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Cobalt (Co) mine production primarily originates from the sediment-hosted copper (Cu) deposits of the Democratic Republic of Congo (DRC). These deposits usually consist of three ore zones with a supergene oxide ore blanket overlying a transition zone which grades into a sulphide zone at depth. Each of these zones display a mineral assemblage with varying gangue mineralogy and, most importantly, a distinct state of oxidation of the mineralisation. This has direct implications for Cu and Co extraction during mineral processing as it dictates which processing method is to be used (i.e., leaching vs. flotation) and affects the performance of these. To optimise resource efficiency, reduce technical risks and environmental impacts, comprehensive understanding of variation of ore mineralogy and texture in the deposit is essential. By defining geometallurgical ore types according to their inferred metallurgical behaviour, this information can serve to classify the resources and improve resource management. To obtain insight into the spatial distribution of mineral grades, it is necessary to develop techniques that have the potential to measure rapidly and, preferably, within the mine at relatively low-cost. In this study, the application of portable Fourier transformed infrared (FTIR) spectroscopy is investigated to measure the mineralogy of drill core samples. A set of samples from a sediment-hosted Cu-Co deposit in DRC was selected to test this approach. Results were validated using automated mineralogy (QEMSCAN). Prediction of gangue and target mineral grades from the FTIR spectra was achieved through partial least squares regression (PLS-R) combined with competitive adaptive reweighted sampling (CARS). It is shown that the modal mineralogy obtained from FTIR can be used to classify the ore according to type of mineralisation and gangue mineralogy into geometallurgical ore types. This classification supports selection of a suitable processing route and is likely to affect the overall process performance.

The transition to a low-carbon economy will be material-intensive. Production of these materials (from mining to manufacturing) incurs environmental costs that vary widely, depending on the geology, mineralogy, extraction routes, type of product, purity of product, background system or manufacturing infrastructure. Understanding the impacts of the raw materials underpinning the low-carbon economy is essential for eliminating any dissonance between the benefits of renewable technologies and the impacts associated with the production of the raw materials. In this Review, we propose an integrated life cycle assessment and geometallurgical approach to optimize the technical performance and reduce the environmental impact of raw material extraction. Life cycle assessments are an effective way of understanding the system-wide impacts associated with material production, from ore in the ground to a refined chemical product ready to be used in advanced technologies such as batteries. In the geometallurgy approach, geologists select exploration targets with resource characteristics that lend themselves to lower environmental impacts, often considering factors throughout the exploration and development process. Combining these two approaches allows for more accurate and dynamic optimization of technology materials resource efficiency, based on in situ ore properties and process simulations. By applying these approaches at the development phase of projects, a future low-carbon economy can be achieved that is built from ingredients with a lower environmental impact. The transition to renewable energy will require a notable quantity of technology metals and materials; however, production of technology materials causes substantial environmental damage. This Review discusses the raw material extraction approaches that optimize technical performance and reduce environmental impact.

Funding Information: The authors would like to thank Boliden Kevitsa mine for industrial support and access to the data. Prof. Kari Heiskanen is acknowledged for his advices, comments, and suggestions. The authors thank Prof. K.H. Esbensen for the very valuable comments provided as part of the pre-examination of TMK Le's PhD manuscript. This research did not receive any specific grant from funding agencies in the public, commercial, or not-for-profit sectors. Publisher Copyright: © 2021 The Authors Due to the rising pressure from governmental regulations, the water recycling rate has increased significantly in mining operations over the last decades, resulting in a high variation of process water quality, which could potentially impact the plant performance. The current effort to assess water quality in mining is shifting from managing water to fulfill environmental regulations (focus on the effluents) to controlling water quality to maintain the operating performance (focus on the water within the process). However, minimal effort has been made to design a dedicated sampling procedure for process water. This study investigates the use of multivariate variography and principal component analysis (PCA) for improving the process water sampling procedure at the Kevitsa Mine, Finland. The aim is to design a sampling procedure for evaluating water quality using two different types of datasets and illustrating the impact of the dataset structure on the sampling design. The results showed that the common spot sampling procedure generated a very high sampling error and was not the best practice for process water. The weekly sampling frequency used at the mine site, suitable for fulfilling environmental regulations was too low to capture the process water variation. Therefore, it is not recommended to use environmental water datasets for operating control purposes. The multivariate variographic analysis revealed the hidden cyclic variation through its ability to summarize the time variations and the correlation between multiple variables that were not visible through the classical univariate variogram approach. However, the number of increments recommended by the global multivariogram became impractically high. Hence, an alternative approach combining PCA to the mutivariogram was used to filter noise from the data and keep the relevant information. This study highlights the benefits of using multivariate variography to improve water sampling procedures in the mining industry and to reduce both operational and environmental risks associated with water quality variability. Thus, this method has the potential to be used in worldwide mining operations as a standard procedure for sampling water to provide reliable results.

More than half of the global cobalt supply from primary sources is currently produced in the Democratic Republic of Congo (DRC) from ores containing copper-cobalt oxides or copper-cobalt sulphides. Where oxide and sulphide cobalt-bearing copper minerals occur together, efficient recovery of copper and cobalt is known to be extremely difficult. This study investigates the flotation behaviour of a mixed oxide-sulphide ore where copper is hosted in sulphides phases such as bornite, chalcopyrite and chalcocite, and oxide phases such as malachite and to a lesser extent chrysocolla. Three cobalt-bearing oxide minerals are observed in the mixed ore, i.e. heterogenite, kolwezite and cupro-asbolane, while carrollite is the only cobalt-bearing sulphide mineral. A two-stage rougher-scavenger flotation process is used in which sulphides are extracted from the ore first. In the second stage, oxides are activated using controlled potential sulphidisation followed by their recovery. Tests are performed with a range of collectors, including xanthate, phosphorodithioate, dithiophosphate, thiocarbamate, and a blend type. A dithiophosphate collector proved to the most successful, achieving recovery of 94% of the carrollite, more than 90% of the copper sulphides, and 70% of the copper oxide minerals. The recovery of the cobalt oxides was less successful, with recovery of roughly half the kolwezite and only 20% of the heterogenite and cupro-asbolane. Despite the generally promising recoveries, the selectivity of the flotation process is relatively low with all the concentrates containing a significant amount of carbonate and silicate minerals. This suggests that a number of improvements require further investigation, notably the application of hydroxamate collectors, depressants and reverse flotation.

As part of a study investigating the influence of mineralogical variability in a sediment hosted copper&ndash, cobalt deposit in the Democratic Republic of Congo on flotation performance, the flotation of nine sulphide ore samples was investigated through laboratory batch kinetics tests and quantitative mineral analyses. Using a range of ore samples from the same deposit the influence of mineralogy on flotation performance was studied. Characterisation of the samples through QEMSCAN showed that bornite, chalcopyrite, chalcocite and carrollite are the main copper-bearing sulphide minerals while carrollite is the only cobalt-bearing mineral. Mineralogical characteristics were averaged per sample to allow for a quantitative correlation with flotation performance parameters. Equilibrium recoveries, rate constants and final grades of the samples were correlated to the feed mineralogy through Multiple Linear Regression (MLR). Target sulphide minerals content and particle size, magnesiochlorite content, carrollite liberation and association of the copper and cobalt minerals with magnesiochlorite and dolomite were used to predict flotation performance. Leave One Out Cross Validation (LOOCV) revealed that the final copper and cobalt grades are predicted with an R2 of 0.80 and 0.93 and Root Mean Square Error of Cross Validation (RMSECV) of 4.41% and 1.34%. The recovery of cobalt and copper with time can be predicted with an R2 of 0.94 for both and an overall test error of 4.70% and 5.14%. Overall, it was shown that quantitative understanding of changes in mineralogy allows for prediction of changes in flotation performance.

Because of their low grades in critical metals such as Light Rare Earth Elements (LREE) or Sn, rare-metal granites are not considered as economic for metal recovery but, when altered, they are often exploited for their industrial minerals. The St Austell rare-metal granite is well known for its world-class kaolin deposits which formed as a result of the extensive weathering and alteration of the underlying granite. The St Austell granite body is composed of several granite components, each having its own accessory minerals assemblage. As a result of the kaolinisation process, some metal-bearing accessory minerals of the granite, such as monazite (LREE) or cassiterite (Sn), are partially liberated from the gangue which allow their pre-concentration in the micaceous residue which is considered as a potential source for critical metals recovery. Similarities with other similar rare-metal granites suggest that topaz granite is the most prospective for disseminated magmatic Sn-Nb-Ta-REE mineralization. However, comparison of the potentiality of 3 granite types i.e., biotite, topaz and tourmaline granites suggest that biotite granites is actually the most prospective due to higher degree of kaolinisation of the biotite granite which favour pre-concentration of its accessory mineral in the micaceous residue. In order to develop a geometallurgical framework for extraction of kaolin and metals from the selected granite component, a field sampling campaign is performed. Core samples are processed in the laboratory using a characterisation program that mimics the full-scale kaolin refining route. Two main products are recovered through this program, viz. MR180 (−180 +53 µm) and P5 (−5 µm), which correspond to a fine micaceous residue and a fine kaolin product respectively. These products are both analysed routinely for major and minor trace elements by XRF and yields are recorded to indicate process performance. A selected number of MR180 samples are also being characterised in terms of particle size by laser light scattering, geochemistry by ICP-MS, and mineralogy by QEMSCAN®. Comparison of characterisation results of MR180 samples and corresponding industrial residue samples shows a good correlation, suggesting that sample analyses are representative for the in-situ deposit and the processing behaviour. Monazite is found to be either fully liberated or fully locked from one sample to the other. Next, pilot-scale gravity concentration tests are performed on micaceous residue samples. Characterisation of the processing products shows that monazite lost in the tailings is mostly locked within tourmaline or micas and is fine grained. Then, predictive regression models for spiral separation performance in terms of recovery, product grade and enrichment as a function of the feed grade are developed for MR180 LREE grade data. Finally, kaolin resources can be classified using quantitative indicators such as yield of the P5 product and the iron oxides content which provides insight into the kaolin quality in terms of whiteness. This geometallurgical classification can be used to delineate zones of interest within the deposit. Although kaolin quality and recovery primarily inform extraction planning, zones which are also of interest for metal recovery can be identified. The proposed model predicts whether the expected LREE grade and recovery satisfy the by-product requirements.

International audience; The Tabuaco tungsten project, located in Northern Portugal, consists of two skarn layers, namely "Main" and "Lower" skarns, which display significantly different mineralogical and geochemical features. Both skarns contain fine-grained disseminated scheelite, but the Lower-skam gangue is dominated by silicates whereas the Main-skarn gangue contains calcium-bearing minerals including fluorite, apatite, and vesuvianite, in close association with scheelite. Preliminary feasibility studies showed that direct separation of calcium-bearing minerals by flotation with fatty acids, which are environmentally benign, was unsuccessful due to their similar surface properties. Several routes for each skarn type were proposed with a strong focus on the Main skarn that displays the most complex mineralogical associations. Enhanced gravity separation (Falcon concentrator), high intensity magnetic separation, and flotation with fatty acids were investigated for the Main Skarn level with the objective of producing a marketable scheelite concentrate. Flotation was thoroughly investigated in terms of depressants, producing optimally a concentrate assaying 9.2% WO3 at 87.9% WO3 recovery from a similar to 1.1% WO3 feed, by means of a 1:1 ratio of sodium carbonate and sodium silicate. Fluorite is the most problematic mineral in flotation with fatty acids and, consequently, new fatty-acids-based collector formulations have been developed to improve the separation contrast between scheelite and fluorite, producing a concentrate assaying 14.1% WO3 with 77.1% WO3 recovery for the fully-optimised conditions. Furthermore, the performance of two types of Falcon bowls (SB and UF) were investigated and optimised for gangue minerals rejection. The Falcon concentrator was used as a desliming and pre-concentrating apparatus: one stage of Falcon SB allowed rejection of 84% of the total amount of fluorite, more than 95% of the slimes, and 85 wt% of the mass. Consequently, it improved significantly the performance of the fully-optimised scheelite flotation as the final concentrate, after one Falcon SB stage and 4 flotation stages, assayed 62.9% WO3 with 59.4% WO3 recovery, which constituted a marketable scheelite concentrate. Both the Falcon SB parameters and the number of fully-optimised flotation stages can be optimised to maximise either the WO3 grade or the WO3 recovery, regarding the product specifications. Finally, the high intensity magnetic separation allowed rejection of around 45 wt% of the total yield prior to the milling stage with only 6.0% WO3 losses, which would decrease significantly the energy consumed during the milling stage.

Scheelite (CaWO4) is one of the main raw material for the production of tungsten. It is usually encountered in skarn deposits where it is commonly associated with other calcium minerals as fluorite, apatite, and calcium silicates. Worldwide, scheelite is upgraded to the chemical grades by direct flotation, but the separation efficiency remains limited due to similar flotation behaviors of scheelite and gangue minerals with fatty acid. The only solutions used to overcome this issue involve high energy consumption or ecotoxic reagents. In the present study, a novel method based on the use of a centrifugal Falcon concentrator was investigated to perform an efficient elimination of gangue minerals and fine particles as well as an acceptable scheelite recovery enabling a decrease of the flotation reagents consumption. The performances of the two types of laboratory Falcon bowls, Falcon UltraFine (UF) and Falcon Semi-Batch (SB), were modeled using the design of experiments (DoE) methodology, which allowed to determine the best operating parameters for each bowl. The separation performances were mainly affected by the rotary speed and the pulp density for the Falcon UF and by the rotary speed and the fluidization pressure for the Falcon SB. Due to the fluidization pressure, the Falcon SB exhibited higher gangue minerals rejection with slightly lower recoveries than the Falcon UF. Overall, the optimized Falcon SB test allowed to reach 71.6%, 22.6%, 17.2%, and 12.6% for scheelite, calcium salts, dense calcium silicates, and light non-calcic silicates respectively while the desliming efficiency reached 98.8%. For comparison purposes, a classical hydrocyclone allowed to attain 89.1%, 89.3%, 79.5%, and 76.5% for scheelite, calcium salts, dense calcium silicates, and light non-calcic silicates respectively while the desliming efficiency reached 52.0%. Theses results can be used reliably to assess the separation performances of an industrial Falcon C which can be regarded, along with Falcon SB, as a sustainable and efficient gangue rejection method for complex W skarn ore, which allows the use of environmentally friendly reagents during downstream flotation stages.

Centrifugal gravity separators, such as the Falcon UltraFine (UF) concentrator, are the most common gravity concentration techniques used for fine particles processing. Hence, understanding the kinetics and separation mechanisms at play within these separators is of paramount interest. Recent research yielded a predictive physical model for the Falcon UF which however does not explain some results obtained with industrial ores. The Falcon UF kinetics have been investigated through the processing of fine-grained ores from the Altenberg tin deposit (Germany), the Tabuaco tungsten deposit (Portugal), a synthetic iron ore as well as results from previous studies on kaolin residues. Results have shown an evolution of Falcon UF performance with time/feed mass in contradiction with the stationary separation hypothesis on which the physical model was based. In terms of Falcon UF separation timing, four phases can be distinguished. First, upon initial feeding of the bowl, particles are trapped or rejected depending on their settling velocity. It yields a relatively ineffective selection according to density so that only ultrafine particles are ejected from the bowl, resulting in the quick growth of the concentrate bed. When the bed reaches a critical size, recovery and enrichment continue to increase through selective resuspension phenomenon that favours the concentration of dense particles and the ejection of larger particles. This way, the bed builds up while the content of concentrate bed surface evolves until resuspension balances the stream of dense material reaching the bed and recovery drops. The evolution of partition curves over time confirmed the low recovery of ultrafine particles during the whole operation but also showed a decrease of coarse particles recovery with time. It suggests that the second separation mechanism is less sensible to particle size compared to the first one and that size even has a negative impact on recovery. Furthermore, erosion figures in furrows are observed in the concentrate bed which may play locally an active role in the separation. These observations suggest that two separation mechanisms are at play. Firstly, differential particles settling within the flowing film which is already accounted for in the existing physical model. Secondly, resuspension of particles from the concentrate bed by the action of a lift force acting preferentially on coarse particles deposited at the surface of the bed and resulting in the rejection of coarser and lower-density particles. The addition of a lift force component to the existing model is discussed and a resuspension criterion is proposed as a guidance of the physics involved in this second separation mechanism. Future developments will require a dynamic model which would need to integrate the evolution of the concentrate bed content over time.

This paper, which is the second in a series of three, follows a previous study that showed that Light Rare Earth Elements (LREE), hosted in monazite and rare-metals (Sn, Nb, W), hosted in cassiterite, rutile and wolframite, respectively, are pre-concentrated in the micaceous residue of a kaolin plant. However, the low grade and fine size range of the metal-bearing minerals make their recovery difficult using unique gravity concentration methods. An enhanced gravity concentration flowsheet was tested, which began by screening the residue into three size fractions: +180 μm, 180–53 μm and −53 μm. The 180–53 μm fraction was treated with a spiral as a roughing unit and then a shaking table as a cleaning unit, whereas the fine −53 μm processing is addressed in the third publication. The design of experiments (DOE) methodology was used to set the best operating parameters. The best heavy mineral recoveries for the spiral concentrate were obtained at a low wash water flow rate and low pulp density. The spiral scavenging tests showed that up to 80% of the heavy minerals can be recovered. However, the mineralogical analysis showed that after three passes, middlings began to be recovered, suggesting that the scavenging operations should not be further continued. Furthermore, up to 70% monazite could be recovered in a single pass. Shaking table testing on the spiral concentrates showed that heavy mineral concentrates with up to 98.5% of heavy minerals can be recovered. However, the concentrate grades were still low for most metals (Sn, Nb and W), with the exception of LREEs, which reached up to 1.6% LREE in the shaking table concentrate. Given the reasonable overall recoveries, the main reason for the low concentrate grades could have been the very low grade of the feed material, which was two to three times less enriched in metals than the average micaceous residues grades. This suggests that the cut-off grade for those metals should be above 100 ppm and points to the critical aspect of metal grade variability in the micaceous residue as a major issue if potential industrial applications are to be considered.

This is the last in a series of papers addressing the by-product potential of Light Rare-Earth Elements (LREE, La, Ce, Nd) and rare-metals (Sn, Nb, W) within the St Austell kaolin deposits (UK). Previous results have shown that heavy minerals are pre-concentrated in the micaceous residue which displays relatively high LREE content (around 170 ppm) hosted by monazite. A combined gravity/flotation concentration flowsheet, starting with classification of the residue into 3 size fractions (+150 μm, 150–53 μm and −53 μm) was developed. This study presents the results of the clay-rich −53 μm size fraction beneficiation by flotation using sodium oleate and hydroxamate as collectors and compares them to results obtained with a Falcon Ultra-Fine (UF) concentrator. The effect of dispersant addition during de-sliming operations on flotation efficiency was also investigated. The LREE grade in the floated product obtained with sodium oleate increased from 3370 ppm to 5350 ppm and recovery from 30% to 80% with dispersant which is attributed to clay coating. Saturation tests performed on the Falcon UF show a positive effect of desliming on separation performance which produces higher LREE grades and recoveries. Comparison of flotation performance and Falcon UF concentration of monazite to previous results obtained with the Falcon Semi-Batch (SB) gravity concentrator on the un-screened raw residue clearly shows that flotation of the fine residue is more efficient than the Falcon UF and Falcon SB. Indeed the Falcon UF produced high recoveries (87–95%) but with almost no enrichment (1.3–1.4 ratios) whereas the Falcon SB concentration of un-screened residue produced higher grades (up to 0.6% LREE) but with low recoveries (17–22%). This study has shown that LREE can be recovered with simple two-stage operations, suggesting that the micaceous residue stream (and the associated tailings dams) could be considered as a potential resource for these metals.

Spiral concentrators are one of the most common gravity processing methods extensively used for the concentration of mineral based on their density, particle size and shape. As for every gravity concentration technique, particle size plays a major role in the separation mechanisms within the spiral. Thus, in addition to classical performance indices, such as grade or recovery, partition curves are also a valuable criterion for quantifying the separation efficiency. In this paper, we study and model the influence of wash water additions and pulp density on heavy and gangue mineral partition curves using kaolin residue enriched in heavy minerals and via a design of experiment (DOE). The results show that coarse particle size recovery curves are affected by a systematic decrease, which is mainly impacted by wash water additions. Partition curve modelling through particle size distribution and DOE regression model fitting allow a better understanding of the effect of wash water on the aforementioned phenomenon for each mineral fraction. Gangue minerals are more affected by this phenomenon, which has been interpreted as a result of the Bagnold effect and secondary flows within the spiral. A decrease in the coarse particle recovery represents a significant source of losses, and a better understanding of this phenomenon is necessary to avoid these losses.

The classical application of theory of sampling (TOS) is univariate. However, most practical situations address multi-analyte issues, in which the common belief is that one should focus a variographic study on the analyte with the most heterogeneous distribution. This paper introduces a multivariogram approach to process sampling and compares it with the classical univariate and multivariate approaches of variograms as applied to principal component analysis (PCA) scores. A case study of low-grade kaolin residue sampling for metallurgical testing is used to illustrate this methodology. A total of eight important properties are analysed, including chemical analytes, size distribution properties and pulp density. The results show that the classical univariate approach can underestimate the global sampling error if the sampling protocol is designed by using only the highest variance property. Variograms that are calculated from PCA scores highlight distinct spatial patterns through variable grouping in a reduced number of variograms. Multivariograms can be used to summarise time variations in multiple analytes and highlight the multivariate time autocorrelation aspects of these analytes. However, the resulting sampling variance is very high, and an alternative approach that applies multivariograms to PCA scores, filters noise from the data, and only keeps relevant data information, which reduces the global sampling variance, is proposed. This case study illustrates the usefulness of multivariate approaches to help multivariate variographic modelling in mineral processing and in many other fields within science and industry that deal with multi-analyte sampling issues.

The vast majority of current global production of cobalt from primary sources originates from extraction of Cu-Co sediment-hosed deposits in the Democratic Republic of Congo. With the progressive depletion of oxidic supergene mineralisation, many operations are progressively starting to extract sulphide ore at depth. Ore in the transition zone between oxides and sulphides usually hosts both oxide and sulphide mineralisation which is extremely difficult to process by conventional flotation. These mixed ores are complex to process by flotation alone because carbonate minerals, such as dolomite and magnesite, may represent a significant proportion of the gangue. Hence, mixed ores are usually processed through flotation of sulphides followed by oxide flotation using controlled potential sulphidisation (CPS) and further processing via acid leaching. The main challenges associated with carbonate-rich mixed ores are the low recovery of cobalt, mostly due to cobalt present in oxide minerals, the floatability of the carbonate gangue minerals, which can make up to 50% of the concentrate, and technical difficulties with regards to industrial application of CPS. This study investigates alternative options to the classic two-step CPS flotation route, including direct reverse flotation of the carbonate gangue at neutral or acid pH, as well as a two-step combined sulphide-reverse flotation route. The latter, including a sulphide flotation stage at neutral pH, followed by an acidic reverse flotation stage at pH 4.5–5, increased the overall copper and cobalt recoveries. The best performance is achieved with reverse flotation, using Na-Oleate and a mixture of phosphoric and sulphuric acid (ratio of 4:1). Overall Cu-Co recoveries of 93.5% and 85.1% respectively were achieved. It is postulated that phosphoric acid acts as a depressant of the Cu-Co oxide minerals through surface passivation while enabling selective flotation of magnesite and dolomite. Further investigation into the effect of phosphoric acid on flotation of these minerals is required. Overall, the results suggest that, following flotation of sulphides, an acidic reverse flotation stage increases the global Cu-Co recovery while producing a concentrate which is suitable for further processing by flotation or acid leaching. Hence, the suggested combined approach could be a viable alternative to aforementioned CPS stage in Cu-Co oxide-sulphide mixed ores flotation plants, especially in operations where the mixed ore contains significant proportions of carbonate gangue minerals.

International audience; This paper analyses output streams of a kaolin plant to select the most valuable waste stream for metal recovery. The results indicate that rare metals (Sn, Nb, W) and Light Rare Earth Elements (LREEs) are pre-concentrated in the micaceous residue which contains a relatively high LREE (La, Ce, Nd) concentration of around 170 ppm. This residue have been characterised using a two-stage gravity pre-concentration process involving a Falcon concentrator and a shaking table to obtain concentrates suitable for mineralogical study. A statistical analysis, combined with qualitative mineralogy, highlights the main mineralogical associations and identifies monazite, cassiterite, Nb-rutile and wolframite as the main metal-bearing minerals. Estimation formulas for Nb and LREE using only the TiO2 and SiO2 content of the wastes, obtained by binary and multiple linear regression, show very accurate results suggesting a possible on-line application. This will provide tools to select the most valuable waste streams or the associated tailing dams (using historical data) for metals recovery. These selected wastes will be used for metallurgical test works to assess the potential for by-product recovery of these metals.

In the theory of sampling, variograms have proven to be a powerful tool to characterise the heterogeneity of 1-dimensional lots. Yet its definition and application in sampling for mineral processing have always been limited to one variable, typically ore grade. However this definition is not adapted to sampling for mineral processing where samples contain multiple properties of interest, i.e. variables, such as multiple element grades, grain size, etc. For such cases, the multivariable variogram, originally developed for spatial data analysis, can be used to summarise time variation of multiple variables (e.g. ore characteristics which are important for the process) and highlights the multivariate time auto-correlation of these variables. A case study of low-grade kaolin residue sampling for gravity processing shows that the multivariogram summarises the overall variability and highlights a periodic phenomenon when all variables are taken into account. This example illustrates the potential of the multivariable variogram compared to the classical approach.

Cobalt (Co) is a transition metal featuring unique physical properties which makes its use critical for many high-tech applications such as high strength materials, magnets and most importantly, rechargeable batteries. The bulk of world cobalt output usually arises as a by-product of extracting other metals, mostly nickel (Ni) and copper (Cu), from a wide variety of deposit types mostly Cu-Co sediment-hosted deposits, but also Ni-Co laterites, Ni-Cu-Co sulphides or hydrothermal and volcanogenic deposits. Significant differences in ore properties (geochemistry, mineralogy, alteration and physical properties) exist between cobalt-containing deposits, as well as within a single deposit, which can host a range of ore types. Variability of cobalt ores makes it challenging to develop a single extraction or treatment process that will be able to accommodate all geometallurgical variation. Overall, there is a lack of fundamental knowledge on cobalt minerals and their processability. The recovery efficiency for cobalt is generally low, in particular for processes involving flotation and smelting, leading to significant cobalt losses to mine tailings or smelter slags. This paper starts by reviewing the main geometallurgical properties of cobalt ores, with a particular focus on ore mineralogy which exerts a significant control over ore processing behaviour and cobalt extraction, such as the oxidation state, i.e. oxide or sulphides which drives the selection of the processing route (leaching vs flotation), and the associated gangue mineralogy, which can affect acid consumption during leaching or flotation performance. The main processing routes and associated specific geometallurgical aspects of each deposit type are presented. The paper concludes on the future cobalt prospects, in terms of primary and secondary resources, cobalt processing and sustainable cobalt sourcing for which further research is needed.