3.1. Characterization of Drill Core Samples
The physical and chemical characterization of the HT showed variations occurring in the different layers of the drill cores, which revealed the variability at different depths of this repository location. In terms of moisture content, some of the upper tailing layers had higher moisture content than lower layers (
Figure 6). This moisture variation results from varying grain size distribution in the tailing layers. It was observed that layers with significantly higher fraction of finer particles (1_2-6 and 1_2-11) had higher moisture content than the layers with relatively coarser particles. In order to illustrate particle size variation with depths, the determined D
80 values for the various layers of the drill core were plotted against the depth, as shown in
Figure 7.
Layer 1_2-2, located between 87 and 144 cm depth, was significantly coarser (D
80 = 681 µm). On the other hand, layers 1_2-6 and 1_2-11 at 303 and 483 cm depth were significantly finer, with D
80 values of 259 and 227 µm, respectively. This variation in grain size distribution with depth is the first indication of the presence of different geometallurgical units, which may have different metallurgical performances [
8], and thus need to be characterized. The variability would likely have an effect on the choice of methods to be considered for reprocessing, such as further grinding to liberate minerals of interest that may be locked up in the coarse particles, as observed with optical microscopy.
The −600 to +297 µm and −297 to +149 µm were the dominating particle size fractions, except the two layers with higher proportions of finer fractions mentioned above (
Figure 8). This variation in particle size is indicative of varying process parameters, such as grinding size and/or changes in the mineralogy during the production period of 1936 to 1963.
The chemical composition of the various particle size fractions was also determined. The major components were SiO2, Al2O3, CaO, and Fe2O3, while W, Cu, S, Sn, Zn, Be, and Bi were the main trace elements. From the elemental concentrations in the particle size fractions, weighted average concentrations in each layer were determined and subsequently calculated for the entire drill core.
Figure 9 illustrates elemental concentrations and mass distributions for layer 1_2-4 as an example, which is the layer with the highest mass distribution of tailings in this drill core (compare
Figure 3). Considering the drill core as a whole, W and Cu, as the main metals of interest, were observed to have high concentrations of 2329 mg/kg and 1427 mg/kg, respectively, in the fine (<75 µm) particle size fraction. All elements were high in the dominating particle size fractions mentioned earlier. Therefore, for purposes of reprocessing these HT, additional steps for ensuring sufficient mineral liberation need to be considered for such coarse tailings particles.
The chemical composition of the drill core was also varying with depth. The highest WO
3 concentration was 0.22 wt.% in layer 1_2-2 (between 87 and 144 cm), as shown in
Figure 10. Since this was the layer with the coarsest tailings particles, it means that during the concentration processes shown in
Figure 1a, most of the scheelite mineral particles should have been lost to these tailings as non-liberated particles. The highest Cu concentration was 1147 mg/kg in layer 1_2-8, at a much deeper depth between 313 and 365 cm. However, it is important to know the mass distribution of metals in the layers, because the concentrations may be high but the actual quantities would be small when the total metal content in the drill core is considered. Therefore, the mass distribution for each element in the layers was calculated as a percentage of the total elemental content in the drill core.
Figure 11 shows the WO
3 and Cu mass distributions in the drill core. It shows that much of the WO
3 and Cu was contained at the depth between 174 and 295 cm, which contain 32 wt.% of the total WO
3 and 29 wt.% of the total Cu.
Since these tailings have been stored in this repository for a long period, sulphur depletion due to oxidation was expected, with the depletion decreasing from top to bottom—even though this would also depend on other factors such as particle size and initial quantity. Therefore, both the concentration and mass distribution of sulphur in each layer was analyzed. The highest S concentration was also in layer 1_2-2 at 1.85%, as shown in
Figure 12 (left). However, with regard to mass distribution, it was observed that much of the S was contained in layer 1_2-4 at the depth between 174 and 295 cm, with 28 wt.% of the total S.
Observing the S depletion trend for the drill core according to
Figure 12 (right), there would be three possible main deposition and oxidation periods where S is seen to have a significant stepwise increase in depth; the first being for the depth 296–483 cm, second 145–295 cm, and third 0–144 cm. Based on the alteration index of minerals and pH/EC in the tailings, the upper section—oxidized acidic zone (pH < 5.5)—showed that pyrrhotite was completely replaced by HFOs, calcite depleted, and occasional yellow rims around scheelite grains [
17]. These trends also show possible mineralogical variations with regard to sulphur content over the production period of 1936 to 1963.
Table 4 summarizes the concentrations of the seven main elements in the 11 layers and the weighted averages in the drill core. The variations in the elemental concentrations and distributions at different depths would have an effect on the grades and recoveries of the concentrates that may be produced from the reprocessing of the HT. Consequently, process parameters would need to be varied depending on what depth the tailings are obtained. For instance, in order to recover much of the W and Cu, the tailings at deeper depths (below 174 cm) must be treated. Blending of tailings before reprocessing may also be necessary for process and product optimization.
