disappear. The Mg stored in the wüstite then reacts with the silica slag in the sample when it approaches its melting point at 1000°C. The magnesium level in the wüstite then approaches a background level which was found to be about 2% after reduction for 2 hours at 1150°C.

Chromium is an important alloying element for stainless steels and other Cr-bearing steels. During the steelmaking process chromium is added to the steels mainly in the form of ferrochrome, which is largely produced by the energy-intensive smelting reduction process of chromite ore in the submerged arc furnace. To reduce the overall energy consumption during the ferrochrome production process and the chromium alloying process, direct chromiumalloying by chromite ore has been proposed. The application of this process will integrate the processes for ferrochrome production and chromium alloying, and thus has the potential to cut the production costs of the Cr-bearing steels by avoiding, or at least partially avoiding, the usage of ferrochrome. Further, this new alloying process has the capacity to improve therecovery of chromium from chromite ore. This thesis presents fundamental studies on the carbothermic reduction of synthetic iron chromite (FeCr2O4) and chromite ore, which aim at designing a direct alloying precursor to be applied in the industrial process. Thermogravimetric Analysis (TGA) experiments have been carried out to investigate the carbothermic reduction processes of FeCr2O4 in the absence/presence of metallic iron, and of chromite ore in the absence/presence of mill scale. In the case of using the mixture ‘FeCr2O4 +iron powder + graphite’, it is found that the presence of metallic iron enhances the reduction of FeCr2O4, and this enhancing effect increases with increasing iron addition. The enhancing effect of iron addition on the reduction of FeCr2O4 is due to the fact that the reduction of component Cr2O3 in FeCr2O4 is enhanced, and this effect is attributed to the presence of solidiron which can decrease the activity of chromium by having chromium in situ dissolved in the iron. In the case of using the mixture ‘chromite ore + petcoke’, it is found that the reduction of iron ions in the chromite ore starts before that of chromium ions in the ore and the reduction of iron ions and chromium ions in the ore overlaps to some degree. (Cr,Fe)7C3 is found to bethe intermediate phase during the reduction and a chromium gradient is found in the spinel phase of the fractional reduced sample at 1673 K. A four-stage reduction process is proposed: one stage involving the reduction of iron ions in the chromite ore and three stages involving the reduction of chromium ions in the ore. The activity aspects of component FeCr2O4 and component MgCr2O4 in the chromite ore have been considered. The difficulty in the reductionof the chromite ore is attributed to the fact that, as the reduction proceeds, the activity of component MgCr2O4 in the fractional reduced ore will decrease to a very low level, which makes the further reduction very difficult. In the case of using the mixture ‘chromite ore + mill scale + petcoke’, it is found that mill scale is reduced to iron before 1573 K. The asreduced iron is disseminated around chromite ore particles and, at the same time, some carbonis dissolved in the iron via diffusion. Reduction of chromite ore is enhanced with the addition of mill scale at temperatures higher than 1623 K, and the enhancing effect increases with increasing mill scale addition. The enhancing effect, in this case, is attributed to the presence of molten Fe-Cr-C phase in the vicinity of chromite ore, which can decrease the activity of chromium by having chromium in situ dissolved into the melt. Induction furnace experiments have been carried out to investigate the effectiveness of some different alloying mixtures. The experimental results have confirmed the necessity ofadjusting the composition of the slag to ensure high chromium yield in the final product and the experimental results show that, by using iron scrap, chromium yield can reach 90%. The present findings have led to the proposal of using ‘chromite ore + mill scale + petcoke’ as alloying precursor for direct chromium alloying. The effectiveness of this alloying precursor needs to be further explored by induction furnace experiments, followed by full scale Electric Arc Furnace experiments.

Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was investigated. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn,Fe)S), and galena (PbS), which are the most abundant sulphide minerals on Earth, generated H2O2 in pulp liquid during wet grinding in the presence or devoid of dissolved oxygen in water and also when the freshly ground solids are placed in water immediately after dry grinding. Pyrite generated more H2O2 than other sulphide minerals and the order of H2O2 production by the minerals found to be pyrite > chalcopyrite > sphalerite > galena. The pH of water influenced the extent of hydrogen peroxide formation where higher amounts of H2O2 are produced at highly acidic pH. The amount of H2O2 formed also increased with increasing sulphide mineral loading and grinding time due to increased surface area and its interaction with water. The sulphide surfaces are highly catalytically active due to surface defect sites and unsaturation because of broken bonds and capable of breaking down the water molecule leading to hydroxyl free radicals. Type of grinding medium on formation of hydrogen peroxide by pyrite revealed that the mild steel produced more H2O2 than stainless steel grinding medium, where Fe2+ and/or Fe3+ ions played a key role in producing higher amounts of H2O2.Furthermore, the effect of mixed sulphide minerals, i.e., pyrite–chalcopyrite, pyrite–galena, chalcopyrite–galena and sphalerite–pyrite, sphalerite–chalcopyrite and sphalerite–galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction in chalcopyrite–pyrite composition. In pyrite–sphalerite, chalcopyrite–sphalerite or galena–sphalerite mixed compositions, the increase in pyrite or chalcopyrite proportion, the concentration of H2O2 increased but with increase in galena proportion, the concentration of H2O2 decreased. Increasing pyrite proportion in pyrite–galena mixture, the concentration of H2O2 increased and also in the mixture of chalcopyrite–galena, the concentration of H2O2 increased with increasing chalcopyrite fraction. The results of H2O2 formation in pulp liquid of sulphide minerals and mixed minerals at different experimental conditions have been explained by Eh–pH diagrams of these minerals and the existence of free metal ions that are equally responsible for H2O2 formation besides surfaces catalytic activity. The results also corroborate the amount of H2O2 production with the rest potential of the sulphide minerals; higher is the rest potential more is the formation of H2O2. Most likely H2O2 is answerable for the oxidation of sulphide minerals and dissolution of non-ferrous metal sulphides in the presence of ferrous sulphide besides the galvanic interactions. This study highlights the necessity of revisiting into the electrochemical and/or galvanic interactions between the grinding medium and sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behaviour in the context of inevitable H2O2 existence in the pulp liquid.