- Email: [email protected]
Recovery of rare earths from wet-process phosphoric acid
Hydrometallurgy 101 (2010) 41–47
Contents lists available at ScienceDirect
Hydrometallurgy j o u r n a l h o m e p a g e : w w w. e l s e v i e r. c o m / l o c a t e / h y d r o m e t
Recovery of rare earths from wet-process phosphoric acid Liangshi Wang, Zhiqi Long ⁎, Xiaowei Huang, Ying Yu, Dali Cui, Guocheng Zhang National Engineering Research Center for Rare Earth Materials, General Research Institute for Nonferrous Metals, and Grirem Advanced Materials Co. Ltd., Beijing, 100088, China
a r t i c l e
i n f o
Article history: Received 30 July 2009 Received in revised form 23 November 2009 Accepted 23 November 2009 Available online 1 December 2009 Keywords: Rare earths extraction Phosphorite Wet-process phosphoric acid D2EHPA
a b s t r a c t Phosphorite ores are a potential resource of rare earths (RE) as well as phosphate so the recovery of rare earths during the wet processing of phosphoric acid is important. This study investigates the influence of operating conditions and crystal modifiers on the leaching of RE and the solvent extraction of RE with organo-phosphorus reagents. The results indicate that lower temperature, higher concentration of phosphoric acid and larger liquid/solid ratio are beneficial to RE enrichment in the phosphoric acid. Surfactant additives which enhance the crystal growth of gypsum also enhanced RE leach recovery about 75% under optimized conditions. Studies on the solvent extraction of RE found that D2EHPA mixtures with neutral organo-phosphorus reagents were antagonistic and that Fe3+ competed strongly over RE. Higher D2EHPA concentration, larger phase ratio, lower temperature and lower phosphoric acid concentration increased the RE extraction efficiency. A negative enthalpy change was found indicating an exothermic extraction reaction. © 2009 Elsevier B.V. All rights reserved.
1. Introduction Phosphorite is a rare earths bearing mineral that may occur as the main component in individual deposits as well as associated with other minerals, such as apatite (Jorjani et al., 2008). The total reserves of phosphorite ores in the world is about 100 billion tons (Ilyin, 1998; Becker, 1983) which on average contain 0.5‰ RE. Table 1 showed the composition of rare earths in various phosphorite reserves. In particular, Russia has abundant reserves of high grade phosphorite ore at Kola which contain up to 1% rare earths and is considered as the primary rare earths resource in the country. Phosphorite is also relatively abundant in China where the Zhijin phosphorite deposit in Guizhou province has reserves of more than 1.6 billion tons containing more than 1.4 million tons. Another deposit is in Fanshan. Therefore it is important to study the recovery of the rare earths from phosphorite which is a phosphate mineral suitable for the manufacture of phosphoric acid. Wet processing, using the digestion of phosphate rocks by sulphuric acid, is the most common process used to produce phosphoric acid. The overall and simplified reaction between calcium phosphate minerals and sulphuric acid (Becker, 1983) can be written as: Ca5 FðPO4 Þ3 þ 5H2 SO4 þ xH2 O ¼ 3H3 PO4 þ 5CaSO4 ·x=5H2 O↓ þ HF↑ where x is 0 for calcium sulphate anhydrate, 0.5 for hemihydrate and 2 for dihydrate, i.e. gypsum. Studies show that the dominant species of
⁎ Corresponding author. Tel.: + 86 10 82241188; fax: + 86 10 62355405. E-mail address: [email protected] (Z. Long). 0304-386X/$ – see front matter © 2009 Elsevier B.V. All rights reserved. doi:10.1016/j.hydromet.2009.11.017
rare earths during the wet-process phosphoric acid (WPA) is phosphate, which has low solubility (Xu, 1995). Therefore, rare earths will be enriched in the phosphor-gypsum with only 20%–30% RE in the phosphoric acid leaching solution. In South Africa (Preston et al., 1996), it was reported that the phosphor-gypsum adsorbed 85% RE during the acidulation of Phalaborwa phosphorite with sulphuric acid. Similarly, in Poland (Jarosinski et al., 1993), it was found that 70% RE entered the calcium sulphate dihydrate (CaSO4·2H2O, DH or gypsum) during dihydrate process for the Kola phosphorite, whilst almost all the rare earth was adsorbed by the calcium sulphate hemihydrate (CaSO4·1/2H2O, HH) during the hemihydrate process. Hence, various recovery techniques of rare earths from the phosphogypsum have been developed. For instance, in Russia the rare earths were leached out from phosphor-gypsum with sulphuric acid and then the leaching solution was evaporated to crystallize rare earths or was neutralized to precipitate the rare earths (Lokshin et al., 2004; Lededev and Lokshin, 2000; Lededev et al., 2000). In Poland, it was reported that rare earths was precipitated with hydrofluoric acid or extracted with nonyl–phenyl phosphoric acid (NPPA) to obtain a 40% concentrate from the phosphor-gypsum leaching solution (Jarosinski et al., 1993). However, the process of the rare earths recovery from phosphor-gypsum is too complicated and uneconomical to be industrialized. On the other hand, rare earths can be recovered directly from phosphoric acid leaching solution and the key problem is to enhance their enrichment. Pessac et al. (1987) mentioned that during the acidulation of the phosphate rock, the leaching efficiency of RE was increased to 56% by introducing aluminum and/or iron and/or silica minerals into the slurry. However, this may lead to difficulties in subsequent stages of phosphoric acid production.