Under the optical microscope, scheelite particles were observed in normal light and with a blue filter (
Figure 13). The scheelite particles in the polished samples were both fine and coarse, as well as liberated and non-liberated, meaning that mineral liberation analysis would be essential in order to develop effective separation methods. Some particles showed rims, which were identified as hydrous ferric oxides [
20]; hence, the recovery of such scheelite mineral particles may be hindered in processes like flotation where reagents need to have contact with the mineral particle, and also in magnetic separation where scheelite would end up being pulled to the paramagnetic fraction. Hence, pre-treatment methods such as scrubbing may be required. The main minerals in which the main elements W, Cu, S, Sn, Zn, Be, Bi, and F were contained were scheelite, chalcopyrite, pyrrhotite, cassiterite, danalite (both Zn and Be), bismuthinite, and fluorspar, respectively [
20,
24]
The particle and compositional characterization of these historical tailings based on the analyzed drill core revealed significant vertical variations in the repository. For this HT repository, the prediction of metallurgical performance based on geometallurgical units would be defined first from the drill core layers perspective. This approach of using the repository layers as samples for metallurgical test work helps to avert technical errors that arise from the use of composite samples, which may not sufficiently represent the repository [
11]. The metallurgical performance of the individual drill core layers provides an effective way of assessing their eventual effect on composite samples. For repository locations with several different layers, such as 1_2, blending of the layers would be inevitable because of the insufficient thickness of some of the layers. The variability would unavoidably affect the choice of the reprocessing methods and the elemental/mineralogical composition of the products [
8,
11].
3.2. Processing Properties
The Yxsjöberg tailings are fine particles (−600 to +38 µm) with high-density valuable minerals such as scheelite (6.01 g/cm
3) and bismuthinite (7 g/cm
3), and low-density gangue minerals such as quartz (2.62 g/cm
3) and calcite (2.71 g/cm
3) [
25]. Therefore, enhanced gravity separation where a centrifugal force is applied to enhance the differential settling velocities between heavy and light particles (−80,000 to +10 µm) would be appropriate for these HT; thus, a Knelson concentrator was used [
19]. The tailings have a high sulphur concentration, and the sulphur occurs mainly in pyrrhotite, which is weakly-to-strongly magnetic. Magnetite, a ferromagnetic mineral, is also present, and as such, LIMS and HIMS were also considered as plausible processing methods. The magnetic separation would enhance the separation of minerals like scheelite and chalcopyrite from pyrrhotite, which is the main Fe–sulphide mineral in the tailings responsible for AMD [
20].
3.2.1. Magnetic Separation Tests
The two samples from layers 1_2-4 and 1_2-8 were used separately in magnetic separation tests. The three products from the magnetic separation, namely, ferromagnetic, paramagnetic, and non-magnetic fractions, had significant visual (color) differences, with the non-magnetic fraction dominated by the orange-, brown-, and white-colored minerals. Based on the knowledge of minerals known to be present in the samples and using EMC, the light-colored minerals in the non-magnetic fraction were identified to be mainly albite, fluorspar, calcite, scheelite, and biotite.
The desired outcome of the magnetic separation test was to have the valuable minerals (scheelite, fluorspar, and chalcopyrite) in the non-magnetic fraction of the HIMS, but the mass recovery to this fraction for both samples was very low, with the highest amount being only 9.2 wt.% from layer 1_2-8. From
Figure 14 and
Table 5, it is observed that much of the W and Cu ended up in the paramagnetic fraction for both layers; meaning the separation of scheelite and chalcopyrite was not achieved as desired. This could be due to the insufficient liberation of scheelite and chalcopyrite minerals from pyrrhotite and/or to liberated small particles being entrapped and entrained by the paramagnetic particles. The insufficient liberation of scheelite was confirmed by the higher W recovery of 83.6 wt.% in the paramagnetic fraction, from the less coarse layer 1_2-8 which had more W retained in the –600 to +297 µm fraction as compared to layer 1_2-4. The paramagnetic fraction had the highest mass recoveries with 87.3 wt.% for layer 1_2-4 and 85.8 wt.% for layer 1_2-8, and based on EMC, approximately 40 wt.% and 30 wt.% of pyrrhotite was in this fraction, respectively, indicating that a high amount of pyrrhotite was ferromagnetic. Sulphur was mostly recovered in the ferromagnetic and paramagnetic fractions with only 1.0 wt.% in the non-magnetic fraction of layer 1_2-8, meaning that pyrrhotite, the main Fe–sulphide mineral in the tailings and responsible for AMD, was retained in the desired magnetic fractions of the LIMS and HIMS. For both layers, the mass recovered by the LIMS was very low, with the highest amount being only 5.0 wt.% from layer 1_2-8. This could indicate a low amount of the ferromagnetic mineral magnetite in the tailings. But considering the high recovery of Fe
2O
3 in the paramagnetic fraction, it could also mean that a larger amount of magnetite was locked up with pyrrhotite and/or other non-paramagnetic minerals, such as cassiterite, danalite, bismuthinite, fluorspar, calcite, and quartz.
Assessing the particle sizes in the products in relation to what was in the feed, it was observed that 97 wt.% of the −600 to +297 µm particles, being the most abundant in these HT, were distributed to the paramagnetic fraction (
Figure 15). This confirms that the minerals of interest, like scheelite, need to be further liberated from this particle size fraction in order to improve mineral separation by magnetic separation.
3.2.2. Gravity Separation Tests
The two layers (1_2-1 and 1_2-8) of drill core 1_2 were individually used for gravity separation tests with the Knelson concentrator, with two separation cycles for each sample. For comparison in metallurgical performance, these samples were quite similar in terms of mass distribution in the core, but slightly different in chemical composition, grain size, and exposure to weathering. The recovery of scheelite in each cycle product was assessed by the amount of W (
Figure 16). It was observed that recovery of W in the concentrates was decreasing with an increasing number of separation cycles, with the highest recovery being 60.6 wt.% in concentrate 1 of layer 1_2-1. The decrease in W recovery with increasing number of separation cycles was due to the decreasing amount of dense and coarse particles that contain W. In this regard, comparison between the two layers showed a higher W recovery in concentrate 1 for layer 1_2-1, which was coarser with higher W content than layer 1_2-8. This can be seen in the products mass balance calculations given in
Table 6.
Even though the recovery of scheelite was significantly favorable with this enhanced gravity separation, it is important to look at the selectivity, i.e., its separation from the other minerals. From
Figure 16, it is observed that in both samples, other than W, at least 30 wt.% of each main element was also recovered to the concentrate fraction. This means that quartz (2.62 g/cm
3), fluorspar (3.13 g/cm
3), calcite (2.71 g/cm
3), chalcopyrite (4.19 g/cm
3), pyrrhotite (4.61 g/cm
3), and danalite (3.43 g/cm
3) were not fully separated from scheelite, despite having much lower mineral densities than scheelite (6.01 g/cm
3). Considering the coarseness of the particles in these samples, the insufficient mineral separation would be attributed to the insufficient liberation of the mineral particles. With regard to the tailing fraction, W distribution was considerably high at 32.7 wt.% in layer 1_2-8, which is an indication of relatively high scheelite distribution into the fine fractions. The scheelite particles may be liberated in the fine fractions but would still end up in the tailings because of the preferential concentration of coarser particles due to their combined effect of size and density, which are important in enhanced gravity separation.
Using products distribution in various particle sizes to further assess the concentration process (
Figure 17), it was confirmed that the dense coarser (–600 to +297 µm) particles were distributed more to concentrate 1 than 2, while the dense finer (<75 µm) particles were higher in concentrate 2 than 1. Therefore, in order to minimize W losses in the fines to the tailing fraction, the particle size range of the feed material to the Knelson concentrator must be narrow; otherwise, many concentration cycles would be needed to optimize the recovery. In this initial metallurgical test work, the feed particle size range was −600 to +38 µm; however, for subsequent tests, division of this size range into narrower ones such as −600 to +297 µm, −297 to +149 µm, −149 to +75 µm, and −75 to +38 µm should be considered in order to improve the recovery and separation efficiency [
19].
3.2.3. Proposed Process Flowsheet
Based on the preliminary results of the physical and chemical characterization, as well as processing tests, a process flowsheet is proposed (
Figure 18). Given the high deportment of mass and economic elements into the coarser particle size fractions (−600 to +149 µm), classification to separate the coarser tailings from the finer ones would be the first step. The threshold for such a classification should be <75 µm because, from optical microscopy and SEM observations, scheelite particles in this size fraction are liberated. Also, during gravity separation, much of the <75 µm ended up in the tailing product; hence, there is a need for prior separation. Basically, scheelite recovery from this particle size fraction would be best through froth flotation [
26]. For the tailing fraction >75 µm, the issue of mineral liberation is evident; hence, regrinding will be required. However, more studies need to be done to determine the optimal mineral liberation size for scheelite in the tailings so that generation of ultrafine particles is minimized. With sufficient mineral liberation, having magnetic separation before gravity separation enhances the separation of magnetite (in addition to pyrrhotite) from scheelite; magnetite density (5.15 g/cm
3) is close to that of scheelite (6.01 g/cm
3) [
25]; hence, their separation with gravity method may not be efficient. Therefore, gravity separation would be applied to the non-magnetic product of magnetic separation.