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Beneficiation studies of a complex REE ore
Minerals Engineering 71 (2015) 55–64
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Beneficiation studies of a complex REE ore Xiaosheng Yang a,⇑, Jacqueline V. Satur b, Kenzo Sanematsu b, Jukka Laukkanen a, Tuula Saastamoinen a a b
Mineral Processing Laboratory, Geological Survey of Finland, Outokumpu, Finland Mineral Resources Research Group, Institute for Geo-Resources and Environment, National Institute of Advanced Industrial Science and Technology, Tsukuba, Japan
a r t i c l e
i n f o
Article history: Received 2 May 2014 Accepted 7 October 2014
Keywords: Rare earth minerals Flotation Modal mineralogy Mineral recovery
a b s t r a c t The beneficiations of a complex rare earth element (REE) ore containing REE minerals of carbonates, silicates and oxides and other REE bearing minerals (e.g. zircon and pyrochlore etc.) by flotation and high gradient magnetic separation (HGMS) were investigated. The recoveries of the different group of minerals in the flotation concentrates and their flotation kinetics were analyzed through the mineralogical studies by MLA. By flotation nearly 90% of La and Ce and 45% of Y were recovered in the final cumulative concentrate with the mass percentage of 11.5%. The mineralogical studies indicated that over 95% of carbonates were recovered in the final cumulative concentrate in which the recoveries of three major REE minerals bastnaesite, parisite and synchysite reached over 96%. The recoveries of silicate REE minerals such as cerite(Ce) and allanite were quite high and reached 86% and 65%, respectively. The recoveries of Nb mineral pyrochlore and Zr minerals zircon and elpidite were very low. HGMS was significantly less selective in recovering La and Ce but was effective in recovering Y. The flow sheet of rougher–cleaner flotation and the flow sheet of rougher flotation–cleaner HGMS were compared. The flow sheet of rougher flotation– cleaner HGMS was showed to be more effective in recovery of Y. Ó 2014 Elsevier Ltd. All rights reserved.
1. Introduction The rare earth (RE) minerals are numerous and can be classified into six classes based on the arrangement of anionic groups such as carbonate minerals bastnaesite and parisite, phosphate minerals monazite and xenotime, and silicate mineral allanite. Meanwhile, some minerals such as Nb-mineral pyrochlore could contain substantial amounts of REEs. But about 95% of the world rare earth resources occur in bastnaesite, monazite, and xenotime (Miyawaki and Nakai, 1996) and by 2002 over 80% of the world rare earth oxide (REO) production was from bastnaesite and 5% from monazite (Gupta and Krishnamurth, 2005). Meanwhile, about 10% of the world REO production from China’s weathered crust elution rare earth deposit or called ion adsorption rare earth ore which occupies over 80% of the world total heavy RE reserve (Chi and Tian, 2007). As the demands of RE materials in the world are increasing (U.S. Geological Survey, 2002) more REE minerals deposits will be in exploitation. For example, preliminary beneficiation tests have been carried out at Kvanefjeld REE deposit containing steenstrupine as the main REE mineral in Greenland and at Norra Kärr REE deposit containing eudialyte as the main REE mineral in Sweden. ⇑ Corresponding author at: Tutkijankatu 1, 83500 Outokumpu, Finland. Tel.: +358 50 348 6063; fax: +358 13 557 557. E-mail address: Jason.yang@gtk.fi (X. Yang). http://dx.doi.org/10.1016/j.mineng.2014.10.005 0892-6875/Ó 2014 Elsevier Ltd. All rights reserved.
Currently, flotation is still the most important beneficiation method used to recover bastnaesite, monazite, xenotime, parisite and other RE minerals in the world by using different hydroxamic acids and fatty acids as the collectors. The main RE minerals at Bayan Obo REE-Nb-Fe deposit, the world largest REE deposit, are bastnaesite and monazite. In the flotation circuit naphthyl hydroxamic acid is used as the collector of bastnaesite and monazite at low alkaline condition (pH9) and temperature 35–45 °C with the addition of sodium silicate as the depressant of silicates. Bastnaesite and monazite separation is realized by further flotation using phthalic acid C6H4(CO2H)2 or benzoic acid C6H5COOH as the collector of bastnaesite and alum as the depressant of monazite (Yu, 2000). At the Mountain Pass deposit, another major REE deposit in the world, the ore contains bastnaesite rich in cerium group REEs and monazite associated with barite, calcite, strontianite, chinalco and apatite etc. Flotation is used with six different conditioning treatments at high solid percentage and high temperature 70–90 °C with reagents added step-wise, soda ash as the pH modifier, sodium fluosilicate as the depressant. Various collectors such as fatty acids (oleic) (1982), distilled tall oil (fatty and rosin acids), dicarboxylic acids (1985) and hydroxamates (1988) have been used (Gupta and Krishnamurth, 2005). Additionally, flotation is also used for placer REE deposit ores to separate monazite and xenotime from other minerals.
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 1 The contents of REEs in the ore sample by different chemical analysis methods. Method La (ppm)
Ce (ppm)
XRF 660.0 ICP558.5 OES ICP-MS 636.5
1060.0
1095.0
Pr (ppm)
Nd (ppm)
Sm (ppm)
Eu (ppm)
Sc (ppm)
Gd (ppm)
Tb (ppm)
Dy (ppm)
Ho (ppm)
Er (ppm)
Tm (ppm)
Yb (ppm)
Y TREE (ppm) (ppm)
HR/LR
280.0 278.0 126.5
489.0
87.0
4.6
<20
83.9
10.9
58.3
11.6
34.0
4.6
28.9
330.0
3000.6
‘18.7: 81.3
The bold value means the content of total REEs.
For carbonatite type of REE deposit ores the adsorption mechanisms of hydroxamic acids and fatty acids as flotation collectors on REE minerals have been investigated. The research by Pradip and Fuerstenau (1991) indicated that the chemical adsorption occurs between fatty acid collectors and REE minerals for carbonatite type of REE ores at high temperature with the addition of soda ash and lignin sulfonate as modifiers. Pavez et al. (1996) concluded that sodium oleate chemisorbs onto the surface of bastnaesite and the adsorption onto the surface of monazite depends on pH. In the pH range above the isoelectric point the chemisorptions are possible. Otherwise, a physical adsorption occurs. However, for octyl hydroxamate chemical adsorptions occur on both minerals. Assis et al. (1996) reported that the selectivity of minerals by flotation with hydroxamates depends on a balance between the solubility of the minerals and the stability of the complex formed between the cation in the lattice and hydroxamate with the cation of the lattice. A favourable condition for the selective utilisation of hydroxamates as collectors is visualised if the mineral to be floated is the most soluble in the system and the chelate formed between the cation in the lattice and hydroxamate is the most stable. For placer deposit ores the flotation mechanisms by using sodium oleate and hydroxamate as the collectors also have been studied. Surface properties and flotation behaviour of xenotime using sodium oleate as a collector were investigated by Cheng et al. (1994). It was found that the adsorption process on xenotime surface is strongly pH dependent. As sodium oleate used as collector for flotation of monazite and xenotime, the point of zero charge (pzc) of monazite and xenotime occurs at pH 5.3 and pH 3.0, and the maximum floatability of both minerals occurs at pH > 7 where both are chemisorbed and xenotime is slightly more floatable than monazite (Cheng et al., 1993). When sodium oleate and hydroxamates were used as collectors for monazite, sodium metasilicate and sodium sulphide as depressants for zircon (ZrSiO4) and rutile (TiO2), the best selectivity condition (effective depression of zircon and rutile and little effect on monazite) was achieved with either oleate or commercial hydroxamate as collector at pH10. The effect of the conditioning temperature was also studied, revealing that, in some instances, a higher temperature enhances the selectivity, but
in other cases the floatability of zircon and rutile is more strongly improved than that of monazite (Pavez and Peres, 1993). As an important beneficiation technique magnetic separation has been applied in REE ore processing to eliminate ferromagnetic minerals such as magnetite and other Fe-oxides prior to further separation steps as well as to separate individual paramagnetic REE minerals (Jordens et al., 2013). Wet low and medium intensity magnetic separators are used in the beneficiation of China’s bastnaesite REE ores to remove Fe-bearing minerals prior to flotation (Yu, 2000; Zhang and Edwards, 2012). Wet high intensity magnetic separators are applied to REE placer deposits to separate paramagnetic monazite and xenotime from non-magnetic heavy minerals such as zircon and rutile (Gupta and Krishnamurth,2005; Jordens et al., 2013; Zhang and Edwards, 2012). However, for the REE ores with fine grain sizes (<100 lm) high gradient magnetic separation (HGMS) has been approved to be a more effective method. In this paper, the beneficiations of a complex REE ore containing REE minerals of carbonates, silicates and oxides and other REE bearing minerals (e.g. zircon and pyrochlore etc.) by flotation and high gradient magnetic separation (HGMS) were investigated. The recoveries of the different group of minerals in the flotation concentrates and their flotation kinetics were analyzed through the mineralogical studies by MLA. The flowsheet of rougher–cleaner flotation and the flowsheet of rougher flotation–cleaner HGMS were compared.
2. Experimental 2.1. Sample and preparation A drill core composite sample around 110 kg from the National Institute of Advanced Industrial Science and Technology (AIST), Japan, was received at GTK Mineral Processing Laboratory in Finland. This is a REE-Zr-Nb ore consists of peralkaline granite intrusions influenced by hydrothermal alteration. The outcropping granites are relatively weathered and unweathered rocks found below 30 m approximately from the surface.
Table 2 Modal mineralogy of the sample.
Sum
Silicate
wt%
Carbonate
wt%
Oxide and other
wt%
K-feldspar Albite Quartz Arfevdsonite Echemannite Gittinsitea Britholitea Elpiditea Zircona
26.61 27.37 33.34 7.57 0.74 0.07 0.12 0.91 0.34 97.07
Parisitea Bastnaesitea Synchysite-Ya Calcite
0.19 0.12 0.02 0.19
Fluorite Rutile Hematite Ilmenite Pyrochlorea Chlorite Unknown
0.51 0.14 0.72 0.05 0.17 0.34 0.48
The bold values mean the sum of percentages of minerals. a REE minerals and REE bearing minerals.