42
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
Table 1 Content of rare earth elements in various phosphate rocks (Krea and Khalaf, 2000; Koopman and Witkamp, 2000). Deposit origin
Florida Pebble
Morocco Khoribga
Algeria Djebel Onk
Russia Kola
Palfos
Kovdor
Jordan
Israel
Lanthanides%
0.29
0.14–0.16
0.13–0.18
0.84–0.88
0.48
0.14
0.01
0.01
Table 2 Content of rare earth elements in various phosphorite phosphate rocks (ppm). Deposit origin
Y2O3
La2O3
CeO2
Pr6O11
Zhijin Fanshan
434.2 248.9
273.2 1034
162.7 2170
45.8 285.0
198.2 1201
35.2 229.6
8.5 56.9
Deposit origin
Ho2O3
Er2O3
Yb2O3
Lu2O3
Sc2O3
∑ REO
LREO
HREO
6.8 –
1295.0 5513.2
723.7 4976.1
571.3 537.1
Zhijin Fanshan
7.4 9.1
20.4 19.8
Tm2O3 2.3 1.9
Nd2O3
12.0 9.7
Since 1950s, solvent extraction methods have been studied to recover trace elements from wet-process phosphoric acid leach solutions, focusing on the recovery of uranium. In 1967, Oak Ridge National Laboratory (ORNL) recovered uranium using a synergistic extractant of DEHPA-TOPO (tri-n-octyl-phosphine oxide), and carried out large scale production (Hurst et al., 1969, 1972; Hurst and Crouse, 1973). Afterwards, liquid Cyanex 923 was used to replace solid TOPO as this offers a materials handling advantage (Rickelton, 1988). Recently, Institute of Atomic Physics in Romania (Bunu and Dumitrescu, 1992; Bunu et al., 1994), University of Blida in Algeria (Krea and Khalaf, 2000) and Rare Earth Development Section of BARC in India (Harvinderpal et al., 2004) studied the liquid–liquid extraction of uranium and lanthanides from wet-process phosphoric acid using a synergistic acidic and neutral organo-phosphorus extractants. As indicated, all the research focused on the uranium recovery. Therefore, this study focuses on both the leaching and solvent extraction method to recover rare earths from phosphorite ores during wet-process phosphoric acid operation. 2. Experimental 2.1. Reagents and solutions The phosphorite ore (Fanshan Phosphorite) used in this study consisted of 33.8% P2O5, 51.0% CaO, and 0.55% REO, 0.8% MgO, 1.5% FeO, 0.5% Al2O3. The rare earth element content is shown in Table 2. The di-(2-ethylhexyl) phosphoric acid (D2EHPA) was purified by nantokite before use (purity N99%) (Zhang et al., 2008). Cyanex 923 was supplied by CYTEC Canada INC. The extractants were diluted with kerosene to the required concentration. Kerosene was sulfonated before use. Stock solutions of each rare earth were prepared by dissolving their oxide with a purity N99.9% in concentrated sulphuric acid and diluting with deionized water. Unless indicated otherwise, all reagents were reagent-grade.
1.4 1.22
Sm2O3
Eu2O3
Gd2O3
Tb4O7
42.9 166.6
5.9 16.9
Dy2O3 35.5 62.8 LREO/HREO 1.27 9.26
by mass balance. RE extraction efficiency was obtained from RE concentrations extracted. 3. Results and discussion 3.1. Leaching of phosphorite concentrate 3.1.1. Influence of WPA process conditions on the RE leaching efficiency Generally, the dominant species of rare earths in phosphoric acid is RE phosphate which has a low solubility, e.g. the solubility product (Ksp) of the LaPO4 is 4 × 10− 23 and that of CePO4 is 1.6 × 10− 23 at 298 K. Nevertheless, the solubility of rare earths in phosphoric acid can be increased by increasing the acid concentration which is a benefit for rare earths leaching (Cetiner et al., 2005 and references therein). A series of experiments performed with increasing concentrations of phosphoric acid ranging from 5% to 25% (by the weight of P2O5) found a significant increase in RE leaching efficiency. The leaching efficiency for the elements is in the order Lab Ceb Pr b Ndb Sm N Gd b Dyb Y which rises with increasing atomic number (except for Y) and is in reasonable agreement with the solubility products of the RE phosphates (Firsching and Brune, 1991; Byrne and Kim, 1993; Liu and Byrne, 1997). As shown in Fig. 1, the RE leaching efficiency increased sharply from 32% when the phosphoric acid contained 5% P2O5, to 71% with phosphoric acid containing 25% P2O5. In addition, as the concentration of phosphoric acid increased, the energy consumption reduced. However, substitution of SO2− ions by HPO2− ions in the gypsum crystal lattice also increases, 4 4
2.2. Leaching studies The wet-process phosphoric acid production was simulated in a batch reactor and the rare earth elements in the phosphoric acid and gypsum were determined by ICP (Inductively Coupled Plasma). The morphology of the gypsum was recorded by scanning electron microscopy, after washing by ethanol and dried at 338 K for 3 h. 2.3. Solvent extraction studies The organic and the aqueous phases were shaken in a separation funnel at constant temperature to reach equilibrium. After phase separation, take aliquot of aqueous solution to analyze rare earths with ICP. The RE concentrations in the organic phase were calculated
Fig. 1. Effect of concentration of phosphoric acid on RE leaching efficiency 343 K, 0.2% NS-2 with 4000 molecule weight, 4:1 ratio of liquid to solid, 1.05 stoichiometric ratio of H2SO4 to CaO.
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
43
Fig. 2. Effect of stoichiometry of H2SO4 to CaO on RE leaching efficiency 343 K, 0.2% NS-2 with 4000 molecule weight, 4:1 ratio of liquid to solid, 21% of phosphoric acid.
Fig. 4. Effect of ratio of liquid to solid on RE leaching efficiency 343 K, 21% P2O5 of phosphoric acid, 0.2% NS-2 with 4000 molecule weight, 1.05 stoichiometric ratio of H2SO4 to CaO.
which results in the loss of P2O5. Thus, about 20%–25% phosphoric acid is optimum for the wet-process phosphoric acid production. According to Becker (1983), sulphuric acid plays an essential role in the crystallization process of dehydrate. With better crystallization, there is less eutectic crystallization and adsorption between lanthanide phosphates and calcium sulphate. Therefore special attention was paid to the effect of the stoichiometric ratio of H2SO4 to CaO that is present in the phosphorite. Fig. 2 shows the effect of varying the mole ratio of H2SO4:CaO from 1.0:1.0 to 1.2:1.0 which is comparable to the industrial mole ratio (0.9:1.0–1.2:1.0), as dictated by the composition and nature of the phosphate rock used (Becker, 1983). The results reveal that RE leaching passes through a maximum when the stoichiometric ratio is about 1.10:1.0 and that excess sulphuric acid affects the solubility and supersaturation of gypsum (El Moussaouiti et al., 1997; Samir et al., 2001). Clearly, the RE leaching efficiency can be enhanced by controlling the crystallization rate and the crystal shape of the calcium sulphate. When the stoichiometric ratio is 1.1, the calcium sulphate is formed with large clusters as shown in Fig. 5(c), which is easily filtered and scrubbed. This observation is consistent with the results reported by El Moussaouiti et al. (1996) who found that calcium sulphate
dihydrate crystals agglomerate more or less depending on the excess of sulphuric acid in sulpho-phosphoric acid solutions. Generally, during the wet-process phosphoric acid production, high sulphuric acid concentration leads to encapsulation and passivation of phosphorite (Becker, 1983). Therefore, the appropriate concentration of sulphuric acid increases the crystallinity of the calcium sulphate to enhance the RE leaching efficiency and at the same time, reduces the substitution of SO2− by HPO2− in the calcium sulphate crystal to 4 4 enhance the productivity of phosphoric acid. A series of experiments was carried out in the temperature range 343 K–358 K to study the effect of temperature on RE leaching efficiency. Lower temperatures lead to stabilized foams, resulting in substantially increased consumption of defoamer. While higher temperatures lead to the corrosion of equipment and to higher solubility of impurities which determine the quality of the phosphoric acid. The results obtained reveal that the RE leaching efficiency rapidly decreases with a small increase of temperature (Fig. 3). It reached 61.4% at 343 K and decreased to 25.1% at 348 K and reduced further with increasing temperature. This is attributed to the sharp decrease in solubility of rare earths phosphate in the phosphoric acid with the increasing temperature (Cetiner et al., 2005). However, in industry, the crystallization conditions would deteriorate and the yield of the phosphoric acid would be lower as the temperature drops below 343 K. Thus, 343 K is considered to be the optimum temperature. A series of experiments was then performed at liquid/solid ratios ranging from 2:1 to 5:1 L/kg at 343 K, with 21% of phosphoric acid and 1.05 stoichiometric ratio of sulphuric acid. The results in Fig. 4 indicate that an increase in the liquid/solid ratio leads to a significant increase in RE leaching efficiency. However, a higher liquid/solid ratio leads to a Table 3 Effect of additives on RE leaching efficiency.
Fig. 3. Effect of temperature on RE leaching efficiency 0.2% NS-2 with 4000 molecule weight, 4:1 ratio of liquid to solid, 21% P2O5 of phosphoric acid, 1.05 stoichiometric ratio of H2SO4 to CaO.
No.
Active additives
RE leaching efficiency (%)
0419-1 0423-1 0423-2 0424-1 0802-2 0427-1 0427-2 0621-1 0716-2
– 0.5%IS-1 0.2%IS-2 0.2%IS-3 0.2%NS-1 0.1%CS-1 0.2%NS-2 0.2%NS-3 1%IA-1
19.8 17.4 23.5 21.4 53.0 29.2 61.4 40.7 60.5
343 K, 4:1 liquid to solid, 21% P2O5 of phosphoric acid and 1.05 stoichiometric ratio H2SO4 to CaO, the content of additives is measured with phosphorite weight percent. NS-2 with 4000 molecule weight.
44
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
Table 4 Effect of NS-2 with 4000 molecule weight on RE leaching efficiency. No.
Amount of NS-2/%
RE leaching efficiency (%)
0626-1 0427-2 0802-1 0626-2
0.1 0.2 0.3 0.5
46.6 61.4 75.2 62.4
343 K, 4:1 liquid to solid, 21% P2O5 of phosphoric acid and 1.05 stoichiometric ratio H2SO4 to CaO.
Table 5 Effect of molecule weight of NS-2 on RE leaching efficiency. No.
Molecule weight of NS-2
RE leaching efficiency (%)
0705-1 0802-1 0705-2
400 4000 20,000
61.2 75.2 73.6
343 K, 4:1 liquid to solid, 21% P2O5 of phosphoric acid, 1.05 stoichiometric ratio H2SO4 to CaO and 0.3% NS-2.
lower RE and P205 content in the pregnant liquor, which subsequently influences further processing. Therefore the optimal liquid/solid ratio is 4:1 L/kg, which gives a 61.4% RE leaching efficiency. 3.1.2. Effect of additives on RE leaching efficiency Most of the fundamental information in the literature regarding the use of additives is directed towards the inhibition of gypsum crystal growth (Smith and Alexander, 1970; Tadros and Mayes, 1979; Amjad, 1987). The additives affected the crystal growth mainly by entering into the crystal lattice, being selectively adsorbed on the interface or changing the crystal surface energy (Sangwal, 1993, 1996, 2002). In this study, the rare earths in the leach solution were enriched by adding non-ionic surfactant NS-1, NS-2 and inorganic additive IA-1, as shown in Table 3. This was mainly because the surfactant increases the aspect ratio of the formed crystals, decreases the induction time, decreases the nucleation rate, increases surface
energy and decreases the percentage of fine crystals (El-Shall et al., 2005). Based on the significant impact of NS-2 on the RE enrichment in phosphoric acid during the production process, the effect of concentration and molecular weight of NS-2 on RE leaching efficiency was studied. The results in Tables 4 and 5 show that the RE leached increases with increasing concentration until a maximum around 75% is reached with 0.3% NS-2; and was optimum with molecular weight of 4000. It appears that NS-2 delayed the onset of nucleation, as evidenced by turbidity, and inhibits nucleation, leading to fewer nuclei. Reduced numbers of nuclei lead to fewer, but larger crystals, as shown in Fig. 5(b), and higher RE leaching efficiency. This observation is consistent with the results in the gypsum growth inhibition literature which shows that the ability of polymers to inhibit crystal growth decreases with increasing molecular weight (Amjad, 1987; Free et al., 1999). 3.2. Solvent extraction for RE recovery 3.2.1. Selection of extractant Recovery of rare earths from wet-process phosphoric acid by solvent extraction is carried out in highly acidic systems. Generally, acid and neutral organo-phosphorus extractants were used with formula of (RO)3PO or R3PO with the extraction ability order of R3PON R2(RO)PON R (RO)2PON (RO)3PO. In this work, Cyanex 923 (a trialkyl-phosphine oxide) was first chosen, but for different phosphoric acid concentrations less than 10% extraction of rare earths was achieved by Cyanex 923. The extraction efficiency of acid organo-phosphorus extractants decreases with the increase of C–P bond energy and the existence of the RO branched chain, while the conjugative effect of a benzene ring is beneficial for extraction. Furthermore, acid organo-phosphorus extractants which have a low pKa value can be used in highly acidic systems, such as D2EHPA, P215, P406, DOPPA and DNPPA. However, uranium and lanthanides are difficult to strip from DOPPA, DNPPA, and need high concentration acid with HF to be used (Krea and Khalaf, 2000; Harvinderpal et al., 2004).
Fig. 5. SEM images of calcium sulphate prepared in different conditions. 343 K, 4:1 ratio of liquid to solid, 21% P2O5 of phosphoric acid and 1.05 stoichiometric ratio of H2SO4 to CaO. a. Blank sample, b. 0.3%NS-2 with 4000 molecule weight, c. 0.2% NS-2 with 4000 molecule weight, 1.1 stoichiometry of H2SO4 to CaO, d. 1%IA-1.
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
45
Table 6 Effect of different extractants on degree extraction of RE. Extractants
0.6 M D2EHPA + 0.2 M
25% P2O5 10% P2O5
1.5 M D2EHPA + 0.2 M
–
TBP
Cyanex 923
TOA
–
TBP
Cyanex 923
– 32.5%
– 23.1%
– 21.9%
– 7.8%
8.4% 65. 5%
6.0% –
6.6% –
298K, phosphoric acid with 1 g/L REO, O/W = 1.
In this work, D2EHPA was chosen as the best extractant and the effect of different mixtures of D2EHPA with TBP, TOA and Cyanex 923 on the extraction of rare earths was examined as shown in Table 6. With all the mixtures examined, the extraction of RE was less than that with D2EHPA alone, which indicates that there was an anti synergistic effect between the neutral organo-phosphorus extractants and D2EHPA. 3.2.2. Influence of D2EHPA concentration In the case of extraction by acidic extractants the reaction may be written as M
n+
P
P
+ q HL⇌ MLn ðHLÞq−n + nH P
þ
ð1Þ P
where HL is the extractant in the organic phase; and MLn ðHLÞq−n represents the metal-extractant complex in the organic phase. The expression for distribution coefficient D is P
D=
MLn ðHLÞq−n ½Mn+
P
= Kex
½ HLq ½H þ n
ð2Þ
where P
Kex =
þ n
½ MLn ðHLÞq−n ½H P
½Mn+ ½ HLq
ð3Þ
Therefore, in general, D increases with increasing extractant concentration and it is preferable to have a high extractant concentration for better RE extraction. However, high concentrations of D2EHPA increase the viscosity of the organic phase and therefore decrease the coefficient of mass transfer. Fig. 6 shows the rare earth extraction (RE%) increases as a function of D2EHPA concentration and reaches 80% with 2.0 M D2EHPA. It can be seen that the extraction efficiency for the elements by D2EHPA is in the order La b Ce b Pr b Nd b Sm b Gd b Dy b Y which rises with increasing atomic number (except for Y) and is consistent with the
Fig. 6. Effect of extractant concentration on degree of extraction of RE 298 K, phosphoric acid with 10% P2O5 and 1 g/L REO, O/W = 1.
Fig. 7. Effect of phosphoric acid concentration on the distribution coefficient of RE 298 K, phosphoric acid with 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
results reported by Sato (1989). However, high extractant concentrations are also favour impurities entering into the organic phase which lowers the separation between the lanthanides and impurities. 3.2.3. Effect of phosphoric acid concentration The effect of phosphoric acid concentration on the extraction of rare earths by 1.5 mol/L D2EHPA was then studied. The acidity of the feed was varied and the rare earths concentration was maintained at 1.0 g/L by spiking with a standard rare earths solution. The log–log plot of D vs. acid concentration (Fig. 7) shows that the distribution coefficient and extraction of RE decreases with increasing phosphoric acid concentration. This is in accordance with the results obtained by Krea using the DOPPA (Krea and Khalaf, 2000). Therefore relatively low concentrations of phosphoric acid are required for efficient RE extraction. Generally, after the purification of wet process phosphoric acid by solvent extraction, the content of P2O5 in raffinate varies from 5% to 10% (Becker, 1983; Hannachi et al., 2007). Hence, the concentration of phosphoric acid was chosen as 10% P2O5 for further solvent extraction studies. 3.2.4. Effect of O/A phase ratio The O/A phase ratio was varied between 1/5 and 5/1 under the following operating conditions: [D2EHPA] = 1.5 mol/L; phosphoric
Fig. 8. Effect of phase ratio on degree extraction of RE 298K, phosphoric acid with 10% P2O5 and 1 g/L REO, 1.5 mol/L D2EHPA.
46
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
ing the stoichiometry of the extracted species. The distribution coefficient of Eq. (2) can be written in a linear form as: P
þ
log D = log Kex + q log½HL−n log½H
Fig. 9. Effect of impurity ions on degree extraction of RE 298 K, phosphoric acid with 10% P2O5 and 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
acid with 10% P2O5 and 1 g/L REO at 298 K and the results are shown in Fig. 8. Increasing the O/A phase ratio from 1/5 to 2/1 led to an increase in the percentage of extracted RE from 37% to 82%, but beyond 2/1, the phase ratio had less effect on the RE extraction. 3.2.5. Effect of impurities The impurity ions in wet-process phosphoric acid mainly include Al3+, Fe3+, Ca2+, Mg2+. During the extraction of lanthanides from phosphoric acid, the impurities are also extracted into the organic phase. In this work, the influence of Al3+, Fe3+, Ca2+, Mg2+ concentration on RE extraction was studied with the concentration range between 0 and 3.0 g/L. The results in Fig. 9 show that the degree of extraction of rare earths decreases very little with increasing Al3+, Ca2+ and Mg2+ concentration but it is evident that Fe3+ has a great effect on RE extraction and there is significant competitive extraction between Fe3+ and rare earth metal ions. 3.3. Mechanism of extraction reaction 3.3.1. Nature of extracted RE-organic complexes Since the species extracted cannot be isolated, an indirect approximation method based on slope-analysis is used for determin-
Fig. 10. Effect of extractant concentration on the distribution coefficient of RE 298 K, phosphoric acid with 10% P2O5 and 1 g/L REO, O/W = 1. (In Figs. 10–12, the extraction efficiency for elements by D2EHPA is in order La b Ce b Pr b Nd b Sm b Gd b Dy b Y, which rises with increasing atomic number of elements (except for Y)).
ð4Þ
According to this relationship, the coefficient q gives the number of extractant molecules, which are bound to the metallic cation. Graphically this represents the slope of a straight line plot of log [D] vs. log ½ HL for constant [H+]. It must be taken into account that organophosphoric acids form dimers at low acidity in non-polar solvents (Xu, 1995). Sato (1989) extracted RE3+ from hydrochloric or nitric acid solutions with D2EHPA and EHEHPA and found that at higher acid concentrations the slope of the straight line plot of log D vs. pH varied between 0 and 3, which suggested that at high acidity these reagents extracted acid by a solvating mechanism. In this work, the plot of log [D] vs. log [D2EHPA] at a phosphoric acid content of 10% P2O5 gave a straight line with an average slope of q close to 2 as shown in Fig. 10. The log–log plot of D and phosphoric acid concentration, shown in Fig. 11, also gives a linear relationship with slope of −3 which indicates that 3 mol of acid are liberated for extraction of 1 mol of rare earths. Based on the above findings, the most plausible extraction equilibrium can be written as RE
3+
P
P
+ H3 PO4 + 2ð HLÞ2 ⇌ REðH 2 PO4 ÞL2 ðHLÞ2 + 3H
þ
ð5Þ
where HL = D2EHPA. It is known that D2EHPA at high acidity extracts H3PO4, by a solvation extraction mechanism. Whilst extraction at low D2EHPA concentration occurs by a cation exchange reaction and releases H+. On the other hand, the phosphoric acid is a moderately strong acid (Ka1 =7.52×10− 3) and 3+ the partial dissociation of H2PO− 4 has a strong binding capacity with RE and is extracted into the organic phase as a complex. 3.3.2. Effect of temperature Fig. 12 shows that the extraction of rare earths by 1.5 mol/L D2EHPA from phosphoric acid containing 1 g/L REO decreases with temperature due to the increasing pKa value of D2EHPA. According to the Van't Hoff equation Δ logD ΔH =− Δð1 = TÞ 2:303R
ð6Þ
The enthalpy change was determined between −4.7 kJ/mol and −9.0 kJ/mol, which indicates that it is an exothermic reaction.
Fig. 11. Effect of phosphoric acid concentration on the distribution coefficient of RE 298 K, phosphoric acid with 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
Fig. 12. Effect of temperature on the distribution coefficient of RE phosphoric acid with 10% P2O5 and 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
4. Conclusions More effective recovery of rare earths from phosphorite ores and RE enrichment in phosphoric acid was obtained by leaching at lower temperatures, with larger liquid/solid ratio and higher concentration of phosphoric acid. Non-ionic surfactant additives led to a significant increase in RE leaching efficiency up to 75% due to enhanced gypsum crystal growth and less adsorption of RE. D2EHPA in kerosene extracted rare earths from the phosphoric acid solution when there was a high concentration of extractant, large phase ratio, lower temperature and relatively low concentration of phosphoric acid. When the amount of Fe3+ ions increased, the degree of RE extraction decreased due to competitive extraction between Fe3+ and rare earth cations, but the decrease was insignificant for the addition of Ca2+, Mg2+ and Al3+ ions. The extraction reaction is exothermic and may be written as: RE
3+
P
P
+ H3 PO4 + 2ð HLÞ2 ⇌ REðH2 PO4 ÞL2 ðHLÞ2 + 3H
þ
where HL is D2EHPA. Acknowledgements The National High Technology Research and Development Program of China (863 Program: 2006AA06Z126) is thanked for the financial aid. Thanks also to Zhaowu Zhu, (CSIRO Minerals, Australia) for his valuable suggestions. References Amjad, Z., 1987. Kinetics of crystal growth of calcium sulfate dihydrate. The influence of polymer composition, molecular weight, and solution pH. Canadian Journal of Chemistry 66, 1529–1536. Becker, P., 1983. Phosphates and Phosphoric Acid: Raw Materials, Technology, and Economics of the Wet Process. Marcel Dekker, NewYork. Bunu, F., Dumitrescu, R., 1992. Simultaneous extraction of rare earth elements and uranium from phosphoric acid. Hydrometallurgy 28 (3), 331–338. Bunu, F., Miu, I., Dumitrescut, R., 1994. Simultaneous recovery and separation of uranium and rare earths from phosphoric acid in a one-cycle extraction-stripping process. Hydrometallurgy 35, 375–389. Byrne, R.H., Kim, K.H., 1993. