Process Mineralogy

A strategic business case analysis and five-year plan are presented here for the Teck Applied Mineralogy group in Trail, British Columbia. Applied mineralogy is a key technical competency for Teck Resources that adds value in the fields of process mineralogy supporting optimization at the mining operations, and in ore characterization supporting evaluation of greenfield and brownfield properties. This 2016-2020 business plan for the applied mineralogy group is cognisant of the current mining industry downturn and recommends only incremental increases in staffing and project budgets during the period. Importantly, the plan recommends improved positioning of the group within Teck to maximize its contributions and a focus on customer relationships and timely delivery of value to projects. It also recommends the development of strategic outsourcing partnerships with commercial providers of applied mineralogy, including development of QA/QC protocols, so that this service is well managed within the company as a whole. There are some key capital expenditures necessary for the group in the next five years, including the replacement of mineralogical instrumentation, but their timing can be managed to coincide with improving corporate economics.

This study provides an up to date review of tannins, specifically quebracho, in mineral processing and metallurgical processes. Quebracho is a highly useful reagent in many flotation applications, acting as both a depressant and a dispersant. Three different types of quebracho are mentioned in this study; quebracho "S" or Tupasol ATO, quebracho "O" or Tupafin ATO, and quebracho "A" or Silvafloc. It should be noted that literature often refers simply to "quebracho" without distinguishing a specific type. Quebracho is most commonly used in industry as a method to separate fluorite from calcite, which is traditionally quite challenging as both minerals share a common ion-calcium. Other applications for quebracho in flotation with calcite minerals as the main gangue source include barite and scheelite. In sulfide systems, quebracho is a key reagent in differential flotation of copper, lead, zinc circuits. The use of quebracho in the precipitation of germanium from zinc ores and for the recovery of ultrafine gold is also detailed in this work. This analysis explores the wide range of uses and methodology of quebracho in the extractive metallurgy field and expands on previous research by Iskra and Kitchener at Imperial College entitled, "Quebracho in Mineral Processing".

Most ore deposits are complex and display a high degree of variability, arising from their inherent geological and mineralogical characteristics, which impact their beneficiation. Process mineralogy has been disruptive in the last two decades for geological exploration and for the mineral processing industry. Minerals are multicomponent elemental systems. Therefore, it is crucial and fundamental to understand the distribution of both minerals and metals they carry. Process mineralogy provides critical quantitative data i.e., mineral mass, liberation, and association of the minerals. Data can be used to determine variability of the ore, help to design, and develop the flowsheet to avoid extensive and time-consuming bench testing, and act as diagnostic to the beneficiation process. Automated mineralogy is commonly coupled with other advanced mineralogical techniques and can be implemented on projects throughout the mining chain, from initial stages of exploration to production. In this paper, we discuss concepts and examples of automated mineralogy applied to exploration and feasibility level.

In the cationic flotation of pyrolusite using dodecyl ammine (DDA), the depressive effect of sodium carbonate and calcium chloride on the calcite mineral was investigated systematically through flotation experiments, FTIR analysis, contact angle measurements and zeta potential tests. The microflotation experiments showed that both depressant agents decrease the flotation recovery of calcite significantly. In addition, sodium carbonate acts as activator agent for pyrolusite, and increases its floatability. The flotation experiments and contact angle measurements indicated that the selective depression effect of sodium carbonate on the calcite mineral is more than that of calcium chloride. As evidenced by zeta potential and FT-IR analysis, sodium carbonate decreases the negative charges on the surface of calcite mineral and subsequently reduces the adsorption of DDA collector through electrostatic forces. At a pH of 7.5, using 2000 g/t DDA and 1500 g/t sodium carbonate, a pyrolusite concentrate containing almost 40% MnO with 71.5% recovery is achieved by carrying out the ore flotation experiments on the tabling pre-concentrate.

The current paper deals with the reprocessing of mineralised mine waste from magnesite processing operations aiming at the recovery of the value. Two cases of mineralised waste were examined. First, waste derived from the removal of "fine" before hand-sorting, with particle size of-40 mm; this waste is usually stockpiled, comprises mostly of relatively coarse particles and is of relatively high-grade in magnesite. The separation methods applied were sorting, sink-float tests and magnetic separation. In general, the results from the application of physical separation methods were satisfactory. Commercial concentrates were obtained from sorting with MgCO3-grade ranging between 94 and 97% MgCO3 and yield about 46%; in case of mixing concentrate with middling, yield rises to 58-64% while grade slightly drops to 91-94%. Equally satisfactory were the results from sink-float separation at sp.gr. 2.88 (grade 95%, yield 45%) but questions arise from the small sp.gr. difference of magnesite and gangues. Finally, the less satisfactory results obtained from magnetic separation (grade around 80-85%, yield 33-65%) may be due to inefficient particles liberation and higher magnetic field is maybe required. Second, mineralised waste consisting of fine particles from size reduction operations were tested. These fines were processed with selective magnetic coating as well as co-agglomeration / magnetic separation techniques. In general, the results of physicochemical processing of fines were very encouraging both on single minerals and their artificial mixtures. In case of magnetic coating, the effect of pH, collector concentration and kerosene dose was studied first on single minerals. Optimum results were obtained for pH=9, dodecylamine solution 4.5x10-5 M and kerosene 2.5 L/t. Tests on artificial mineral mixtures at optimum conditions and magnetic separation at 0.8 A resulted in non-magnetic product with MgCO3-grade higher than 99%. The parameters examined in co-agglomeration / magnetic separation tests were the same as with magnetic coating. Best conditions for co-agglomeration on single minerals were pH=8, dodecylamine 2 kg/t, kerosene 4 L/t. Tests on mineral mixture reveals that the MgCO3-grade increases with dodecylamine but the yield drops.

The basic principle of a Knelson Concentrator (KC) is the differential settling velocities of particles in a medium under a centrifugal field. Whether the ore particles remain in the KC bowl depends on the terminal settling velocity (v) of the particles and the fluidization water flow velocity (V). Both variables are functions of the material properties and the operating parameters. This study reports the relations between the retained mass of quartz in each ring of the KC bowl (obtained by measuring the quartz mass ring by ring) and the rotational speed, fluidization water flow rate, and particle size. The retained mass of quartz in each ring was then modeled by approximating the fluidization water velocity and defining a new ratiov V /. Based on the experimental data, the functional relation between the quartz mass in each ring of the KC bowl and the feed particle properties/ operating conditions was fitted by a Weibull model. The model established by the ring-by-ring approach gives more accurate predictions than a previous model, and a model based on the quartz mass in the entire bowl.

The results of the study of rare-metal (Bi, Te, Se. Be, In) mineralization of skarn deposits (Sn, Zn) in the Pitkäranta Mining District, genetically related to the Salmi anorthosite-rapakivi granite batholiths of Early Riphean age are reported. Minerals and their chemical composition were identified on the base of optical microscopy as well as electron microanalysis. The diversity of rare-metal ore mineralization (native metals, oxides, and hydroxides, carbonates, tellurides, selenides, sulfides, sulphosalts, borates, and silicates) in Pitkäranta Mining District ores is indicative of considerable variations in the physicochemical conditions of their formation controlled by the discrete-pulse-like supply of fluids. Bismuth, wittichenite, and matildite are the most common rare-metal minerals. Sulfosalts of the bismuthinite-aikinite series are represented only by its end-members. The absence of solid solution exsolution structures in sulfobismuthides suggests that they crystallized fro...

Aurocyanide complex adsorption by naturally occurring carbon in Goldstrike ore has been implicated in a form of gold refractoriness known as Apreg-robbingB. The aim of this work was to establish a relationship between the aurocyanide uptake capacity of Goldstrike ore carbonaceous matter and the preg-robbing behavior of the parent ore. The effect of varying contact time between the aqueous gold-cyanide and the carbonaceous matter on the uptake of aurocyanide complexes was also investigated. In column adsorption experiments, the adsorbed gold concentration from 100 mL of a gold-cyanide solution depended on the flow rate of the solution. The concentration of adsorbed gold increased with solution flow rate, reaching a maximum concentration at a flow rate of approximately 0.05 mLrmin. As solution flow rate increased to more than 0.05 mLrmin, the concentration of gold adsorbed on the carbon decreased. Desorption of adsorbed gold into 75-mL sodium cyanide solution also depended on solution flow rate. Maximum desorption of gold occurred at a flow rate of approximately 0.14 mLrmin. The gold adsorbed at the lowest and highest flow rates was more slowly desorbed than the additional gold adsorbed at the intermediate flow rates. The concentration of gold remaining on the carbon after desorption at 0.14 mLrmin correlated to the amount of gold adsorbed at the highest and lowest flow rates in the adsorption experiments. This amount of slowly desorbed gold correlated to the preg-robbing behavior of the ore, whereas the amount of Ž. additional adsorbed gold i.e., gold adsorbed at intermediate flow rates was similar for all of the ores. For all of the carbonaceous matters studied, the concentration of gold sorbed in both batch and column experiments increased after autoclaving.

In this paper, the modeling of the liberation of scheelite is presented. A pattern of concentration experiments was performed to investigate the scheelite liberation and distribution density calculation procedure. In this work, one sample from a Mittersill tungsten ore was studied. This work describes a method for determining the downstream milling energy requirements for rod mill products based on a Bond mill test performance. The grade distribution of particles at a given size fraction was calculated using a predictive liberation model. The concentration behavior of these particles in size fractions was evaluated using batch concentrate tests. The recovery of particles in size/grade classes, image analysis using mineral liberation analysis (MLA), and function calculations were implemented for the modeling of the liberation. By describing the size, grade, and recovery data of particles in size/grade classes, a technique for the measurement of distribution functions was developed th...

A combination of microprobe proton induced x-ray emission (μ-PIXE), x-ray fluorescence (XRF), energy dispersive spectroscopy-scanning electron microscopy (SEM-EDS) and x-ray diffraction (XRD) techniques have been applied in the analyses of tailings of tin (Sn) mining activities in the Jos Plateau, Nigeria. This paper reviews the elemental composition and distribution maps across single mineral grains. The microscopic data obtained are discussed to understand the mineral phases; and the provenance of economically important recoverable metals, associated with the major indicator mineral element, contained in the investigated samples. Caution is required when using automated elemental analysis of the portable XRF instrument, of which user has no control. The obtained results must always be confirmed using another analytical method.

A combination of microprobe proton induced x-ray emission (μ-PIXE), x-ray fluorescence (XRF), energy dispersive spectroscopy-scanning electron microscopy (SEM-EDS) and x-ray diffraction (XRD) techniques have been applied in the analyses of tailings of tin (Sn) mining activities in the Jos Plateau, Nigeria. This paper reviews the elemental composition and distribution maps across single mineral grains. The microscopic data obtained are discussed to understand the mineral phases; and the provenance of economically important recoverable metals, associated with the major indicator mineral element, contained in the investigated samples. Caution is required when using automated elemental analysis of the portable XRF instrument, of which user has no control. The obtained results must always be confirmed using another analytical method.

The beneficiation of rare earth element (REE) minerals may include many different beneficiation unit operations, but the final process choice typically depends on the inherent mineral characteristics. The Nechalacho REE deposit contains multiple low specific gravity, diamagnetic silicate gangue minerals (predominantly feldspars and quartz) and high specific gravity, iron oxide minerals (magnetite and hematite). The valuable REE minerals (REM) in the deposit consist of a variety of relatively high specific gravity, paramagnetic minerals. A process has been proposed to concentrate the value REM through a combination of gravity (rejecting silicate gangue) and magnetic (rejecting iron oxide gangue) separation steps prior to froth flotation. This work employed a laboratory-scale process including two different gravity separation steps (Knelson and Falcon centrifugal concentrators) followed by a series of varying intensity wet drum magnetic separation steps as well as dry induced roll magnetic separation and wet high intensity magnetic separation steps. The resultant fractions have been characterized by XRD, ICP-MS, and QEMSCAN to identify the optimum fraction for downstream flotation separation. The combination of a Knelson Concentrator with low intensity wet drum magnetic separation was found to efficiently concentrate REM while also rejecting high specific gravity iron oxide minerals. A novel finding from this study is the concentration of heavy REE-bearing zircon into coarse size fractions after grinding.

Ore characterization is crucial for efficient and profitable production of mineral products from an ore deposit. Analysis is typically performed at various scales (meter to microns) in a sequential fashion, where sample volume is reduced with increasing spatial resolution due to the increasing costs and run times of analysis. Thus, at higher resolution, sampling and data quality become increasingly important to represent the entire ore deposit. In particular, trace metal mineral characterization requires high-resolution analysis, due to the typical very fine grain sizes (sub-millimeter) of trace metal minerals. Automated Mineralogy (AM) is a key technique in the mining industry to quantify process-relevant mineral parameters in ore samples. Yet the limitation to two-dimensional analysis of flat sample surfaces constrains the sampling volume, introduces an undesired stereological error, and makes spatial interpretation of textures and structures difficult. X-ray computed tomography (XCT) allows three-dimensional imaging of rock samples based on the x-ray linear attenuation of the constituting minerals. Minerals are visually differentiated though not chemically classified. In this study, decimeter to millimeter large ore samples were analyzed at resolutions from 45 to 1 μm by AM and XCT to investigate the potential of multi-scale correlative analysis between the two techniques. Mineralization styles of Au, Bi-minerals, scheelite, and molybdenite were studied. Results show that AM can aid segmentation (mineralogical classification) of the XCT data, and vice versa, that XCT can guide (sub-)sampling (e.g., for heavy trace minerals) for AM analysis and provide three-dimensional context to the two-dimensional quantitative AM data. XCT is particularly strong for multi-scale analysis, increasingly higher resolution scans of progressively smaller volumes (e.g., by minicoring), while preserving spatial reference between (sub-)samples. However, results also reveal challenges and limitations with the segmentation of the XCT data and the data integration of AM and XCT, particularly for quantitative analysis, due to their different functionalities. In this study, no stereological error could be quantified as no proper grain separation of the segmented XCT data was performed. Yet, some well-separated grains exhibit a potential stereological effect. Overall, the integration of AM with XCT improves the output of both techniques and thereby ore characterization in general.

The growth of demand for metallic minerals has faced with the need for new techniques and improving technologies for all mine life cycle operations. Nowadays, the exploitation of old tailings and mine wastes facilities could represent a solution to this demand, with economic and environmental advantages. W-Sn Panasqueira Mine has been operating for more than 100 years. Its first processing plant “Rio” was located near Zêrere river being the mineral processing residues deposited on the top hillside on the margin of this river in Cabeço do Pião tailings dam. The lack of maintenance and monitoring of this enormous structure in the last twenty years represents high risks to the environment and population of the surrounding region. The re-mining of the tailings by hydrometallurgical methods was considered, in order to satisfy these two enounced conditions - metals demand and environmental risk, aiming for the sale of the metal to pay the environmental intervention. Field samples campaign...

The Sar Cheshmeh porphyry copper deposit located in southern central Iran, with the sulphide ore reserves of about 1.22 billion tons averaging 0.69% Cu and 0.027% Mo, ranks with the great copper deposits of the world. Minor amounts of gold, silver, and selenium are also present in the ore. The current operation at Sar Cheshmeh includes open-pit mining, milling, smelting, and copper refinery, with a rated production of 158,000 t/y of cathode copper. Concentrator flowsheet consists of crushing, grinding, bulk flotation of copper sulphides and molybdenite, differential flotation of molybdenite, and dewatering circuits, capable of processing 43,000 t/d of about 1% Cu ore, and producing approximately 1,300 t/d of 30% copper concentrate containing 0.8-1.2 ppm Au and 30-40 ppm Ag, as well as a molybdenum concentrate of 55% Mo. In this study a comprehensive sampling of the concentrator plant was carried out and over 130 samples of mill products were analyzed for 48 elements. The metallurgical balances of copper and gold during the sampling periods show that, in compare with 80% copper recovery, about 44% of the gold is recovered as the by-product into the copper concentrate, and the remaining is lost to the tailing. In this manuscript the geology and current plant practice at Sar Cheshmeh are briefly described, and an assessment of metallurgical performance of the concentrator plant at various time intervals with special reference to the behavior of gold in the mill products are presented.

Acceptable zircon for composite formulation in the aerospace industry requires that the mineral contains a minimum of 65% zirconia (ZrO2). Despite having vast deposits of zircon, Nigeria’s aerospace industry has historically relied primarily on imported mild steel tubes for solid rocket
motor cases (SRMCs) construction, resulting in three major challenges: low strength-to-weight ratio, pressure, and temperature containment. In this study, the Arikya zircon deposit located in northern Nigeria was investigated with the aim of upgrading low-grade zircon ore using magnetic and gravity separation processes for use in composite formulation for SRMCs. The dry high-intensity magnetic separator (DHIMS) produced a ZrO2 grade of 52.48%, recovery of 57.99%, and an enrichment ratio of 0.78 with a separation efficiency of 0.56, while the air-floating separator (AFS) generated the highest of 65.52% ZrO2 grade with 70.81% recovery and enrichment ratio of 1.25 with a separation efficiency
of 0.25. The ZrO2 content increased from 40.77 to 65.52% after beneficiation. Iron oxide and titanium dioxide contaminants at 0.73 and 0.83% were reduced to 0.66 and 0.54%, respectively, while the specific
gravity increased from 4.4 to 4.6 g/cm3. The ZrO2 content and specific gravity were improved to the minimum standard specified for zirconia-reinforced composite application and competed effectively with industrially/globally accepted zircon. These results demonstrated the efficacy of combining DHIMS and AFS to upgrade the low-grade zircon ore from Arikya, Nasarawa State.

In this paper, the modeling of the liberation of scheelite is presented. A pattern of concentration experiments was performed to investigate the scheelite liberation and distribution density calculation procedure. In this work, one sample from a Mittersill tungsten ore was studied. This work describes a method for determining the downstream milling energy requirements for rod mill products based on a Bond mill test performance. The grade distribution of particles at a given size fraction was calculated using a predictive liberation model. The concentration behavior of these particles in size fractions was evaluated using batch concentrate tests. The recovery of particles in size/grade classes, image analysis using mineral liberation analysis (MLA), and function calculations were implemented for the modeling of the liberation. By describing the size, grade, and recovery data of particles in size/grade classes, a technique for the measurement of distribution functions was developed that relates beta distribution, a model for the function based on the incomplete beta function, and a solution to produce liberation modeling. It was shown that the predicted results agreed well with the observed results. With a procedure for measuring the liberation, it was possible to carry out the first experimental measurement of the beta distribution. This liberation/concentrate model has wide potential applications for metallurgy and plant design, where the liberation modeling is to be determined with the distribution density solution to the predictive mineral liberation function equation, which includes the liberation of ore samples and their liberation characteristics.

Baryte is a common gangue mineral in many ore systems. Here, we report on a study of baryte chemistry in samples of ore and processing materials (flotation feed, flotation concentrate, flotation tailings, concentrate leach discharge, and tailings leach discharge) from the Olympic Dam Cu-U-Au-Ag deposit, South Australia. Elements that commonly substitute for Ba in the baryte lattice, including Sr and Ca, are measured in variable concentrations reflecting grain-scale zonation and heterogeneity at the scale of the deposit but appear unaffected during processing. Variation in the concentration of some other elements, notably Cu, reflect both the heterogeneous nature of flotation feed and the intimately intergrown character of the sulfide-sulfate assemblage. Measured Pb concentrations in baryte progressively increase during processing from flotation feed to flotation concentrate, and particularly in concentrate leach discharge. Such data suggest that, during sulfuric acid leaching, baryte contained within the concentrate is able to incorporate quantities of Pb that have been mobilized during breakdown of Pb-bearing minerals (notably U-minerals containing radiogenic lead). This takes place via surface adsorption followed by rapid coupled dissolution-replacement driven incorporation throughout the grain. Results suggest that baryte may scavenge non-target elements during processing and contribute to an understanding of mobility, mineralogical location, and evolution in the deportment of radionuclides through the processing cycle.

Egyptian oil shale from Wadii El-Nakhil, Red sea region was upgraded using enhanced gravity separation. The oil shale sample was characterized physically and chemically to determine its mineral content and characteristics. The sample includes quartz, siderite, apatite, anhydrite and calcite. The clay mineral is mainly represented by kaolinite while the organic matter is 30%. The ground sample (less than 50 microns) was classified into two fractions. The coarser was higher than 25 µm while the finer was less than 25 µm. The lower and upper levels of both the centrifugal force and water pressure have been suggested to construct the design for Falcon Concentrator type SB-40. The coarse concentrate of 42% kerogen with 94.35% recovery was achieved at 60 Hz (equivalent to G-force 176) and water pressure of 4 Psi from feed of 29% kerogen. The fine concentrate of 38.46% kerogen with 85.4% recovery was achieved at 70 Hz (equivalent to G-force 243) and water pressure of 2 Psi from feed of 33% kerogen.

Rare earth element (REE)-bearing carbonatite deposits commonly contain a wide range of different REE-and REE-bearing minerals associated with various gangue matrices. In order to select the most-suitable mineral processing technique for these deposits, it is essential to identify and quantify the minerals of interest, including their liberation, associations and grain size distribution, along with whole rock compositions. These data are also vital for ore feed optimisation and metallurgical troubleshooting during and after designing a mineral processing flowsheet. This paper summarises the key mineralogical parameters needed before conducting metallurgical beneficiation tests, using the Songwe Hill carbonatite deposit as an example. This REE ore deposit consists of poorly-liberated synchysite-(Ce), which hosts the light rare earth elements including Nd plus some heavy rare earths and well-liberated apatite, which hosts 50% of Gd, 63% of Dy and 71% of Y (heavy rare earth elements) in the deposit. For all REE heavier than Gd, apatite is the most important REE host, however, for the two REE where data are available in both synchysite-(Ce) and apatite (Dy and Y), synchysite-(Ce) still accommodates > 25% of the whole-rock HREE content. Both of these ore minerals are associated with ankerite, calcite, and to a lesser extent with iron oxides/carbonates, K-feldspar, strontianite and baryte. According to the quantitative mineralogical data, the possibility of using gravity separation, magnetic separation, froth flotation and leaching to process Songwe Hill carbonatite ore is discussed and a potential beneficiation flowsheet is presented.

An essential operation in the mineral processing of copper ores into concentrates is blending, as it guarantees a constant feed for the flotation cells, increases metal recovery rate and reduces tailings. In this study, copper ores from Huelva province (Spain) were investigated by quantitative XRD (X-ray diffraction) methods to optimize blending and detect penalty minerals, which can affect flotation and concentrate quality. The Rietveld method in combination with cluster analysis, PLSR and more traditional chemical analysis provide a more complete and in-depth characterization of the ore and the whole process. The mineralogical monitoring can be fully automated to enable real-time decision making.

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Se obsevaron particulas finas en relaves generados a partir de la flotacion de minerales. Mediante el uso de barrido de emision de campo (FE-SEM), se observaron las particulas finas a nivel submicronico con alta resolucion. Para la observacion de la seccion transversal, las muestras de relaves fueron preparados por la seccion pulida. Con la combinacion de estas tecnicas, se determino la distribucion de los metales pesados y metaloides en las particulas finas. Como resultado, cuando se observaron las superficies de las particulas, la deteccion de Pb, Sb, Fe como se indica que estos elementos coexisten o existen por separado en las particulas finas 1-3um. En la observacion de la seccion transversal, se encontro que las concentraciones de Zn y Fe diferente en las particulas.

Repositories of historical tungsten mining tailings pose environmental risks, but are also potential resources for valuable metals. They still contain large tonnages of useful minerals and metals, reflecting the inefficient extraction methods and/or low metal prices at the time they were mined. The focus of this study is to evaluate the technical viability of reprocessing the tailings to recover some of the contained valuable minerals and metals, as well as reducing the negative environmental impact associated with the tailings. Geometallurgical studies were conducted on drill core samples taken from the Smaltjärnen tailings repository of the closed Yxsjöberg tungsten mine, Sweden. The collected samples were characterized physically, chemically, and mineralogically. Knelson concentrator dry low-and high-intensity magnetic separation methods were tested as potential beneficiation methods. The tailings are dominated by the −600 to +149 µm particles. The highest concentration of tungsten (W) was 0.22% WO 3. Using a Knelson concentrator, scheelite (main W mineral) recovery was enhanced, with 75 wt.% tungsten recovered in the 34 wt.% heavy concentrate. Only 1.0 wt.% sulphur (S) reported to the non-magnetic fraction. Based on the findings, a methodology and a preliminary process flowsheet for reprocessing the tailings is proposed.

Where a licence is displayed above, please note the terms and conditions of the licence govern your use of this document. When citing, please reference the published version. Take down policy While the University of Birmingham exercises care and attention in making items available there are rare occasions when an item has been uploaded in error or has been deemed to be commercially or otherwise sensitive.

The Sar Cheshmeh porphyry copper deposit located in southern central Iran, with the sulphide ore reserves of about 1.22 billion tons averaging 0.69% Cu and 0.027% Mo, ranks with the great copper deposits of the world. Minor amounts of gold, silver, and selenium are also present in the ore. The current operation at Sar Cheshmeh includes open-pit mining, milling, smelting, and copper refinery, with a rated production of 158,000 t/y of cathode copper. Concentrator flowsheet consists of crushing, grinding, bulk flotation of copper sulphides and molybdenite, differential flotation of molybdenite, and dewatering circuits, capable of processing 43,000 t/d of about 1% Cu ore, and producing approximately 1,300 t/d of 30% copper concentrate containing 0.8-1.2 ppm Au and 30-40 ppm Ag, as well as a molybdenum concentrate of 55% Mo. In this study a comprehensive sampling of the concentrator plant was carried out and over 130 samples of mill products were analyzed for 48 elements. The metallurgical balances of copper and gold during the sampling periods show that, in compare with 80% copper recovery, about 44% of the gold is recovered as the by-product into the copper concentrate, and the remaining is lost to the tailing. In this manuscript the geology and current plant practice at Sar Cheshmeh are briefly described, and an assessment of metallurgical performance of the concentrator plant at various time intervals with special reference to the behavior of gold in the mill products are presented.

In this study, enrichment possibility of apatite-bearing iron ore sample was investigated to obtain iron and phosphate concentrate, separately. It was determined that the raw ore contains 35.75% Fe and 5.36 P2O5. Magnetite in the ore can be enriched with low-intensity magnetic separator, and apatite can be enriched from the non-magnetic product by flotation. For this purpose, the raw ore was ground below 106 m size and then subjected to magnetic separation (with two cleaner and one scavenger), and following flotation (with three cleaner and one scavenger). According to the experimental result, a magnetic concentrate containing 63.55% Fe and 1.65% P2O5, and phosphate concentrate containing 25.33% P2O5 and 6.45%Fe were obtained. The results show that there are still difficulties to obtain plausible iron and phosphate concentrates in terms of phosphorus and iron impurities for iron and phosphate concentrate, respectively. It is concluded that it is hard to separate the iron and phosphorus phases from each other, possibly due to complex mineralogical composition and poor mineral liberation of the ore.

In this study, the effect of an external magnetic field on cationic flotation of quartz from magnetite was investigated by using a magnetic micro-flotation column. For this purpose, a micro-flotation column jacketed with three coils was fabricated to create a funnel-shaped magnetic field. Both the theoretical magnetic field strengths and the magnetic forces were calculated. The results from flotation experiments with 6 Â 10 À5 M dodecylamine as collector using a binary mixture of quartz and magnetite as a feed material show that the separation efficiency increased from 0% without the magnetic field to 88% in the presence of the magnetic field. The significant enhancement in separation efficiency is evident.

This study compares the efficiency of different methods in beneficiation of iron ore from Doğanşehir (Malatya) region. It is a low grade ore with 27.43 % Fe content that requires concentration to meet the specifications of blast furnace feed. The ore mainly contains magnetite, ferro-actinolite and magnesioferrite. The collected samples were classified into different size fractions after size reduction and then subjected to different gravimetric and magnetic separation methods for concentration. The results showed that particle sizes affected the concentrations to a large extend, but wet magnetic separation yielded comparably better results than gravimetric methods. A concentrate assaying 65.66 % Fe and 0.38 % K2O+Na2O was obtained with 78.11 % recovery by wet magnetic separation. It was concluded that concentrates meeting blast furnace specifications could be obtained from this low grade iron ore. It was also concluded that the proposed separation flow sheet can be applied to simila...

Characterization of heterogeneous materials, such as particles from mechanically processed waste printed circuit boards, is a challenging task. The majority of characterization methods either give average information or information that is very limited and in a tiny area of specific interest. That said, capturing such heterogeneity is significantly important for any kind of processes. Degree of liberation, indicating how much the target component is liberated from the non-valuable components, is a key property to determine the success of subsequent process for valuable material recovery. This work analyzed the degree of liberation of metals within the products of hammer milling process via the combination of image acquisition and analysis. The digital microscope and a scanning electron microscope (SEM) coupled with the energy dispersive spectroscopy (EDS) were used for image acquisition and elemental mapping, in order to evaluate the selective liberation under different milling conditions (i.e., feed mass, milling time) for different metals (mainly Cu and Al) and particle size fractions. The obtained liberation degree was also modelled and determined the liberation parameters that were compared. The results showed that the degree of liberation significantly depend on the milling conditions and metals we analyzed, and well correlated with the selective metal enrichment behavior. Results between the two methods showed some similarities and discrepancies. The advantages and disadvantages of the above two methods were identified and discussed in the paper, in addition to their methodological developments.

Alkaline rock-type rare-earth (RE) ores have significant utilisation value. However, the exploitation of such resources faces great challenges owing to the complex mineral and element assemblages. Composite collectors exhibit excellent performances, which may provide solutions to the flotation problem of alkaline rock-type RE ores. Therefore, 16 collectors typically used in RE ores flotation were selected. Flotation tests were performed to identify collectors with high selectivity and collection ability for RE minerals, then nine composite collectors were prepared by combining the satisfactory collectors. The flotation performances of single and composite collectors for RE minerals were examined, and the composite collector FA301 with different carbon chain lengths was identified as the best one. When FA301 was applied in optimal conditions of slurry temperature, grinding size, collector and inhibitor dosage, RE concentrate yield of 6.29%, REO grade of 32.013%, and recovery of 59.02...

This article aimed to separate ilmenite (FeTiO3) mineral from tin tailing applying a single splitter IRMS (Induced Roll Magnetic Separator). Ilmenite mineral is the substantial main source for TiO2. This work used air table middling for feeding. The mineral components of middling feeding from air table using grain counting analysis were found as follows: cassiterite (48.61%), ilmenite (21.36%), monazite (18.56%), pyrite (4.60%), zircon (5.85%), quartz (0.71%), anatase (0.27%), and tourmaline (0.02%), It was found that electrical current and opening of single splitter affected the degree of separation addressing to ilmenite recovery and ilmenite grade. The finding showed that current of 15 Ampere and single splitter with opening 4.25 cm yielded ilmenite recovery more than 74%. The high grade ilmenite (90.46 %) and recovery of 29.38% was obtained using 5 Ampere  with single splitter opening of 1.0 cm. Up to date, the study on ilmenite separation from tailing only focused on the effect...

The Bear Lodge Project is an important rare earth deposit in the United States. While extensive previous studies have settled on a crushing, screening, gravity and magnetic separation process, this study sets out to investigate an alternative flowsheet based on flotation and wet high intensity magnetic separation (WHIMS) to effectively beneficiate the rare earth oxide (REO) content of the Bear Lodge ore. Mineral characterization found ancylite was the dominant rare earth mineral, associated mainly with calcite and strontianite. Electrokinetic studies on the effects of pH, concentrations of various ions (Sr 2+ , HCO 3 − and CO 3 2−) and hydroxamate concentrations were performed to establish the electric nature of the Bear Lodge ore. The isoelectric points (I.E.P) of the material in distilled water was around 5.27. The Sr 2+ and CO 3 2− ions in solution significantly affected the surface charge of the material. Adsorption studies suggested that the mechanism of hydroxamate adsorption is chemisorption, as hydroxamate adsorption increased on the Bear Lodge ore with an increase in temperature. WHIMS was employed to remove the iron content to reduce the interference of iron in following flotation process and consumption of hydroxamic acid. After cleaner flotation a REO grade of 11.2% at 72.7% recovery from a feed material of 4.5% REO was obtained. In light of the loss of REO in WHIMS process, it is possible to produce a concentrate containing 11.2% REO grade at 61.2% recovery.

To increase resource efficiency, mining residues–especially tailings–have come into the focus of research, companies, and politics. Tailings still contain varying amounts of unextracted elements of value and minerals that were not of economic interest during production. As for primary mineral deposits, only a small share of tailings offers the possibility for an economic reprocessing. To minimize exploration expenditure, a stepwise process is followed during exploration, to estimate the likelihood of a project to become a mine or in this case a reprocessing facility. During this process, costs are continuously estimated at least in an order of magnitude. Reprocessing flowsheets for copper mine tailings in Chile were developed and costs and revenues of possible products from reprocessing were examined for a rough economic assessment. Standard cost models with capex and opex for flotation, leaching, and magnetic separation were adopted to the needs of tailings reprocessing. A copper t...

The primary raw material of the steel industry is iron. This paper aims to optimize magnetite recovery from wet tailings by increasing the iron content in the concentrate of the line. To manage tailings, a Wet Tailing Processing (WTP) line constructed at Gol-e-Gohar Iron Ore Company to recover the magnetite. The dominant crystalline phases in these tailings were quartz, albite, talc, hematite, and calcite. The line feed is 45 microns, which is not suitable for the gravity method. Thus, separation can achieve using only the magnetic method. Because of the high iron content in the tailings, a wet magnetic separator is used. According to the results, the proposed medium-intensity separator and the associated circuit modifications increase iron recovery from 7 to 30 percent; resulting in 150 tons of annual production; preventing loss of iron through concentrator plant tailings, and increasing the Blain number by 50 to 100 units in the hematite plant. Furthermore, water consumption is significantly reduced by replacing old wet tailings of the concentrator plant with new wet tailings as the feed, which is another significant achievement of this research. Instead of fresh water, saline water with flow rate of 250 cubic meters per hour are used.

is the largest underground tin mining operation in the world. The main methods for processing the ore are gravity separation and flotation. In 2016 a new addition to the plant was an X-ray transmission (XRT)based ore sorting island. The objective was to reject waste from a marginal development waste dump, and it is now also treating low grade material from the underground mine. XRT sorting discriminates particles on a planar projection of the density of matter attenuating X-ray radiation. Physical separation is achieved by means of an array of high speed air jets. The feed grade of approximately 0.6% tin is concentrated to 2.8% tin in the product at recovery of 90% in about 19% mass yield to product. This enriched product fraction from the sensor-based ore sorting plant feeds into the main plant. The plant, which is designed for 3600 tonnes per day, went into operation in 2016. The total capital expenditures of 24 M$ US were paid back within 4 months. The projected contribution to the production of refined tin for the year 2017 was 6000 tonnes. In summary, the positive impacts were:-Value added from previously sub-economic waste that can be mined with lower costs.-Increase of productivity in the main wet sections of the plant.-Reduction in cutoff grade and increase in reserves.-Reduction of environmental impacts. The paper describes the project from concept, through the development phase, including test work confirming feasibility, to a discussion of operational data and includes shared experiences from operation and maintenance.

This study is conducted with the aim of investigating the efficiency of open and closed-circuit molybdenite ore comminution processes (primary and secondary mill, respectively), through mineralogical study of mills feed and product. For this purpose, particle size distribution, minerals distribution, degree of liberation and interlocking of minerals in mills feed and product were studied. According to the results, chalcopyrite, molybdenite, pyrite and covellite constitute the major part of the mineral composition of open-circuit mill feed. Minerals at the mill product, in the order of abundance include liberated molybdenite particles, liberated chalcopyrite and interlocked chalcopyrite with pyrite, liberated and interlocked pyrite particles, and associated silicate gangues. The d50 values of the feed and product particles of the open-circuit mill are equal to 13.80 and 13.40 microns, respectively. Degree of liberation of molybdenite for the feed and product of this mill is almost th...

Estonia holds the largest unexploited sedimentary phosphate rock reserves in the EU. Compared to other sedimentary and also igneous phosphate rocks the Estonian phosphorite is particularly outstanding by its remarkably low Cd (1-5 ppm) and trace U content (~50 ppm) and also contains valuable elements such as rare earth elements (REEs), which could be used in important high-technologies in the future. In this paper beneficiation by reverse flotation of the silicates for three types of Estonian phosphorite ores was investigated. The apatite concentrates with grades of over 33 wt% P 2 O 5 at recoveries of 84.8 to 89.2% were achieved. The concentrates had low MgO contents and grades of FeO ranged from 1.29 up to 2.54 wt%. The study gives an overview of the specifics of enrichment of Estonian shelly phosphorite ores, provides a ground for future in-depth and pilot-scale experiments to find proper solutions for the Estonian phosphorite processing and will be applied in the feasibility studies of the phosphorite industrialization.

The Pindos ophiolite complex, located in northern Greece, hosts small podiform chromitites characterized by very low platinum group element (PGE) grades. PGE (excluding Os) analyses from four chromitite samples, collected for this study, are below 400 ppb. Consequently, we used the technique of hydroseparation to concentrate PGM from Pindos chromitites. More specifically, we investigated two separate composite samples from Pefki and Milia chromitites. The Pefki concentrate contains 67 PGM that include secondary Ru-bearing minerals, ruarsite (RuAsS), laurite (Ru,Os)S 2 , irarsite (Ir,Ru,Rh,Pt)AsS, alloys of Os-Ir-Ru, hollingworthite (Rh,Pt,Pd)AsS, paolovite (Pd 2 Sn), braggite (Pd,Pt,Ni)S, sperrylite (PtAs 2) and four unnamed PGM, electrum (Au,Ag) and native silver. The investigation of the Milia concentrate yields fifty one grains of PGM including primary laurite, Os-Ir alloys, erlichmanite, secondary Ru-bearing minerals, irarsite, Ru-alloys, ruarsite and Ru-based metals sulphides. PGM occur as both single and polyphase particles in both concentrates. The bulk of mineralization for grains between 5 and 15 lm is finer-grained in Milia (64.7%) than in Pefki (41.8%). The latter concentrate hosts considerably more altered PGM grains than the former. The hydroseparation process has recovered significantly more, as well as novel, PGM grains than the in situ mineralogical examination of single chromitite samples from the neighboring Korydallos occurrence. Although, most of the PGM occur as free particles and in situ textural information is lost, single grain textural evidence is observed. The mineralogical and grain size differences between the two samples may reflect styles of mineralization and indicate significant remobilization of PGE in Pefki. The latter possibility is suggested by the presence of secondary PGM, which may be related to the different alteration processes that affect the Pefki and Milia chromitites. In summary, this study provides significant information on the particles, grain size and associations of PGM, which are critical with respect to the petrogenesis and mineral processing of these minerals.

It is known that ore containing cassiterite (SnO 2) has been our most important source of tin since antiquity and its successful separation continuously pose problems to mineral processors. The situation is more pronounced since the depletion of the more easily recoverable alluvial reserves forces us to work with the more complex deposits such as hardrock cassiterite ores. In order to understand more about the challenges in processing complex tin ore deposits, a metasedimentary rock ore sample from a mine in Malaysia was used in this study. Chemical analysis by wet method shows that SnO 2 content in the sample was 2.86%, while for mineralogical analysis, the x-ray diffractogram (XRD) of the sample had identified that besides cassiterite, the sample also contained minerals such as quartz (SiO 2) and clinochlore. Furthermore, the FESEM (field emission scanning electron microscope) micrograph analysis carried out on a polish section of the sample indicated that the fine cassiterite particles (approximately 80 µm) were found to be disseminated in the quartz minerals. Prior to the separation processes, grindability studies were carried-out on crushed samples to liberate the cassiterite from other gangue minerals and at the same time, avoid producing high percentage of fines. For the separation of tin from gangue minerals on the ground samples, two stages of gravity separations by shaking tables were carried out. The first stage was run on ground samples and for the second stage, the middling product from the first stage was re-tabled. Magnetic separation process on Concentrate 1 (stage 1) and Concentrate 2 (stage 2) products from the shaking table increased the grade of SnO 2 to 46.85% and 61.90% respectively (as a non-magnetic products). Further concentration process of these non-magnetic products by high tension separator, increased the grade of SnO 2 from 85.05% to 98.77%.

Phosphate fines (size ≤ 45 μm), from the Red Sea Region, Egypt, were subjected to beneficiation process to recover the phosphorite grains from such fines. The sample is characterized by low P 2 O 5 content (18.78%) associated with gangue minerals such as silica (22.77% SiO 2) and carbonate (2.01% MgO) indicating its low grade. Phosphate pre-concentrate was prepared by gravity separation, using Falcon, where 63.37% of the phosphate, with 24.94% P 2 O 5 , 1.46% MgO, and 10.54% SiO 2 , was firstly recovered reducing the mass flow to the subsequent beneficiation process. However, the gravity tailing was the feed for the reverse phosphate flotation using bench scale column flotation where oleic acid was used as a carbonate gangue collector and amylase enzyme as a phosphate depressant. Under the appropriate flotation conditions (0.1% amylase, 5•10 −4 mol/dm 3 oleic acid, and temp. 30°C), phospho-concentrate assaying 0.61% MgO, 13.14% SiO 2 , and 27.85% P 2 O 5 with a P 2 O 5 recovery of 55.45%, was finally obtained without the use of expensive depressants, e.g., phosphoric acid or sodium silicate. A tentative flow sheet for the whole process was postulated.