0.52
2.41
100.00
X. Yang et al. / Minerals Engineering 71 (2015) 55–64
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Fig. 1. SEM images of REE minerals, bastnaesite, synchysite, parisite, britholite, pyrochlore and zircon.
The sample was crushed by a jaw and a roll crusher to 1.5 mm and split into 1.5 kg amount of subsamples for flotation and magnetic separation experiments. One of the subsamples was further split into 100 g amount of samples for chemical and mineralogical analyses. 2.2. Mineralogical analysis A Mineral Liberation Analyzer (MLA) was used for mineralogical analysis which consists of the standard modern SEM (FEI Quanta 600) with the energy dispersive X-ray analyzer (EDAX Genesis with two detectors) and the software package developed originally by JKTech (Australia). The modal mineralogy of the samples was analyzed by the XMOD-STD method.
2.4. High gradient magnetic separation (HGMS) A laboratory high gradient magnetic separator (Sala HGMS 1015-20 SCR) was used for the magnetic separation experiments. The matrix type used was 3.5 XRO with the aperture size of 850 lm. The HGMS experiment consisting of one rougher and one scavenger was carried out at the magnetic intensities of 0.05 T and 0.3 T, respectively. Additionally, HGMS was used as the cleaner for the concentration of flotation rougher concentrate. Three times of cleaning by HGMS were applied at the intensities of 2.0, 0.5, and 0.05 T. 3. Results and discussion 3.1. Chemical analysis
2.3. Flotation and reagents Flotation experiments were carried out using an Outotec laboratory flotation machine. The samples of 1.5 kg with the size of 1.5 mm were ground at a laboratory rod mill as the head samples for flotation. The flotation cells with the volumes of 4 and 1.5 liters were used in the rougher and cleaner flotation, respectively. Tap water was used for wet grinding and flotation. The room temperature (20–23 °C) was applied for flotation conditioning. Hydroxamate type of reagent Aero6494, fatty acid type of reagent Aero704 and sulfosuccinamate type of reagent Aero 845 from Cytec and sodium oleate (Na-oleate) powder (P82 wt% fatty acids as oleic acid) manufactured by Sigma–Aldrich were tested as the collectors. Sodium silicate (Na2SiO3) and starch were used as the depressants of silicate minerals during flotation. Aero 6494, Aero 704 and Aero 845 were used without dilution and sodium oleate was diluted into 5 wt% solution for use. The value of pH was adjusted by using soda ash and sulfuric acid.
Chemical analyses of the sample by XRF, ICP-OES and ICP-MS were conducted at Labtium Oy Finland. The contents of rare earth elements (REEs) by different chemical analysis methods are shown in Table 1. The content of total REEs (TREE) by ICP-MS is 3000 ppm or 0.3% in which the content of the heavy REEs (Gd to Y) is 562 ppm or 0.056%. The ratio of the heavy REEs to the light REEs is 18.7– 81.3%. 3.2. Modal mineralogy The data of modal mineralogy of the sample are shown in Table 2. A total of 20 minerals were identified by MLA in which silicates take up 97.1%, carbonates 0.52%, and oxides and other minerals 2.41% in mass. The identified REE minerals and REE-bearing minerals are parisite, bastnaesite, synchysite-(Y), pyrochlore, gittinsite, britholite, elpidite and zircon. REEs are distributed not only in carbonates and oxides (parisite, bastnaesite, synchysite-(Y),
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Table 3 REE flotation performance comparison at the grinding size of P80 = 50 lm and 80 lm.
a
Collector
Grinding size, P80 lm
La %
Rec%
%
Rec%
%
Rec%
%
Rec%
Aero 704, 845a
50 80
0.37 0.41
55.76 57.63
0.61 0.66
56.65 57.98
0.09 0.10
30.56 30.13
1.07 1.16
52.42 53.76
Aero 6494a
50 80
0.12 0.11
52.49 54.60
0.19 0.17
53.50 55.14
0.03 0.03
36.66 38.20
0.34 0.32
50.81 52.53
Ce
Y
REE (La, Ce, Y)
Dosage 200 g/t.
pyrochlore) but also in silicate minerals (gittinsite, britholite, elpidite and zircon). Light REEs are mainly contained in carbonate minerals (parisite, bastnaesite) and silicate mineral (britholite) while Y and heavy REEs are mainly contained in Zr silicate minerals (zircon, elpidite, gittinsite), Nb mineral (pyrochlore) and carbonate mineral (synchysite-(Y)). The SEM images of REE minerals in Fig. 1 show the appearance of bastnaesite, synchysite, parisite, britholite, pyrochlore and zircon in fine grains. 3.3. Determination of flotation conditions 3.3.1. Grinding size The previous mineralogical analyses by AIST in Japan on the grain size of REE minerals in the ore have shown that the reasonable grinding size for the beneficiation would be around 80% passing 50–100 lm. To select a suitable grinding size the flotation tests were carried out at the grinding size of P80 = 50 lm and P80 = 80 lm by using Aero 704, 845 (in the ratio 1:1) and Aero 6494 as the collectors at the dosage of 200 g/t and Na2SiO3 as the depressant at the dosage of 300 g/t. The desliming of 20 lm by the hydrodynamic settling method was applied before flotation. The results showed in Table 3 indicated that the flotation performances of the REE minerals at the two grinding sizes were not obviously different. 3.3.2. Collector type To select a proper collector the flotation tests using Aero 704, 845 (in the ratio of 1:1), Na-oleate and Aero 6494 as the collectors were carried out in the presence of sodium silicate (Na2SiO3) as the depressant. The grinding size was P80 = 50 lm and the temperature during pulp conditioning was at 20–23 °C. When the dosages
of collectors varied in the range of 70–200 g/t the different flotation recoveries of the total REE (La, Ce, Y) were obtained. The curves of the grade versus recovery of REE (La, Ce, Y) in the concentrate for different collectors are shown in Fig. 2. It was demonstrated that higher recovery and grade of REE (La, Ce, Y) were obtained by using Na-oleate as the collector. That is, comparing with Aero 704, 845 and Aero 6494, Na-oleate was more selective for REE flotation for this ore.
3.3.3. Depressant dosage Because of the high content of silicates in the ore the depression of silicate minerals is important for REE flotation. Two levels of 300 g/t and 50 g/t of Na2SiO3 dosage were tested using Na-oleate as the collector. The flotation results were shown in Fig. 3. It is obvious that higher grade and recovery of REE (La, Ce, Y) in flotation concentrates were obtained at dosage 300 g/t than 50 g/t. By the flotation experiments discussed above on the grinding size, collector type and Na2SiO3 dosage the flotation conditions were determined: grinding size P80 = 80 lm, collector Na-oleate at dosage 200–300 g/t and depressant Na2SiO3 at dosage 300 g/t. Starch could be added to improve the depression of Fe-oxides.
3.4. Rougher flotation and results The rougher flotation experiment was carried out at the grinding size of P80 = 80 lm. The particle size distribution of the feed measured by a laser sizer is shown in Table 4. Na-oleate was used as the collector at the dosage of 270 g/t and Na2SiO3 and starch as the silicate depressants at the dosages of 300 g/t and 150 g/t. The flotation pH was natural in 8.8–9.3 and the pulp temperature for conditioning and flotation was at 20–23 °C.
Fig. 2. Effect of collector type on the flotation of REE (La, Ce, Y).
X. Yang et al. / Minerals Engineering 71 (2015) 55–64
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Fig. 3. Effect of Na2SiO3 on the flotation of REE (La, Ce, Y).
Table 4 Particle size distribution of the feed. Particle size, lm Cumulative percentage, vol.%
3.19 10.00
12.40 20.00
38.88 50.00
72.79 75.00
81.18 80.00
103.90 90.00
The flow sheet of the experiment is shown in Fig. 4. The total flotation time was 28 min. Six froth concentrates (concentrate 1– 6) were respectively pulled in the flotation periods of 0–2 min, 2–4 min, 4–6 min, 6–9 min, 9–14 min, and 14–28 min and the tail was the material left in the cell. The flotation products of Conc 1–6 and Tail were filtered, tried, weighed and sent for assay by XRF. The recoveries of elements in the products were calculated based on the weights and assays. The element contents of the feed were recalculated based on the data of products. All the experimental data are shown in Table 5. For analyzing the experimental errors the element contents of the feed by XRF (Analysis feed) are listed in Table 5. The standard deviation (SD) between the calculated feed (X1) and the analysis feed (X2) was determined by
SD ¼
rffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi 1 X ðX i X m Þ2; 21
where Xm is the mean value of X1 and X2. The standard deviations between the calculated feed and the analysis feed on La, Ce and Y are shown in Fig. 5.
The weights of cumulative concentrates and the grades and recoveries of elements in the cumulative concentrates calculated from the experimental data in Table 5 are shown in Table 6. The mass percentage of the final cumulative concentrate Conc (1,2,3,4,5,6) was 11.5%. Nearly 90% of La and Ce were recovered but the recovery of Y was only 42%. The relationships of grade versus recovery of La, Ce and Y in the cumulative concentrates are shown in Fig. 6.
3.5. Mineralogical studies of flotation products In order to investigate the flotation recoveries and kinetics of different minerals, the mineralogical analyses by MLA were conducted on the concentrates 1, 3 and 5 and the tail. The modal mineralogy of these three concentrates and the tail were obtained. Based on the trend lines of mineral contents in flotation products versus flotation time the modal mineralogy of concentrates 2, 4 and 6 were achieved by calculation. The modal mineralogy of all the concentrates and the tail are shown in Table 7. Comparing the data in Table 7 with the modal mineralogy of ore in Table 2 all the major identified minerals and REE minerals are consistent. But some silicate minerals such as smectite, garnet, and allanite, and carbonate mineral siderite, were not identified in the analysis of ore. Among them allanite is a REE mineral. It is revealed in Table 7 that as the increase of flotation time the grades of carbonate and oxide minerals decrease and the grades of silicate minerals increase in the flotation concentrates from concentrate 1 to concentrate 6.
