Amenability for processing of oolitic iron ore concentrate for phosphorus removal

Beneficiation routes aimed at dephosphorisation of oolitic gravity magnetic concentrate and involving a combination of roasting, re-grinding, magnetic separation and water and acid leaching are investigated. Roasting was carried out at 900 °C for 1 h without or with lime or sodium hydroxide as roasting additives. When additives were used, cement phases of Si–Al–Na–Ca–O type were detected as well as the mineral giuseppettite. During the thermal process sodium silicate is liquefied and the newly formed phases coat the oolites and penetrate inside the cracks. Energy Dispersive Spectroscopy analysis has indicated that the zone surrounding the oolites consists of Na, Al and Si phases with part of phosphorus being captured there. As a result of the alkaline roasting, goethite is partly transformed to magnetite and this reduction is reinforced with an increase in sodium hydroxide dosage. Investigation of redistribution of phosphorous shows that it could be only partly separated if leaching is not accompanied by re-grinding and physical separation. The recommended dosage of the reductive agent for the final flowsheet is 8 mass% ratio to concentrate. Grinding to a mean size of 0.040 mm, with water and acid leaching and double magnetic separation creates conditions to obtain a high-quality iron concentrate with 65.97% Fe and recovery of 92.43%, with simultaneous decrease in the phosphorus content from 0.71% to 0.05%.     

An innovative process for extracting boron and simultaneous recovering metallic iron from ludwigite ore

Ludwigite ore has not yet been utilized on an industrial scale due to its complex mineralogy and fine mineral dissemination in China. Boron–iron separation and dissolution activity of boron-bearing minerals in alkaline liquor are the two key issues in the utilization of ludwigite ore, governing the boron recovery as well as operating cost. This paper proposes an innovative process for extraction of boron and iron from ludwigite ore based on coal-based direct reduction process with sodium carbonate (Na₂CO₃). The novel process involves reduction roasting, combined leaching and grinding of reduced ludwigite ore, followed by magnetic separation of leach residue, and experimental validation for each of the processing steps is demonstrated. Alkali-activation of boron and metallization of iron were synchronously achieved during carbothermic reduction of ludwigite ore in the presence of Na₂CO₃. Consequently, boron was readily extracted in the form of sodium metaborate (NaBO₂) with water at room temperature during ball mill grinding, and metallic iron powder was recovered from the leaching-filtering residue by magnetic separation. Boron extraction of 72.1% and iron recovery of 95.7% with corresponding iron grade of 95.7% in the magnetic concentrate were achieved when ludwigite ore was reduced with 20% sodium carbonate at 1050 °C for 60 min.     

A novel method to recover zinc and iron from zinc leaching residue

A novel method to recover zinc and iron from zinc leaching residue (ZLR) by the combination of reduction roasting, acid leaching and magnetic separation was proposed. Zinc ferrite in the ZLR was selectively transformed to ZnO and Fe₃O₄ under CO, CO₂ and Ar atmosphere. Subsequently, acid leaching was carried out to dissolve zinc from reduced ZLR while iron was left in the residue and recovered by magnetic separation. The mineralogical changes of ZLR during the processes were characterized by XRF, TG, XRD, SEM–EDS and VSM. The effects of roasting and leaching conditions were investigated with the optimum conditions obtained as follows: roasted at 750 °C for 90 min with 8% CO and CO/CO + CO₂ ratio at 30%; leached at 35 °C for 60 min with 90 g/l sulfuric acid and liquid to solid ratio at 10:1. The iron was recovered by magnetic separation with magnetic intensity at 1160 G for 20 min. Under the optimum operation, 61.38% of zinc was recovered and 80.9% of iron recovery was achieved. This novel method not only realized the simultaneous recovery of zinc and iron but also solved the environmental problem caused by the storage of massive ZLR.     

A non-invasive technique for sorting of alumina-rich iron ores

This paper describes an Infrared (IR) thermography based technique for sorting of iron ores consisting of alumina-rich particles of relatively low thermal absorptivity as compared to iron-rich particles in the ores. The technique primarily consists of selection of iron ores with Fe compositions ranging from 59 to 69 wt.% and alumina (Al₂O₃) from 0.35 to 8.85 wt.%, crushing the ores up to the particle size range around 10 mm. The iron ore fines are uniformly heated using heat source of wavelength ranging from 10⁻² to 10⁻⁶ m for a period of time sufficient to create a difference in infrared emission between the ore particles. The thermal image of the heated ores is captured by IR thermography. The alumina-rich iron ore particles are heated up less as the thermal absorptivity of these ores is less than the ores with high iron content. Thus, the alumina-rich iron ore particles can be identified by observing the temperature profile and/or thermal image of these ores. This technique of ore recognition can be useful in improving the feed quality of iron ore to the blast furnace in iron and steel industries by rejecting the alumina-rich ores through modification in the presently existing processes.     

Use of oxidation roasting to control NiO reduction in Ni-bearing limonitic laterite

Use of limonitic laterite as an iron source in conventional ironmaking is restricted due to its gangue composition and small particle size. Even direct reduction cannot effectively produce direct reduced iron (DRI) because NiO would be reduced together with iron oxide to form Fe–Ni. A small amount of Ni (about 2 wt.%) in DRI degrades the physical properties of final steel products. The current study investigated how oxidation roasting of limonitic laterite ores affected NiO reduction, with the goal of producing Ni-free DRI and Ni-bearing slag. Ni-bearing slag can be a good secondary Ni resource. Oxidation roasting made NiO inert under H₂ reduction at 900 °C by forming Ni-olivine. Optimum roasting temperature was proposed by examining phase transformation of limonitic laterite ores during heating and by FactSage calculation of the equilibrium Ni fraction in Ni-bearing phases. Furthermore, the effect of Mg–silicate forming additives on the control of NiO reducibility was clarified to maximize the suppression of NiO reduction. Among various additives such as MgSiO₃, Mg₂SiO₄ and Fe–Ni smelting slag, Ni-free olivine-typed flux was found to be the most effective form of Ni-olivine because Ni–Mg ion exchange between Ni-bearing phase and Ni-free olivine occurs more readily than other Ni-olivine formation schemes. Finally, the mechanism of Ni-olivine formation during roasting was studied using a diffusion couple test. Calculated diffusivity values of Ni in Mg₂SiO₄ indicated that the two major routes of Ni-olivine formation while roasting limonitic laterite ore are (1) Ni partitioning within Mg–Ni silicate before crystallization and (2) Ni diffusion from spinel to Ni free olivine after crystallization.     

Comprehensive recovery of the components of ferritungstite base on reductive roasting with mixed sodium salts,water leaching and magnetic separation

Ferritungstite ores have great commercial value because of the huge reserve and high content of W, Mo and Fe. But their economic recovery has long been a challenge due to its complex mineralogy and heterogeneous. The current study investigated how reductive roasting of ferritungstite ores with mixed sodium salts affected the phase evolution of W, Mo and Fe through Micro-area XRD and Powder XRD, with the goal of comprehensive transformation of ferritungstite. Reductive roasting with mixed sodium salts at 800 °C transformed ferritungstite to Na₂WO₄ and magnetite (Fe₃O₄), which were easily recovered by water leaching and magnetic separation. Furthermore, a lot of pores and gaps rather than sintering or agglomeration was observed in the ore particles after roasting by SEM-EDS, which was beneficial for the water leaching of W and Mo. As a result, 96.40% of W and 96.64% of Mo were extracted after water leaching, while an iron concentrate with an Fe content of 55.65% and recovery of 83.30% was obtained after magnetic separation. These results suggested such process would be applicable to the comprehensive recovery of valuable metals from ferritungstite ores, as well as similar tungsten ores and scraps.     

Bioleaching of six nickel sulphide ores with differing mineralogies in stirred-tank reactors at 30 °C

A bioleaching study was conducted with six nickel sulphide ores from different geographical locations across Canada. Mineralogical and chemical examination revealed considerable variability between the samples, particularly in the silicate phases. The ores contain 0.3–1% nickel, primarily in pentlandite and secondarily in pyrrhotite. Copper is present primarily in chalcopyrite, and cobalt in pentlandite. The ores were subjected to the same crushing and grinding procedure, and bioleached under the same conditions for 3 weeks with a mixed culture of iron- and sulphur-oxidizing bacteria. Stirred-tank experiments with finely ground ore (−147 μm) at 30 °C were conducted to assess the effect of pH (2–5) and the impact of the bacteria. Nickel extraction from pentlandite and pyrrhotite during bioleaching at pH 2 and 3 was generally good (49–86% after 3 weeks), and cobalt extraction tracked nickel extraction over most conditions. All six ores showed a similar response to a change in pH; an increase in pH from 2 to 3 resulted in approximately the same nickel and cobalt extraction (within statistical error), and a statistically significant reduction in sulphuric acid consumption, dissolved iron, and magnesium extraction.     

Calcination of low-grade laterite for concentration of Ni by magnetic separation

With the continuous depletion of high-grade nickel ores such as millerite and niccolite, nickeliferous laterites have become the major source for the production of nickel metal. However, only 42% of the world’s production of nickel comes from laterites, since the concentration of Ni is relatively low (ca. 2 wt.%). In addition, other metals, such as magnesium, iron and silicon can be found in laterite, which make the concentration of nickel even more difficult.In this study, a low-grade nickeliferous laterite ore was first calcinated and then processed by using a wet magnetic separator in order to recover nickel. Since, the ore contains both Ni and Fe, the calcination of laterite is effective in altering the crystalline structure of Fe species and therefore its magnetic properties, which in turn enable the selective concentration of nickel by magnetic separation that is an easy and environmentally-friendly technique. The experimental results have indicated the importance of carefully controlling: (1) the calcination temperature; (2) the pulp density and (3) applied magnetic field strength. The main finding of this work was that magnetic separation is effective in recovering 48% of nickel from laterite, increasing the Ni grade in the recovered product from 1.5% to 2.9%, when prior to the separation the ore was calcinated at 500 °C for 1 h.     

The use of diffuse reflectance spectroscopy for the characterization of iron ores

The aim of this work was to develop a diffuse reflectance methodology for quantifying minerals in powdered iron ores, which is a key quality control requirement for these materials. Selected samples ranging widely in their concentrations of hematite (as specularite and martite), goethite, magnetite, and quartz were collected in mines from the Iron Quadrangle, Minas Gerais State, and also in the Carajás region, Pará State, Brazil. A chemometric analysis based on the concentrations of the different minerals as determined with a combination of conventional methods (chemical analysis, X-ray diffraction, Mössbauer spectroscopy, light-reflected microscopy, and magnetic susceptibility) and the principal components derived from the diffuse reflectance spectra in the visible range was performed. Principal component regression analysis provided successful calibration for the concentrations of goethite (r² = 0.94; standard error of validation (SEv) = 4.2%) and hematite (r² = 0.89; SEv = 7.4%), in addition to good estimates for quartz (r² = 0.83; SEv = 7.4%), specularite (r² = 0.80; SEv = 11.6%), and martite (r² = 0.78; SEv = 10.6%). Our results suggest that diffuse reflectance spectroscopy is a promising tool for the simultaneous determination of minerals in iron ores within a few minutes only.     

Selective pressure leaching of Fe (II)-rich limonitic laterite ores from Indonesia using nitric acid

The selective extraction of nickel and cobalt over iron from an Indonesian limonitic laterite was investigated using nitric acid pressure leaching (NAPL). The mineralogical analysis showed that the major minerals were goethite and magnetite, and the content of the divalent iron was as high as 7.06%. Nickel and cobalt were mainly distributed in these two minerals; however, the distribution was non-uniform. A series experiments were conducted to examine the basic parameters and propose the optimal conditions for the extraction. When the ore was treated via HPAL under the optimal condition, the extracted nickel and cobalt were less than 75%, and the iron concentration in the leach liquor was over 12.5 g/L. By contrast, over 85% of nickel and cobalt were extracted and about 1.8 g/L iron was achieved using NAPL. The loss of nickel and cobalt can be mainly attributed to the undissolved magnetite and manganese minerals. The leaching process of NAPL is a dissolution–oxidation–precipitation mechanism, and in this process nitric acid acts as both a lixiviant and an oxidant. The formation of hematite results in a low iron concentration in the leach liquor without oxygen injected. Meanwhile, the oxidation and the precipitation of dissolved divalent iron results in a calculated savings in acid consumption of about 120 kg nitric acid per ton of ore can be obtained, which is equal to over 93 kg of sulfuric acid per ton of ore. Moreover, lower residual acid (20 g/L nitric acid) is a significant advantage of NAPL. The iron residues had a high iron content (>56 wt%) with no sulfur, making it suitable as raw materials for ironmaking.     

Infrared thermography: An approach for iron ore gradation

There are insufficient high-grade iron ores currently being mined to meet world demand for steel production. In order to meet raw material demand in India, lower grade ores with high alumina contents are being crushed and beneficiated, mainly by gravity techniques requiring water. However, the scarcity of water in the mining areas warrants the development of some dry gradation techniques for iron ores so that the inferior ore specimens can be rejected in order to improve the grade of the concentrate. The present gradation of ores by mineralogical/chemical methods is time-consuming and cumbersome. This paper presents an Infrared (IR) thermography-based non-invasive technique for the faster gradation of iron ores, taking into account the variation in thermal absorptivity of the ore constituents. Iron ore samples from the Joda, Noamundi and Barsua mines, with Fe contents ranging from 52 to 67 wt.%, were selected and crushed to around 10 mm particle size. The crushed iron ores were uniformly heated using a microwave oven, for a time period sufficient to create a difference between the ore particles in the extent of their respective infrared emissions. The thermal images of the heated particles were captured by IR thermography and the peak temperature of each ore particle was obtained from the thermal profile. A computer program was developed for ore gradation based on the peak temperature of each ore particle which corresponds to its iron content. A threshold was selected through chemical verification of the ores and accordingly the ores were categorized as high-, medium- and low-grade.     

Ligand-promoted dissolution of serpentine in ultramafic nickel ores

The ligands catechol, citrate, EDTA, oxalate and tiron were investigated for their ability to improve the dissolution of serpentine in ultramafic nickel ores at neutral to alkaline pH to enhance mineral carbon sequestration. It is desirable to leach magnesium from serpentine in ores at neutral to alkaline pH so that leaching and carbonation can be conducted at the same pH value, and ultimately so that the reagent requirements of mineral carbon sequestration can be reduced. Both solution modeling and experimental work were conducted. The solution modeling revealed that each of the ligands studied is able to enhance the solubility of magnesium at the desired pH, with the order of effectiveness being EDTA > tiron > citrate > catechol > oxalate. Experimentally it was shown that the ligands studied could improve both the total amount and rate of magnesium leaching from ultramafic nickel ores. The order of ligand effectiveness based on the experimental work for the Pipe ore was EDTA ⩾ tiron > catechol > oxalate ⩾ citrate. For the OK ore, the order was tiron > EDTA = catechol > oxalate > citrate. Overall, the ligands studied in this work, particularly EDTA, tiron, and catechol, appear promising for enhancing the dissolution of serpentine in ultramafic nickel ores at neutral to alkaline pH.     

Recovery of zinc and cadmium from industrial waste by leaching/cementation

Metallic zinc production from sulfide zinc ore is comprised by the stages of ore concentration, roasting, leaching, liquor purification, electrolysis and melting. During the leaching stage with sulfuric acid, other metals present in the ore in addition to zinc are also leached. The sulfuric liquor obtained in the leaching step is purified through impurities cementation. This step produces a residue with a high content of zinc, cadmium and copper, in addition to lead, cobalt and nickel. This paper describes the study of selective dissolution of zinc and cadmium present in the residue, followed by the segregation of those metals by cementation. The actual sulfuric solution, depleted from the electrolysis stage of metallic zinc production, was used as leaching agent. Once the leaching process variables were optimized, a liquor containing 141 g/L Zn, 53 g/L Cd, 0.002 g/L Cu, 0.01 g/L Co and 0.003 g/L Ni was obtained from a residue containing 30 wt.% Zn, 26 wt.% Cd, 7 wt.% Cu, 0.35 wt.% Co and 0.32 wt.% Ni. The residue mass reduction exceeded 80 wt.%. Cementation studies investigated the influence of temperature, reaction time, zinc concentration in feeding solution, pH of feeding solution and metallic zinc excess. After that such variables were optimized, more than 99.9% of cadmium present in liquor was recovered in the form of metallic cadmium with 97 wt.% purity. A filtrate (ZnSO₄ solution) containing 150 g/L Zn and 0.005 g/L Cd capable of feeding the electrolysis zinc stage was also obtained.     

Direct extraction of nickel and iron from laterite ores using the carbonyl process

The carbonyl method of refining nickel and iron was invented more than 100 years ago and has been used for refining of nickel commercially. CVMR® developed the process of direct extraction of nickel and iron from laterite ores as metal carbonyls which in turn produced pure nickel and iron metals.CVMR®’s carbonyl technology has been applied to several types of limonite and saprolite ores containing other metals such as copper, cobalt and PGE. The process consists of reducing the ore with hydrogen, extracting of iron and nickel in the form of volatile metal carbonyls, separating the metal carbonyls and producing high purity nickel and iron metals; and production of copper, cobalt and PGE concentrate by gravity or magnetic separation. Economic evaluation of this process shows significant increase in cash flow. The CVMR® process does not produce liquid waste and does not require tailing dumps. This makes CVMR®’s process attractive for projects in areas that are environmentally sensitive, or have a high level of rainfall.     

Agglomeration and column leaching behaviour of goethitic and saprolitic nickel laterite ores

Processability of complex, low-grade nickel (Ni) laterite ores via heap leaching is very limited due to some intractable geotechnical and hydrological challenges such as poor heap porosity/permeability and structural stability. This work presents some investigations on laboratory batch drum agglomeration and continuous column leaching behaviour of saprolitic (SAP) and goethitic (G) Ni laterite ores as part of the quest for an effective ore pre-treatment process for enhanced heap leaching. As a focus, the effect of ore mineralogy/chemistry on the agglomeration and column leaching behaviour of −2 mm (crushed from −15 mm run-of-mine) G and SAP Ni laterite ores was examined. To produce ∼5–40 mm agglomerates in <15 min, the SAP ore required a higher H₂SO₄ (30 wt.%) binder dosage compared with the G ore, although both ores displayed substantially similar, coalescence-controlled agglomeration mechanism. The resulting G agglomerates were more robust than the SAP ones based upon their compressive strength and acidic solution soak test measurements. However, over 100 days of continuous column leaching, the structural stability of the SAP agglomerate bed was slightly greater than that of G agglomerates, reflecting a lesser slump of the former. The pregnant leach solution analysis revealed greater Ni/Co extraction rates from the SAP than the G agglomerates. Whilst the total mass of acid consumed per ton dry ore processed was greater for the SAP ore, the total kg acid per kg Ni extracted was markedly lower. Incongruent leaching of gangue minerals’ constituent elements (e.g., Fe, Mn, Mg, Al and Si) occurred and contributed significantly to the overall acid consumption. The findings show the relevance of agglomeration and column leaching tests for providing useful information for plant designing and optimization of Ni laterite heap leaching operations.     

铁矿石碳热还原过程中铅锌脱除及含铁物相的演变规律

针对含铅0.39%、含锌0.30%的铁矿,采用碳热还原脱除铅锌杂质,利用X射线衍射、扫描电子显微镜及能谱分析等检测手段考察了铁矿还原焙烧过程的反应行为及物相演变规律。结果表明,该铁矿中铅主要以氧化铅和铅铁矾形式存在,锌主要以氧化锌形式存在; 升高焙烧温度及延长焙烧时间均有利于铅锌脱除; 在1 200 ℃下焙烧60 min时,铁矿中铅和锌脱除率均在90%以上。含铅锌铁矿在碳热还原焙烧过程中会生成中间产物铁橄榄石,并最终转变为金属铁和游离的氧化硅固溶体。还原焙烧产物经磁场强度80 kA/m弱磁选可获得铁品位91.91%和铁回收率84.78%的铁精矿,且铁精矿中铅和锌含量分别为0.01%和0.03%,可作为电炉炼钢原料使用。     

Adsorption of copper sulphate on PGM-bearing ores and its influence on froth stability and flotation kinetics

Copper sulphate is used as an activator in the flotation of base metal sulphides as it promotes the interaction of collector molecules with mineral surfaces. It has been used as an activator in certain platinum group mineral (PGM) flotation operations in South Africa although the mechanisms by which improvements in flotation performance are achieved are not well understood. Some investigations have suggested these changes in flotation performance are due to changes in the froth phase rather than activation of minerals by true flotation in the pulp zone. In the present study, the effect of copper sulphate on froth stability was investigated on two PGM containing ores, namely Merensky and UG2 (Upper Group 2) ores from the Bushveld Complex of South Africa. Froth stability tests were conducted using a non-overflowing froth stability column. Zeta potential tests and ethylenediaminetetraacetic acid (EDTA) tests were used to confirm the adsorption of reagents onto pure minerals commonly found in the two ores. The results of full-scale UG2 concentrator on/off copper sulphate tests are also presented. The UG2 ore showed a substantial decrease in froth stability in the order of reagent addition: no reagents > copper > xanthate > copper + xanthate, while Merensky ore showed a slight decrease. It was shown through zeta potential measurements that copper species were to be found on plagioclase, chromite, talc and pyrrhotite surfaces and through EDTA extraction that this copper was in the form of almost equal amounts of Cu(OH)₂ and chemically reacted copper ions on the Merensky and UG2 ore surfaces. In certain cases, the presence of copper sulphate and xanthate substantially increased the recovery, and therefore the implied hydrophobicity, of pure minerals in a frothless microflotation device. It was, therefore, proposed that increases in hydrophobicity beyond an optimum contact angle for froth stability, were the cause of instabilities in the froth phase and these were found to impact grade and recovery in a full-scale concentrator. Differences in the extent of froth phase effects between the different ores can be attributed to differences in mineralogy.     

Effect of amine and starch dosages on the reverse cationic flotation of an iron ore

In iron ore concentration, reverse cationic flotation of quartz has been successfully employed for particles below 150 μm previously deslimed. Amine and starch are used, respectively, as quartz collector and iron oxides depressant. Understanding the mechanisms of reagents interaction is relevant to improve the separation selectivity, especially for high amine dosages. The term clathrate was used to explain this interaction, meaning a molecular compound in which molecules of one species occupy the empty spaces in the lattice of the other species, resulting in the depression of hydrophobic minerals. Laboratory scale experiments were carried out with itabirite iron ore in three different size ranges. The clathrate formation between molecules of amine and starch may explain the increase of SiO₂ content in the concentrates of the coarse size range (−150 + 45 μm) due to an increase in amine dosage.     

Possibilities for Co(III) dissolution from an oxidized ore through simultaneous bioleaching of pyrite

Solubilisation of Co(III) from a heterogenite met in copper cobaltiferous oxide ore has been realized through reductive leaching using ferrous iron generated via bio-oxidation of pyrite. Biotic and abiotic experiments at various pulp densities and redox potentials have been performed and results compared. Cobalt leaching at elevated redox potential is possible, offering cost reduction benefits due to reduced consumption of ferrous iron. At elevated potential of 625 mV, however, the initial rate of cobalt leaching has been found as 115 mg/(g ore)⋅(24 h), lower than the rate of 865 mg/(g ore)⋅(24 h) registered at 505 mV. Less stochiometric amount of ferrous iron was required when cobalt leaching was coupled to pyrite bioleaching, with 75% of cobalt recovered for 12 h at the optimally found conditions. It could be inferred that the Fe³⁺–Fe²⁺ cycle exists and is efficiently maintained through bacterial presence in the studied system.