Fig. 4. Flow sheet of rougher flotation.
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 5 Assays of the flotation products and the calculated feed. Product
Weight g
Conc 1 21.10 Conc 2 20.30 Conc 3 16.30 Conc 4 19.10 Conc 5 31.10 Conc 6 64.50 Tail 1327.80 Calc’d feed (X1) 1500.20 Analysis feed (X2) Standard deviation (SD)
La
Ce
Y
Fe
SiO2
ZrO2
C
%
ppm
Rec%
ppm
Rec%
ppm
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
1.41 1.35 1.09 1.27 2.07 4.30 88.51 100.00
19,400 8600 5100 2810 1460 642 80 609 660 36.00
44.80 19.11 9.10 5.87 4.97 4.53 11.62 100.00
31,000 14,200 8500 4680 2480 1093 140 1002 1060 40.73
43.50 19.17 9.21 5.94 5.13 4.69 12.36 100.00
2190 1630 1350 1010 750 482 180 276 280 2.85
11.16 7.99 5.32 4.66 5.63 7.51 57.73 100.00
3.00 4.52 5.57 6.08 6.67 6.39 3.25 3.53 3.16 0.26
1.20 1.73 1.71 2.19 3.92 7.78 81.47 100.00
22.30 37.00 46.00 56.60 64.20 69.24 75.30 72.99 73.90 0.64
0.43 0.69 0.68 0.99 1.82 4.08 91.31 100.00
0.29 0.45 0.57 0.60 0.63 0.55 0.33 0.35 0.35 0.00
1.17 1.73 1.77 2.18 3.73 6.74 82.67 100.00
2.97 2.04 1.79 1.21 0.82 0.50 0.09 0.22 0.21 0.01
19.10 12.62 8.89 7.05 7.77 9.81 34.76 100.00
had not only high recoveries of over 96% in the final cumulative concentrate but also high flotation kinetics. At the time of 6 min their recoveries reached 81–90%. After the flotation of carbonate REE minerals two silicate REE minerals of cerite(Ce) and allanite were floated at slower kinetics. The recoveries of cerite(Ce) and allanite reached 79% and 48% at the flotation time of 15 min. However, the recoveries of one Nb mineral pyrochlore and two Zr minerals zircon and elpidite were very low and their flotation kinetics were the slowest. Because these minerals are the main bearing minerals of Y and other heavy REEs the low recoveries of them caused the low recoveries of Y and other heavy REEs in concentrates. 3.6. High gradient magnetic separation (HGMS) Fig. 5. Standard deviations between the calculated feed and the analysis feed on La, Ce and Y.
As shown in Table 8 the recoveries of minerals in the cumulative concentrates at different flotation times were calculated based on the modal mineralogy in Table 7 and the mass percentages of flotation products. Over 95% of carbonates were recovered in final cumulative concentrate in which the recoveries of three major REE minerals bastnaesite, parisite, synchysite and synchysite-(Y) reached over 96%. Meanwhile, very high recoveries of calcite and siderite (94% and 92% respectively) were obtained in the final cumulative concentrate. The recovery of fluorite reached as high as 98% but the recoveries of other oxide minerals such as pyrochlore and Fe-oxide were below 30%. Although the recoveries of major silicates minerals quartz, plagioclase, K-feldspar and arfvedsonite were quite low in the range of 7–17% those of silicate REE minerals such as cerite(Ce) and allanite were quite high and reached 86% and 65%. This means the collector Na-oleate was quite selective for the recovery of silicate REE minerals from other major silicates. The recoveries of different REE minerals in cumulative concentrates at different times are shown in Fig. 7. The carbonate group of REE minerals (bastnaesite, parisite, synchysite and synchysite-(Y))
As shown in Fig. 6 the recovery of Y in the rougher flotation concentrate was only 45%. The poor recovery of Y can be explained by the poor recoveries of Y bearing minerals pyrochlore, zircon and elpidite as discussed above. A high gradient magnetic separation (HGMS) experiment was carried out. The matrix type of 3.5 XRO with the aperture size of 850 lm was used. Two stages of rougher and scavenger separations were applied at the intensities of 0.05 T and 0.3 T, respectively. The rougher and scavenger magnetic concentrates were combined as the final concentrate. The recoveries of La, Ce and Y versus concentrate mass for flotation and HGMS are shown in Fig. 8. Comparing to flotation HGMS was significantly less selective for recoveries of La and Ce. Only 74.7% recovery of La and Ce was obtained at a high mass percentage (35.8%). But the recovery of Y reached 67.1%. That is, HGMS could be more selective for the recoveries of pyrochlore, zircon and elpidite than flotation using Na-oleate as the collector. 3.7. Cleaning process and comparison of flow sheets Although the results of HGMS could be improved by optimizing the conditions it is not capable to replace flotation at rougher stage due to low La and Ce recoveries and grades. After rougher flotation
Table 6 Weights of cumulative concentrates, and grades and recoveries of elements in the cumulative concentrates. Cumulative conc
Conc Conc Conc Conc Conc Conc
1 (1,2) (1,2,3) (1,2,3,4) (1,2,3,4,5) (1,2,3,4,5,6)
Weight
La
Ce
Y
Fe
SiO2
ZrO2
C
g
%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
21.10 41.40 57.70 76.80 107.90 172.40
1.41 2.76 3.85 5.12 7.19 11.49
1.94 1.41 1.16 0.94 0.71 0.47
44.80 63.90 73.00 78.87 83.84 88.38
3.10 2.28 1.87 1.52 1.16 0.76
43.50 62.67 71.88 77.82 82.95 87.64
0.22 0.19 0.18 0.16 0.13 0.10
11.16 19.15 24.47 29.13 34.76 42.27
3.00 3.75 4.26 4.71 5.28 5.69
1.20 2.93 4.64 6.83 10.75 18.53
22.30 29.51 34.17 39.75 46.79 55.19
0.43 1.12 1.80 2.79 4.61 8.69
0.29 0.37 0.43 0.47 0.52 0.53
1.17 2.90 4.67 6.86 10.59 17.33
2.97 2.52 2.31 2.04 1.69 1.24
19.10 31.72 40.61 47.66 55.43 65.24
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Fig. 6. Relationships of grade versus recovery of La, Ce and Y in cumulative concentrates.
Table 7 Modal mineralogy of flotation products by MLAa. Mineral Silicate
Carbonate
Oxide and other
Product Flotation time, min
Conc 1 0–2
Conc 2 2–4
Conc 3 4–6
Conc 4 6–9
Conc 5 9–14
Conc 6 14–28
Tail >28
Feed 33.34 27.37 26.61 7.57 0.07
Quartz Plagioclase (Albite) K-felspar Arfvedsonite Gittinsite Smectite Garnet Cerite(Ce) (Britholite) Allanite Zircon Elpidite Other Total
7.48 6.01 4.15 4.07
15.81 12.21 7.26 7.81
18.80 15.00 10.75 11.07
23.45 17.06 12.41 14.16
26.33 20.59 15.80 17.46
32.03 22.51 23.56 16.88
35.54 23.06 26.91 8.68
1.92 0.68 1.68 0.21 0.58 0.15 0.02 26.95
4.22 0.91 1.34 0.19 0.89 0.24
5.68 1.00 1.02 0.17 1.10 0.33 0.03 64.95
6.54 0.93 0.70 0.19 1.20 0.35
6.18 0.61 0.29 0.18 1.17 0.42 0.04 89.07
4.72 0.27 0.13 0.11 0.75 0.51
3.64 0.02 0.01 0.01 0.45 0.52 0.01 98.84
Calcite Siderite Bastnaesite(Ce) Parisite Synchysite(Y) Synchysite(Ce) Other Total
16.09 1.43 9.52 4.93 0.39 3.24 0.04 35.63
Fluorite Pyrochlores Fe-oxide Other Total
34.52 0.26 0.94 1.70 37.42 1.94 3.10
50.88 11.22 2.16 4.81 3.39 0.26 1.52
76.98
7.56 2.43 1.65 1.67 0.12 0.45 0.13 14.01
4.47 2.25 0.40 1.25 0.09 0.34
2.12 1.14 0.23 0.28 0.02 0.15 0.10 4.04
9.83 0.30 1.52
26.16
16.86 0.36 1.65 2.18 21.05
3.48 0.26 1.47 1.68 6.89
0.86 1.42
0.51 0.85
0.28 0.47
0.15 0.25
23.37 24.61 0.30 1.25
101.74 0.93 0.49 0.10 0.12 0.01 0.06
0.12 0.34 0.91 0.74 97.07
0.04 0.02 0.01 0.00 0.00 0.00 0.01 0.08
0.52
2.88
0.03 0.08 0.43 0.54 1.08
0.51 0.17 0.72 1.01 2.41
0.06 0.11
0.01 0.01
0.06 0.10
1.71 1.49 0.19 1.20
0.19 0.12 0.19 0.02
REE La Ce a
Flotation conditions: Na-oleate 270 g/t, Na2SiO3 300 g/t, starch 150 g/t, pH 8.8–9.3. Unit of mineralogy: wt%.
the combined concentrate of Conc 1,2,3,4,5,6 was respectively cleaned by flotation and HGMS. The flow sheet is shown in Fig. 9. In the cleaner flotation Na-oleate was used as the collector and three time cleaners were applied. In the cleaner HGMS stage three time cleaners were applied at the intensities of 2.0, 0.5 and 0.05 T and the matrix type was 3.5 XRO with the aperture size of 850 lm. The testwork results for the two flow sheets are shown in Fig. 10. The recoveries of La and Ce were 68–73% and not obviously different for the two flow sheets. The grade of La plus Ce reached
4.2%. But the flow sheet of flotation–HGMS was showed to be more effective for recovery of Y. That is, comparing to the flow sheet of flotation–flotation a higher recovery of Y was obtained by the flow sheet of flotation–HGMS at the same recovery of La and Ce. The assays of the two cleaned concentrates from the flow sheets of flotation–flotation and flotation–HGMS are shown in Table 9. The grades of REE (La, Ce, Y) for the two flow sheets were 4.9% and 4.5%, respectively, at the recoveries of 53.9% and 67.4%. The
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 8 Recoveries of minerals in cumulative concentrates. Conc 1
Quartz Plagioclase K-felspar Arfvedsonite Smectite Garnet Cerite(Ce) Allanite Zircon Elpidite Silicate Calcite Siderite bastnaesite(Ce) Parisite Synchysite(Y) Synchysite(Ce) Carbonate Fluorite Pyrochlores Fe-oxide Oxide and other
Conc 1,2
Conc 1,2,3
Conc 1,2,3,4
Conc 1,2,3,4,5
Conc 1,2,3,4,5,6
Calc’d feed
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
7.48 6.01 4.15 4.07 1.92 0.68 1.68 0.21 0.58 0.15 26.93 16.09 1.43 9.52 4.93 0.39 3.24 35.59 34.52 0.26 0.94 35.72
0.31 0.38 0.23 0.62 0.71 11.14 27.69 10.29 1.62 0.42 0.39 35.37 12.39 55.87 42.32 44.09 56.13 38.53 37.65 3.68 2.51 26.22
11.56 9.05 5.67 5.91 3.05 0.79 1.51 0.20 0.73 0.19 38.67 13.70 1.79 7.21 4.18 0.33 2.40 29.60 29.66 0.28 1.09 31.04
0.93 1.11 0.61 1.77 2.23 25.59 48.83 19.59 4.03 1.06 1.11 59.09 30.42 83.04 70.35 72.85 81.54 62.87 63.48 7.70 5.73 44.70
13.61 10.73 7.11 7.36 3.79 0.85 1.37 0.19 0.83 0.23 46.09 11.97 1.97 5.64 3.47 0.27 1.85 25.16 26.05 0.30 1.25 27.60
1.53 1.84 1.07 3.07 3.86 38.37 61.72 26.10 6.41 1.78 1.84 71.93 46.68 90.52 81.40 83.33 87.63 74.48 77.69 11.55 9.14 55.40
16.05 12.31 8.42 9.05 4.48 0.87 1.21 0.19 0.92 0.26 53.77 10.10 2.04 4.34 2.91 0.23 1.47 21.09 22.01 0.30 1.32 23.63
2.40 2.81 1.69 5.02 6.06 52.31 72.11 34.48 9.47 2.67 2.85 80.83 64.30 92.64 91.10 92.93 92.91 83.11 87.40 15.38 12.82 63.14
19.02 14.69 10.55 11.48 4.97 0.80 0.94 0.19 0.99 0.31 63.93 7.80 1.78 3.15 2.15 0.17 1.09 16.15 16.67 0.29 1.36 18.32
4.00 4.71 2.97 8.94 9.45 67.16 79.10 47.55 14.32 4.38 4.77 87.71 78.80 94.64 94.61 96.69 96.68 89.39 92.99 20.79 18.61 68.78
23.89 17.62 15.42 13.50 4.87 0.60 0.64 0.16 0.90 0.38 77.98 5.23 1.30 2.01 1.39 0.11 0.71 10.75 10.99 0.25 1.30 12.55
8.03 9.03 6.92 16.81 14.81 80.76 85.65 64.75 20.80 8.76 9.29 93.93 91.78 96.51 97.76 100.00 100.00 95.06 97.98 28.86 28.41 75.25
34.20 22.43 25.59 9.23 3.78 0.09 0.09 0.03 0.50 0.50 96.43 0.64 0.16 0.24 0.16 0.01 0.08 1.30 1.29 0.10 0.53 1.92
100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00
The bold values mean the sum of percentages of silicate group minerals, the sum of percentages of carbonate group minerals and the sum of percentages of oxide and other group minerals.
Fig. 7. Recoveries of REE minerals in cumulative concentrates versus flotation time.
Fig. 8. Recoveries of REE (La, Ce) and Y versus concentrate mass for flotation and HGMS.
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Fig. 9. Flow sheet of cleaning.
Fig. 10. La plus Ce and Y grade versus recovery for rougher–cleaner flotation flow sheet and rougher flotation–cleaner HGMS flow sheet.
Table 9 Assays of the concentrates from flow sheets of flotation–flotation and flotation–HGMS. Element
Flotation–Flotation conc Flotation–HGMS conc Feed
La
REE (La, Ce, Y)
Fe
%
Rec%
Ce %
Rec%
Y %
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
1.76 1.59 0.06
59.30 72.84 100.00
2.85 2.60 0.10
60.50 75.07 100.00
0.24 0.27 0.03
17.98 27.92 100.00
4.85 4.45 0.19
53.85 67.40 100.00
2.72 12.48 3.51
1.62 10.10 100.00
11.50 37.06 73.09
0.33 1.44 100.00
1.11 4.76 9.94
0.23 1.36 100.00
0.38 0.86 0.35
2.27 7.02 100.00
3.47 3.19 0.21
35.19 44.08 100.00
grade of REE (La, Ce, Y) in the ore from Table 1 was 0.19%. So that, the enrichment ratios for the two processes were 25.7 and 23.7, respectively. These are quite high values of beneficiation enrichment for this type of complex ore. Table 9 also indicates that the both concentrates had high contents of C and the flotation–HGMS concentrate had a high content of Fe. That is, calcite and Fe-oxides were also concentrated in the cleaned concentrates. To further increase the grade of REE in the concentrates calcite and Fe-oxides should be eliminated. Further flotation by using more selective collectors or properly controlling pH value could be effective to remove calcite and further magnetic separation by adjusting the magnetic intensity or gradient could work to decrease the content of Fe-oxides. Acid leaching (sulfuric or hydrochloric acids) at high concentration after roasting (200–400 °C) could be used for the downstream metallurgical processing of the concentrates. REE carbonate minerals (bastneasite, parisite etc.) should be dissolved in the acids at high recoveries. Due to high content of calcite a pre-leaching with diluted HCl at room temperature may be needed to dissolve calcite and purify the material prior to REE leaching. The leachability of REEs in silicate minerals may be low and the caustic pre-treatments with calcium hydroxide or sodium hydroxide should be needed to transform them into soluble salts prior to leaching.
SiO2
Al2O3
ZrO2
C
4. Conclusions Using Na-oleate as the collector at the dosage of 270 g/t and Na2SiO3 and starch as the silicate depressants at the dosages of 300 g/t and 150 g/t, nearly 90% of La and Ce were recovered but the recovery of Y was 45% in the final cumulative concentrate with the mass percentage of 11.5%. The mineralogical studies indicated that over 95% of carbonates were recovered in the final cumulative concentrate in which the recoveries of three major REE minerals bastnaesite, parisite and synchysite reached over 96%. They showed also high flotation kinetics. At the flotation time of 6 min their recoveries reached 81–90%. Meanwhile, the recoveries of silicate REE minerals such as cerite(Ce) and allanite were quite high and reached 85.6% and 64.7%, respectively. Their flotation kinetics was slower than that of carbonate REE minerals and the recoveries of cerite(Ce) and allanite reached 79.1% and 47.6% at the flotation time of 15 min. However, the recoveries of Nb mineral pyrochlore and Zr minerals zircon and elpidite were very low. The cumulative concentrate of flotation was cleaned by flotation and high gradient magnetic separation HGMS. The grades of REE (La, Ce, Y) for the two flow sheets were 4.9% and 4.5%, respectively, at the recoveries of 53.9% and 67.4%. The enrichment ratios for the two processes were 25.7 and 23.7 for the beneficiation of
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REE (La, Ce, Y). Comparing the results by the two flow sheets a higher recovery of Y was obtained by the flow sheet of flotation– HGMS at the same recovery of La and Ce. References Assis, S.M. et al., 1996. Utilisation of hydroxamates in minerals froth flotation. Miner. Eng. 9 (1), 103–114. Cheng, Ta-Wui, Holtham, P.N., Tran, Tam, 1993. Froth flotation of monazite and xenotime. Miner. Eng. 6 (4), 341–351. Cheng, Ta-Wui et al., 1994. The surface properties and flotation behavior of xenotime. Miner. Eng. 7 (9), 1085–1098. Chi, Ruan, Tian, Jun, 2007. Introduction of weathered crust elutiondeposited rare earth ore. J. Chinese Rare Earth Soc. 25 (6), 641–649 (Chinese). Gupta, C.K., Krishnamurth, N., 2005. Extractive Metallurgy of Rare Earths. CRC Press. Jordens, A., Cheng, Y.P., Waters, K.E., 2013. A review of the beneficiation of rare earth element bearing minerals. Miner. Eng. 41, 97–114.
Miyawaki, R., Nakai, I., 1996. Crystal chemical aspects of rare earth minerals. In: Jones, Adrian P., Wall, Frances, Terry Williams, C. (Eds.), Rare Earth Mineral: Chemistry, origin and ore deposits. Chapman & Hall, pp. 21–37. Pavez, O., Peres, A.E.C., 1993. Effect of sodium metasilicate and sodium sulphide on the flotability of monazite-zircon-rutile with oleate and hydroxamates. Miner. Eng. 6 (1), 69–78. Pavez, O. et al., 1996. Technical note adsorption of oleate and octyl-hydroxamate onto rare-earths minerals. Miner. Eng. 9 (3), 357–366. Pradip, Fuerstenau, D.W., 1991. The role of inorganic and organic reagents in the flotation separation of rare-earth ores. Int. J. Miner. Eng. 32, 1–22. U.S. Geological Survey, 2002. Rare earth elements-critical resources for high technology, USGS Fact Sheet 087-02. Yongfu Yu, 2000. Mineral processing techniques of rare earths and development in China, West China Exploration Engineering, series No.63, pp. 1–4. (Chinese). Zhang, J., Edwards, C., 2012. A review of rare earth mineral processing technology. In: 44th Annual Meeting of the Canadian Mineral Processors. CIM, Ottawa, pp. 79–102.
Contents lists available at ScienceDirect
Minerals Engineering journal homepage: www.elsevier.com/locate/mineng
Beneficiation studies of a complex REE ore Xiaosheng Yang a,⇑, Jacqueline V. Satur b, Kenzo Sanematsu b, Jukka Laukkanen a, Tuula Saastamoinen a a b
Mineral Processing Laboratory, Geological Survey of Finland, Outokumpu, Finland Mineral Resources Research Group, Institute for Geo-Resources and Environment, National Institute of Advanced Industrial Science and Technology, Tsukuba, Japan
a r t i c l e
i n f o
Article history: Received 2 May 2014 Accepted 7 October 2014
Keywords: Rare earth minerals Flotation Modal mineralogy Mineral recovery
a b s t r a c t The beneficiations of a complex rare earth element (REE) ore containing REE minerals of carbonates, silicates and oxides and other REE bearing minerals (e.g. zircon and pyrochlore etc.) by flotation and high gradient magnetic separation (HGMS) were investigated. The recoveries of the different group of minerals in the flotation concentrates and their flotation kinetics were analyzed through the mineralogical studies by MLA. By flotation nearly 90% of La and Ce and 45% of Y were recovered in the final cumulative concentrate with the mass percentage of 11.5%. The mineralogical studies indicated that over 95% of carbonates were recovered in the final cumulative concentrate in which the recoveries of three major REE minerals bastnaesite, parisite and synchysite reached over 96%. The recoveries of silicate REE minerals such as cerite(Ce) and allanite were quite high and reached 86% and 65%, respectively. The recoveries of Nb mineral pyrochlore and Zr minerals zircon and elpidite were very low. HGMS was significantly less selective in recovering La and Ce but was effective in recovering Y. The flow sheet of rougher–cleaner flotation and the flow sheet of rougher flotation–cleaner HGMS were compared. The flow sheet of rougher flotation– cleaner HGMS was showed to be more effective in recovery of Y. Ó 2014 Elsevier Ltd. All rights reserved.
1. Introduction The rare earth (RE) minerals are numerous and can be classified into six classes based on the arrangement of anionic groups such as carbonate minerals bastnaesite and parisite, phosphate minerals monazite and xenotime, and silicate mineral allanite. Meanwhile, some minerals such as Nb-mineral pyrochlore could contain substantial amounts of REEs. But about 95% of the world rare earth resources occur in bastnaesite, monazite, and xenotime (Miyawaki and Nakai, 1996) and by 2002 over 80% of the world rare earth oxide (REO) production was from bastnaesite and 5% from monazite (Gupta and Krishnamurth, 2005). Meanwhile, about 10% of the world REO production from China’s weathered crust elution rare earth deposit or called ion adsorption rare earth ore which occupies over 80% of the world total heavy RE reserve (Chi and Tian, 2007). As the demands of RE materials in the world are increasing (U.S. Geological Survey, 2002) more REE minerals deposits will be in exploitation. For example, preliminary beneficiation tests have been carried out at Kvanefjeld REE deposit containing steenstrupine as the main REE mineral in Greenland and at Norra Kärr REE deposit containing eudialyte as the main REE mineral in Sweden. ⇑ Corresponding author at: Tutkijankatu 1, 83500 Outokumpu, Finland. Tel.: +358 50 348 6063; fax: +358 13 557 557. E-mail address: Jason.yang@gtk.fi (X. Yang). http://dx.doi.org/10.1016/j.mineng.2014.10.005 0892-6875/Ó 2014 Elsevier Ltd. All rights reserved.
Currently, flotation is still the most important beneficiation method used to recover bastnaesite, monazite, xenotime, parisite and other RE minerals in the world by using different hydroxamic acids and fatty acids as the collectors. The main RE minerals at Bayan Obo REE-Nb-Fe deposit, the world largest REE deposit, are bastnaesite and monazite. In the flotation circuit naphthyl hydroxamic acid is used as the collector of bastnaesite and monazite at low alkaline condition (pH9) and temperature 35–45 °C with the addition of sodium silicate as the depressant of silicates. Bastnaesite and monazite separation is realized by further flotation using phthalic acid C6H4(CO2H)2 or benzoic acid C6H5COOH as the collector of bastnaesite and alum as the depressant of monazite (Yu, 2000). At the Mountain Pass deposit, another major REE deposit in the world, the ore contains bastnaesite rich in cerium group REEs and monazite associated with barite, calcite, strontianite, chinalco and apatite etc. Flotation is used with six different conditioning treatments at high solid percentage and high temperature 70–90 °C with reagents added step-wise, soda ash as the pH modifier, sodium fluosilicate as the depressant. Various collectors such as fatty acids (oleic) (1982), distilled tall oil (fatty and rosin acids), dicarboxylic acids (1985) and hydroxamates (1988) have been used (Gupta and Krishnamurth, 2005). Additionally, flotation is also used for placer REE deposit ores to separate monazite and xenotime from other minerals.
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 1 The contents of REEs in the ore sample by different chemical analysis methods. Method La (ppm)
Ce (ppm)
XRF 660.0 ICP558.5 OES ICP-MS 636.5
1060.0
1095.0
Pr (ppm)
Nd (ppm)
Sm (ppm)
Eu (ppm)
Sc (ppm)
Gd (ppm)
Tb (ppm)
Dy (ppm)
Ho (ppm)
Er (ppm)
Tm (ppm)
Yb (ppm)
Y TREE (ppm) (ppm)
HR/LR
280.0 278.0 126.5
489.0
87.0
4.6
<20
83.9
10.9
58.3
11.6
34.0
4.6
28.9
330.0
3000.6
‘18.7: 81.3
The bold value means the content of total REEs.
For carbonatite type of REE deposit ores the adsorption mechanisms of hydroxamic acids and fatty acids as flotation collectors on REE minerals have been investigated. The research by Pradip and Fuerstenau (1991) indicated that the chemical adsorption occurs between fatty acid collectors and REE minerals for carbonatite type of REE ores at high temperature with the addition of soda ash and lignin sulfonate as modifiers. Pavez et al. (1996) concluded that sodium oleate chemisorbs onto the surface of bastnaesite and the adsorption onto the surface of monazite depends on pH. In the pH range above the isoelectric point the chemisorptions are possible. Otherwise, a physical adsorption occurs. However, for octyl hydroxamate chemical adsorptions occur on both minerals. Assis et al. (1996) reported that the selectivity of minerals by flotation with hydroxamates depends on a balance between the solubility of the minerals and the stability of the complex formed between the cation in the lattice and hydroxamate with the cation of the lattice. A favourable condition for the selective utilisation of hydroxamates as collectors is visualised if the mineral to be floated is the most soluble in the system and the chelate formed between the cation in the lattice and hydroxamate is the most stable. For placer deposit ores the flotation mechanisms by using sodium oleate and hydroxamate as the collectors also have been studied. Surface properties and flotation behaviour of xenotime using sodium oleate as a collector were investigated by Cheng et al. (1994). It was found that the adsorption process on xenotime surface is strongly pH dependent. As sodium oleate used as collector for flotation of monazite and xenotime, the point of zero charge (pzc) of monazite and xenotime occurs at pH 5.3 and pH 3.0, and the maximum floatability of both minerals occurs at pH > 7 where both are chemisorbed and xenotime is slightly more floatable than monazite (Cheng et al., 1993). When sodium oleate and hydroxamates were used as collectors for monazite, sodium metasilicate and sodium sulphide as depressants for zircon (ZrSiO4) and rutile (TiO2), the best selectivity condition (effective depression of zircon and rutile and little effect on monazite) was achieved with either oleate or commercial hydroxamate as collector at pH10. The effect of the conditioning temperature was also studied, revealing that, in some instances, a higher temperature enhances the selectivity, but
in other cases the floatability of zircon and rutile is more strongly improved than that of monazite (Pavez and Peres, 1993). As an important beneficiation technique magnetic separation has been applied in REE ore processing to eliminate ferromagnetic minerals such as magnetite and other Fe-oxides prior to further separation steps as well as to separate individual paramagnetic REE minerals (Jordens et al., 2013). Wet low and medium intensity magnetic separators are used in the beneficiation of China’s bastnaesite REE ores to remove Fe-bearing minerals prior to flotation (Yu, 2000; Zhang and Edwards, 2012). Wet high intensity magnetic separators are applied to REE placer deposits to separate paramagnetic monazite and xenotime from non-magnetic heavy minerals such as zircon and rutile (Gupta and Krishnamurth,2005; Jordens et al., 2013; Zhang and Edwards, 2012). However, for the REE ores with fine grain sizes (<100 lm) high gradient magnetic separation (HGMS) has been approved to be a more effective method. In this paper, the beneficiations of a complex REE ore containing REE minerals of carbonates, silicates and oxides and other REE bearing minerals (e.g. zircon and pyrochlore etc.) by flotation and high gradient magnetic separation (HGMS) were investigated. The recoveries of the different group of minerals in the flotation concentrates and their flotation kinetics were analyzed through the mineralogical studies by MLA. The flowsheet of rougher–cleaner flotation and the flowsheet of rougher flotation–cleaner HGMS were compared.
2. Experimental 2.1. Sample and preparation A drill core composite sample around 110 kg from the National Institute of Advanced Industrial Science and Technology (AIST), Japan, was received at GTK Mineral Processing Laboratory in Finland. This is a REE-Zr-Nb ore consists of peralkaline granite intrusions influenced by hydrothermal alteration. The outcropping granites are relatively weathered and unweathered rocks found below 30 m approximately from the surface.
Table 2 Modal mineralogy of the sample.
Sum
Silicate
wt%
Carbonate
wt%
Oxide and other
wt%
K-feldspar Albite Quartz Arfevdsonite Echemannite Gittinsitea Britholitea Elpiditea Zircona
26.61 27.37 33.34 7.57 0.74 0.07 0.12 0.91 0.34 97.07
Parisitea Bastnaesitea Synchysite-Ya Calcite
0.19 0.12 0.02 0.19
Fluorite Rutile Hematite Ilmenite Pyrochlorea Chlorite Unknown
0.51 0.14 0.72 0.05 0.17 0.34 0.48
The bold values mean the sum of percentages of minerals. a REE minerals and REE bearing minerals.
0.52
2.41
100.00
X. Yang et al. / Minerals Engineering 71 (2015) 55–64
57
Fig. 1. SEM images of REE minerals, bastnaesite, synchysite, parisite, britholite, pyrochlore and zircon.
The sample was crushed by a jaw and a roll crusher to 1.5 mm and split into 1.5 kg amount of subsamples for flotation and magnetic separation experiments. One of the subsamples was further split into 100 g amount of samples for chemical and mineralogical analyses. 2.2. Mineralogical analysis A Mineral Liberation Analyzer (MLA) was used for mineralogical analysis which consists of the standard modern SEM (FEI Quanta 600) with the energy dispersive X-ray analyzer (EDAX Genesis with two detectors) and the software package developed originally by JKTech (Australia). The modal mineralogy of the samples was analyzed by the XMOD-STD method.
2.4. High gradient magnetic separation (HGMS) A laboratory high gradient magnetic separator (Sala HGMS 1015-20 SCR) was used for the magnetic separation experiments. The matrix type used was 3.5 XRO with the aperture size of 850 lm. The HGMS experiment consisting of one rougher and one scavenger was carried out at the magnetic intensities of 0.05 T and 0.3 T, respectively. Additionally, HGMS was used as the cleaner for the concentration of flotation rougher concentrate. Three times of cleaning by HGMS were applied at the intensities of 2.0, 0.5, and 0.05 T. 3. Results and discussion 3.1. Chemical analysis
2.3. Flotation and reagents Flotation experiments were carried out using an Outotec laboratory flotation machine. The samples of 1.5 kg with the size of 1.5 mm were ground at a laboratory rod mill as the head samples for flotation. The flotation cells with the volumes of 4 and 1.5 liters were used in the rougher and cleaner flotation, respectively. Tap water was used for wet grinding and flotation. The room temperature (20–23 °C) was applied for flotation conditioning. Hydroxamate type of reagent Aero6494, fatty acid type of reagent Aero704 and sulfosuccinamate type of reagent Aero 845 from Cytec and sodium oleate (Na-oleate) powder (P82 wt% fatty acids as oleic acid) manufactured by Sigma–Aldrich were tested as the collectors. Sodium silicate (Na2SiO3) and starch were used as the depressants of silicate minerals during flotation. Aero 6494, Aero 704 and Aero 845 were used without dilution and sodium oleate was diluted into 5 wt% solution for use. The value of pH was adjusted by using soda ash and sulfuric acid.
Chemical analyses of the sample by XRF, ICP-OES and ICP-MS were conducted at Labtium Oy Finland. The contents of rare earth elements (REEs) by different chemical analysis methods are shown in Table 1. The content of total REEs (TREE) by ICP-MS is 3000 ppm or 0.3% in which the content of the heavy REEs (Gd to Y) is 562 ppm or 0.056%. The ratio of the heavy REEs to the light REEs is 18.7– 81.3%. 3.2. Modal mineralogy The data of modal mineralogy of the sample are shown in Table 2. A total of 20 minerals were identified by MLA in which silicates take up 97.1%, carbonates 0.52%, and oxides and other minerals 2.41% in mass. The identified REE minerals and REE-bearing minerals are parisite, bastnaesite, synchysite-(Y), pyrochlore, gittinsite, britholite, elpidite and zircon. REEs are distributed not only in carbonates and oxides (parisite, bastnaesite, synchysite-(Y),
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 3 REE flotation performance comparison at the grinding size of P80 = 50 lm and 80 lm.
a
Collector
Grinding size, P80 lm
La %
Rec%
%
Rec%
%
Rec%
%
Rec%
Aero 704, 845a
50 80
0.37 0.41
55.76 57.63
0.61 0.66
56.65 57.98
0.09 0.10
30.56 30.13
1.07 1.16
52.42 53.76
Aero 6494a
50 80
0.12 0.11
52.49 54.60
0.19 0.17
53.50 55.14
0.03 0.03
36.66 38.20
0.34 0.32
50.81 52.53
Ce
Y
REE (La, Ce, Y)
Dosage 200 g/t.
pyrochlore) but also in silicate minerals (gittinsite, britholite, elpidite and zircon). Light REEs are mainly contained in carbonate minerals (parisite, bastnaesite) and silicate mineral (britholite) while Y and heavy REEs are mainly contained in Zr silicate minerals (zircon, elpidite, gittinsite), Nb mineral (pyrochlore) and carbonate mineral (synchysite-(Y)). The SEM images of REE minerals in Fig. 1 show the appearance of bastnaesite, synchysite, parisite, britholite, pyrochlore and zircon in fine grains. 3.3. Determination of flotation conditions 3.3.1. Grinding size The previous mineralogical analyses by AIST in Japan on the grain size of REE minerals in the ore have shown that the reasonable grinding size for the beneficiation would be around 80% passing 50–100 lm. To select a suitable grinding size the flotation tests were carried out at the grinding size of P80 = 50 lm and P80 = 80 lm by using Aero 704, 845 (in the ratio 1:1) and Aero 6494 as the collectors at the dosage of 200 g/t and Na2SiO3 as the depressant at the dosage of 300 g/t. The desliming of 20 lm by the hydrodynamic settling method was applied before flotation. The results showed in Table 3 indicated that the flotation performances of the REE minerals at the two grinding sizes were not obviously different. 3.3.2. Collector type To select a proper collector the flotation tests using Aero 704, 845 (in the ratio of 1:1), Na-oleate and Aero 6494 as the collectors were carried out in the presence of sodium silicate (Na2SiO3) as the depressant. The grinding size was P80 = 50 lm and the temperature during pulp conditioning was at 20–23 °C. When the dosages
of collectors varied in the range of 70–200 g/t the different flotation recoveries of the total REE (La, Ce, Y) were obtained. The curves of the grade versus recovery of REE (La, Ce, Y) in the concentrate for different collectors are shown in Fig. 2. It was demonstrated that higher recovery and grade of REE (La, Ce, Y) were obtained by using Na-oleate as the collector. That is, comparing with Aero 704, 845 and Aero 6494, Na-oleate was more selective for REE flotation for this ore.
3.3.3. Depressant dosage Because of the high content of silicates in the ore the depression of silicate minerals is important for REE flotation. Two levels of 300 g/t and 50 g/t of Na2SiO3 dosage were tested using Na-oleate as the collector. The flotation results were shown in Fig. 3. It is obvious that higher grade and recovery of REE (La, Ce, Y) in flotation concentrates were obtained at dosage 300 g/t than 50 g/t. By the flotation experiments discussed above on the grinding size, collector type and Na2SiO3 dosage the flotation conditions were determined: grinding size P80 = 80 lm, collector Na-oleate at dosage 200–300 g/t and depressant Na2SiO3 at dosage 300 g/t. Starch could be added to improve the depression of Fe-oxides.
3.4. Rougher flotation and results The rougher flotation experiment was carried out at the grinding size of P80 = 80 lm. The particle size distribution of the feed measured by a laser sizer is shown in Table 4. Na-oleate was used as the collector at the dosage of 270 g/t and Na2SiO3 and starch as the silicate depressants at the dosages of 300 g/t and 150 g/t. The flotation pH was natural in 8.8–9.3 and the pulp temperature for conditioning and flotation was at 20–23 °C.
Fig. 2. Effect of collector type on the flotation of REE (La, Ce, Y).
X. Yang et al. / Minerals Engineering 71 (2015) 55–64
59
Fig. 3. Effect of Na2SiO3 on the flotation of REE (La, Ce, Y).
Table 4 Particle size distribution of the feed. Particle size, lm Cumulative percentage, vol.%
3.19 10.00
12.40 20.00
38.88 50.00
72.79 75.00
81.18 80.00
103.90 90.00
The flow sheet of the experiment is shown in Fig. 4. The total flotation time was 28 min. Six froth concentrates (concentrate 1– 6) were respectively pulled in the flotation periods of 0–2 min, 2–4 min, 4–6 min, 6–9 min, 9–14 min, and 14–28 min and the tail was the material left in the cell. The flotation products of Conc 1–6 and Tail were filtered, tried, weighed and sent for assay by XRF. The recoveries of elements in the products were calculated based on the weights and assays. The element contents of the feed were recalculated based on the data of products. All the experimental data are shown in Table 5. For analyzing the experimental errors the element contents of the feed by XRF (Analysis feed) are listed in Table 5. The standard deviation (SD) between the calculated feed (X1) and the analysis feed (X2) was determined by
SD ¼
rffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi 1 X ðX i X m Þ2; 21
where Xm is the mean value of X1 and X2. The standard deviations between the calculated feed and the analysis feed on La, Ce and Y are shown in Fig. 5.
The weights of cumulative concentrates and the grades and recoveries of elements in the cumulative concentrates calculated from the experimental data in Table 5 are shown in Table 6. The mass percentage of the final cumulative concentrate Conc (1,2,3,4,5,6) was 11.5%. Nearly 90% of La and Ce were recovered but the recovery of Y was only 42%. The relationships of grade versus recovery of La, Ce and Y in the cumulative concentrates are shown in Fig. 6.
3.5. Mineralogical studies of flotation products In order to investigate the flotation recoveries and kinetics of different minerals, the mineralogical analyses by MLA were conducted on the concentrates 1, 3 and 5 and the tail. The modal mineralogy of these three concentrates and the tail were obtained. Based on the trend lines of mineral contents in flotation products versus flotation time the modal mineralogy of concentrates 2, 4 and 6 were achieved by calculation. The modal mineralogy of all the concentrates and the tail are shown in Table 7. Comparing the data in Table 7 with the modal mineralogy of ore in Table 2 all the major identified minerals and REE minerals are consistent. But some silicate minerals such as smectite, garnet, and allanite, and carbonate mineral siderite, were not identified in the analysis of ore. Among them allanite is a REE mineral. It is revealed in Table 7 that as the increase of flotation time the grades of carbonate and oxide minerals decrease and the grades of silicate minerals increase in the flotation concentrates from concentrate 1 to concentrate 6.
Fig. 4. Flow sheet of rougher flotation.
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 5 Assays of the flotation products and the calculated feed. Product
Weight g
Conc 1 21.10 Conc 2 20.30 Conc 3 16.30 Conc 4 19.10 Conc 5 31.10 Conc 6 64.50 Tail 1327.80 Calc’d feed (X1) 1500.20 Analysis feed (X2) Standard deviation (SD)
La
Ce
Y
Fe
SiO2
ZrO2
C
%
ppm
Rec%
ppm
Rec%
ppm
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
1.41 1.35 1.09 1.27 2.07 4.30 88.51 100.00
19,400 8600 5100 2810 1460 642 80 609 660 36.00
44.80 19.11 9.10 5.87 4.97 4.53 11.62 100.00
31,000 14,200 8500 4680 2480 1093 140 1002 1060 40.73
43.50 19.17 9.21 5.94 5.13 4.69 12.36 100.00
2190 1630 1350 1010 750 482 180 276 280 2.85
11.16 7.99 5.32 4.66 5.63 7.51 57.73 100.00
3.00 4.52 5.57 6.08 6.67 6.39 3.25 3.53 3.16 0.26
1.20 1.73 1.71 2.19 3.92 7.78 81.47 100.00
22.30 37.00 46.00 56.60 64.20 69.24 75.30 72.99 73.90 0.64
0.43 0.69 0.68 0.99 1.82 4.08 91.31 100.00
0.29 0.45 0.57 0.60 0.63 0.55 0.33 0.35 0.35 0.00
1.17 1.73 1.77 2.18 3.73 6.74 82.67 100.00
2.97 2.04 1.79 1.21 0.82 0.50 0.09 0.22 0.21 0.01
19.10 12.62 8.89 7.05 7.77 9.81 34.76 100.00
had not only high recoveries of over 96% in the final cumulative concentrate but also high flotation kinetics. At the time of 6 min their recoveries reached 81–90%. After the flotation of carbonate REE minerals two silicate REE minerals of cerite(Ce) and allanite were floated at slower kinetics. The recoveries of cerite(Ce) and allanite reached 79% and 48% at the flotation time of 15 min. However, the recoveries of one Nb mineral pyrochlore and two Zr minerals zircon and elpidite were very low and their flotation kinetics were the slowest. Because these minerals are the main bearing minerals of Y and other heavy REEs the low recoveries of them caused the low recoveries of Y and other heavy REEs in concentrates. 3.6. High gradient magnetic separation (HGMS) Fig. 5. Standard deviations between the calculated feed and the analysis feed on La, Ce and Y.
As shown in Table 8 the recoveries of minerals in the cumulative concentrates at different flotation times were calculated based on the modal mineralogy in Table 7 and the mass percentages of flotation products. Over 95% of carbonates were recovered in final cumulative concentrate in which the recoveries of three major REE minerals bastnaesite, parisite, synchysite and synchysite-(Y) reached over 96%. Meanwhile, very high recoveries of calcite and siderite (94% and 92% respectively) were obtained in the final cumulative concentrate. The recovery of fluorite reached as high as 98% but the recoveries of other oxide minerals such as pyrochlore and Fe-oxide were below 30%. Although the recoveries of major silicates minerals quartz, plagioclase, K-feldspar and arfvedsonite were quite low in the range of 7–17% those of silicate REE minerals such as cerite(Ce) and allanite were quite high and reached 86% and 65%. This means the collector Na-oleate was quite selective for the recovery of silicate REE minerals from other major silicates. The recoveries of different REE minerals in cumulative concentrates at different times are shown in Fig. 7. The carbonate group of REE minerals (bastnaesite, parisite, synchysite and synchysite-(Y))
As shown in Fig. 6 the recovery of Y in the rougher flotation concentrate was only 45%. The poor recovery of Y can be explained by the poor recoveries of Y bearing minerals pyrochlore, zircon and elpidite as discussed above. A high gradient magnetic separation (HGMS) experiment was carried out. The matrix type of 3.5 XRO with the aperture size of 850 lm was used. Two stages of rougher and scavenger separations were applied at the intensities of 0.05 T and 0.3 T, respectively. The rougher and scavenger magnetic concentrates were combined as the final concentrate. The recoveries of La, Ce and Y versus concentrate mass for flotation and HGMS are shown in Fig. 8. Comparing to flotation HGMS was significantly less selective for recoveries of La and Ce. Only 74.7% recovery of La and Ce was obtained at a high mass percentage (35.8%). But the recovery of Y reached 67.1%. That is, HGMS could be more selective for the recoveries of pyrochlore, zircon and elpidite than flotation using Na-oleate as the collector. 3.7. Cleaning process and comparison of flow sheets Although the results of HGMS could be improved by optimizing the conditions it is not capable to replace flotation at rougher stage due to low La and Ce recoveries and grades. After rougher flotation
Table 6 Weights of cumulative concentrates, and grades and recoveries of elements in the cumulative concentrates. Cumulative conc
Conc Conc Conc Conc Conc Conc
1 (1,2) (1,2,3) (1,2,3,4) (1,2,3,4,5) (1,2,3,4,5,6)
Weight
La
Ce
Y
Fe
SiO2
ZrO2
C
g
%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
21.10 41.40 57.70 76.80 107.90 172.40
1.41 2.76 3.85 5.12 7.19 11.49
1.94 1.41 1.16 0.94 0.71 0.47
44.80 63.90 73.00 78.87 83.84 88.38
3.10 2.28 1.87 1.52 1.16 0.76
43.50 62.67 71.88 77.82 82.95 87.64
0.22 0.19 0.18 0.16 0.13 0.10
11.16 19.15 24.47 29.13 34.76 42.27
3.00 3.75 4.26 4.71 5.28 5.69
1.20 2.93 4.64 6.83 10.75 18.53
22.30 29.51 34.17 39.75 46.79 55.19
0.43 1.12 1.80 2.79 4.61 8.69
0.29 0.37 0.43 0.47 0.52 0.53
1.17 2.90 4.67 6.86 10.59 17.33
2.97 2.52 2.31 2.04 1.69 1.24
19.10 31.72 40.61 47.66 55.43 65.24
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Fig. 6. Relationships of grade versus recovery of La, Ce and Y in cumulative concentrates.
Table 7 Modal mineralogy of flotation products by MLAa. Mineral Silicate
Carbonate
Oxide and other
Product Flotation time, min
Conc 1 0–2
Conc 2 2–4
Conc 3 4–6
Conc 4 6–9
Conc 5 9–14
Conc 6 14–28
Tail >28
Feed 33.34 27.37 26.61 7.57 0.07
Quartz Plagioclase (Albite) K-felspar Arfvedsonite Gittinsite Smectite Garnet Cerite(Ce) (Britholite) Allanite Zircon Elpidite Other Total
7.48 6.01 4.15 4.07
15.81 12.21 7.26 7.81
18.80 15.00 10.75 11.07
23.45 17.06 12.41 14.16
26.33 20.59 15.80 17.46
32.03 22.51 23.56 16.88
35.54 23.06 26.91 8.68
1.92 0.68 1.68 0.21 0.58 0.15 0.02 26.95
4.22 0.91 1.34 0.19 0.89 0.24
5.68 1.00 1.02 0.17 1.10 0.33 0.03 64.95
6.54 0.93 0.70 0.19 1.20 0.35
6.18 0.61 0.29 0.18 1.17 0.42 0.04 89.07
4.72 0.27 0.13 0.11 0.75 0.51
3.64 0.02 0.01 0.01 0.45 0.52 0.01 98.84
Calcite Siderite Bastnaesite(Ce) Parisite Synchysite(Y) Synchysite(Ce) Other Total
16.09 1.43 9.52 4.93 0.39 3.24 0.04 35.63
Fluorite Pyrochlores Fe-oxide Other Total
34.52 0.26 0.94 1.70 37.42 1.94 3.10
50.88 11.22 2.16 4.81 3.39 0.26 1.52
76.98
7.56 2.43 1.65 1.67 0.12 0.45 0.13 14.01
4.47 2.25 0.40 1.25 0.09 0.34
2.12 1.14 0.23 0.28 0.02 0.15 0.10 4.04
9.83 0.30 1.52
26.16
16.86 0.36 1.65 2.18 21.05
3.48 0.26 1.47 1.68 6.89
0.86 1.42
0.51 0.85
0.28 0.47
0.15 0.25
23.37 24.61 0.30 1.25
101.74 0.93 0.49 0.10 0.12 0.01 0.06
0.12 0.34 0.91 0.74 97.07
0.04 0.02 0.01 0.00 0.00 0.00 0.01 0.08
0.52
2.88
0.03 0.08 0.43 0.54 1.08
0.51 0.17 0.72 1.01 2.41
0.06 0.11
0.01 0.01
0.06 0.10
1.71 1.49 0.19 1.20
0.19 0.12 0.19 0.02
REE La Ce a
Flotation conditions: Na-oleate 270 g/t, Na2SiO3 300 g/t, starch 150 g/t, pH 8.8–9.3. Unit of mineralogy: wt%.
the combined concentrate of Conc 1,2,3,4,5,6 was respectively cleaned by flotation and HGMS. The flow sheet is shown in Fig. 9. In the cleaner flotation Na-oleate was used as the collector and three time cleaners were applied. In the cleaner HGMS stage three time cleaners were applied at the intensities of 2.0, 0.5 and 0.05 T and the matrix type was 3.5 XRO with the aperture size of 850 lm. The testwork results for the two flow sheets are shown in Fig. 10. The recoveries of La and Ce were 68–73% and not obviously different for the two flow sheets. The grade of La plus Ce reached
4.2%. But the flow sheet of flotation–HGMS was showed to be more effective for recovery of Y. That is, comparing to the flow sheet of flotation–flotation a higher recovery of Y was obtained by the flow sheet of flotation–HGMS at the same recovery of La and Ce. The assays of the two cleaned concentrates from the flow sheets of flotation–flotation and flotation–HGMS are shown in Table 9. The grades of REE (La, Ce, Y) for the two flow sheets were 4.9% and 4.5%, respectively, at the recoveries of 53.9% and 67.4%. The
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Table 8 Recoveries of minerals in cumulative concentrates. Conc 1
Quartz Plagioclase K-felspar Arfvedsonite Smectite Garnet Cerite(Ce) Allanite Zircon Elpidite Silicate Calcite Siderite bastnaesite(Ce) Parisite Synchysite(Y) Synchysite(Ce) Carbonate Fluorite Pyrochlores Fe-oxide Oxide and other
Conc 1,2
Conc 1,2,3
Conc 1,2,3,4
Conc 1,2,3,4,5
Conc 1,2,3,4,5,6
Calc’d feed
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
wt%
Rec%
7.48 6.01 4.15 4.07 1.92 0.68 1.68 0.21 0.58 0.15 26.93 16.09 1.43 9.52 4.93 0.39 3.24 35.59 34.52 0.26 0.94 35.72
0.31 0.38 0.23 0.62 0.71 11.14 27.69 10.29 1.62 0.42 0.39 35.37 12.39 55.87 42.32 44.09 56.13 38.53 37.65 3.68 2.51 26.22
11.56 9.05 5.67 5.91 3.05 0.79 1.51 0.20 0.73 0.19 38.67 13.70 1.79 7.21 4.18 0.33 2.40 29.60 29.66 0.28 1.09 31.04
0.93 1.11 0.61 1.77 2.23 25.59 48.83 19.59 4.03 1.06 1.11 59.09 30.42 83.04 70.35 72.85 81.54 62.87 63.48 7.70 5.73 44.70
13.61 10.73 7.11 7.36 3.79 0.85 1.37 0.19 0.83 0.23 46.09 11.97 1.97 5.64 3.47 0.27 1.85 25.16 26.05 0.30 1.25 27.60
1.53 1.84 1.07 3.07 3.86 38.37 61.72 26.10 6.41 1.78 1.84 71.93 46.68 90.52 81.40 83.33 87.63 74.48 77.69 11.55 9.14 55.40
16.05 12.31 8.42 9.05 4.48 0.87 1.21 0.19 0.92 0.26 53.77 10.10 2.04 4.34 2.91 0.23 1.47 21.09 22.01 0.30 1.32 23.63
2.40 2.81 1.69 5.02 6.06 52.31 72.11 34.48 9.47 2.67 2.85 80.83 64.30 92.64 91.10 92.93 92.91 83.11 87.40 15.38 12.82 63.14
19.02 14.69 10.55 11.48 4.97 0.80 0.94 0.19 0.99 0.31 63.93 7.80 1.78 3.15 2.15 0.17 1.09 16.15 16.67 0.29 1.36 18.32
4.00 4.71 2.97 8.94 9.45 67.16 79.10 47.55 14.32 4.38 4.77 87.71 78.80 94.64 94.61 96.69 96.68 89.39 92.99 20.79 18.61 68.78
23.89 17.62 15.42 13.50 4.87 0.60 0.64 0.16 0.90 0.38 77.98 5.23 1.30 2.01 1.39 0.11 0.71 10.75 10.99 0.25 1.30 12.55
8.03 9.03 6.92 16.81 14.81 80.76 85.65 64.75 20.80 8.76 9.29 93.93 91.78 96.51 97.76 100.00 100.00 95.06 97.98 28.86 28.41 75.25
34.20 22.43 25.59 9.23 3.78 0.09 0.09 0.03 0.50 0.50 96.43 0.64 0.16 0.24 0.16 0.01 0.08 1.30 1.29 0.10 0.53 1.92
100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00 100.00
The bold values mean the sum of percentages of silicate group minerals, the sum of percentages of carbonate group minerals and the sum of percentages of oxide and other group minerals.
Fig. 7. Recoveries of REE minerals in cumulative concentrates versus flotation time.
Fig. 8. Recoveries of REE (La, Ce) and Y versus concentrate mass for flotation and HGMS.
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X. Yang et al. / Minerals Engineering 71 (2015) 55–64
Fig. 9. Flow sheet of cleaning.
Fig. 10. La plus Ce and Y grade versus recovery for rougher–cleaner flotation flow sheet and rougher flotation–cleaner HGMS flow sheet.
Table 9 Assays of the concentrates from flow sheets of flotation–flotation and flotation–HGMS. Element
Flotation–Flotation conc Flotation–HGMS conc Feed
La
REE (La, Ce, Y)
Fe
%
Rec%
Ce %
Rec%
Y %
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
%
Rec%
1.76 1.59 0.06
59.30 72.84 100.00
2.85 2.60 0.10
60.50 75.07 100.00
0.24 0.27 0.03
17.98 27.92 100.00
4.85 4.45 0.19
53.85 67.40 100.00
2.72 12.48 3.51
1.62 10.10 100.00
11.50 37.06 73.09
0.33 1.44 100.00
1.11 4.76 9.94
0.23 1.36 100.00
0.38 0.86 0.35
2.27 7.02 100.00
3.47 3.19 0.21
35.19 44.08 100.00
grade of REE (La, Ce, Y) in the ore from Table 1 was 0.19%. So that, the enrichment ratios for the two processes were 25.7 and 23.7, respectively. These are quite high values of beneficiation enrichment for this type of complex ore. Table 9 also indicates that the both concentrates had high contents of C and the flotation–HGMS concentrate had a high content of Fe. That is, calcite and Fe-oxides were also concentrated in the cleaned concentrates. To further increase the grade of REE in the concentrates calcite and Fe-oxides should be eliminated. Further flotation by using more selective collectors or properly controlling pH value could be effective to remove calcite and further magnetic separation by adjusting the magnetic intensity or gradient could work to decrease the content of Fe-oxides. Acid leaching (sulfuric or hydrochloric acids) at high concentration after roasting (200–400 °C) could be used for the downstream metallurgical processing of the concentrates. REE carbonate minerals (bastneasite, parisite etc.) should be dissolved in the acids at high recoveries. Due to high content of calcite a pre-leaching with diluted HCl at room temperature may be needed to dissolve calcite and purify the material prior to REE leaching. The leachability of REEs in silicate minerals may be low and the caustic pre-treatments with calcium hydroxide or sodium hydroxide should be needed to transform them into soluble salts prior to leaching.
SiO2
Al2O3
ZrO2
C
4. Conclusions Using Na-oleate as the collector at the dosage of 270 g/t and Na2SiO3 and starch as the silicate depressants at the dosages of 300 g/t and 150 g/t, nearly 90% of La and Ce were recovered but the recovery of Y was 45% in the final cumulative concentrate with the mass percentage of 11.5%. The mineralogical studies indicated that over 95% of carbonates were recovered in the final cumulative concentrate in which the recoveries of three major REE minerals bastnaesite, parisite and synchysite reached over 96%. They showed also high flotation kinetics. At the flotation time of 6 min their recoveries reached 81–90%. Meanwhile, the recoveries of silicate REE minerals such as cerite(Ce) and allanite were quite high and reached 85.6% and 64.7%, respectively. Their flotation kinetics was slower than that of carbonate REE minerals and the recoveries of cerite(Ce) and allanite reached 79.1% and 47.6% at the flotation time of 15 min. However, the recoveries of Nb mineral pyrochlore and Zr minerals zircon and elpidite were very low. The cumulative concentrate of flotation was cleaned by flotation and high gradient magnetic separation HGMS. The grades of REE (La, Ce, Y) for the two flow sheets were 4.9% and 4.5%, respectively, at the recoveries of 53.9% and 67.4%. The enrichment ratios for the two processes were 25.7 and 23.7 for the beneficiation of
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REE (La, Ce, Y). Comparing the results by the two flow sheets a higher recovery of Y was obtained by the flow sheet of flotation– HGMS at the same recovery of La and Ce. References Assis, S.M. et al., 1996. Utilisation of hydroxamates in minerals froth flotation. Miner. Eng. 9 (1), 103–114. Cheng, Ta-Wui, Holtham, P.N., Tran, Tam, 1993. Froth flotation of monazite and xenotime. Miner. Eng. 6 (4), 341–351. Cheng, Ta-Wui et al., 1994. The surface properties and flotation behavior of xenotime. Miner. Eng. 7 (9), 1085–1098. Chi, Ruan, Tian, Jun, 2007. Introduction of weathered crust elutiondeposited rare earth ore. J. Chinese Rare Earth Soc. 25 (6), 641–649 (Chinese). Gupta, C.K., Krishnamurth, N., 2005. Extractive Metallurgy of Rare Earths. CRC Press. Jordens, A., Cheng, Y.P., Waters, K.E., 2013. A review of the beneficiation of rare earth element bearing minerals. Miner. Eng. 41, 97–114.
Miyawaki, R., Nakai, I., 1996. Crystal chemical aspects of rare earth minerals. In: Jones, Adrian P., Wall, Frances, Terry Williams, C. (Eds.), Rare Earth Mineral: Chemistry, origin and ore deposits. Chapman & Hall, pp. 21–37. Pavez, O., Peres, A.E.C., 1993. Effect of sodium metasilicate and sodium sulphide on the flotability of monazite-zircon-rutile with oleate and hydroxamates. Miner. Eng. 6 (1), 69–78. Pavez, O. et al., 1996. Technical note adsorption of oleate and octyl-hydroxamate onto rare-earths minerals. Miner. Eng. 9 (3), 357–366. Pradip, Fuerstenau, D.W., 1991. The role of inorganic and organic reagents in the flotation separation of rare-earth ores. Int. J. Miner. Eng. 32, 1–22. U.S. Geological Survey, 2002. Rare earth elements-critical resources for high technology, USGS Fact Sheet 087-02. Yongfu Yu, 2000. Mineral processing techniques of rare earths and development in China, West China Exploration Engineering, series No.63, pp. 1–4. (Chinese). Zhang, J., Edwards, C., 2012. A review of rare earth mineral processing technology. In: 44th Annual Meeting of the Canadian Mineral Processors. CIM, Ottawa, pp. 79–102.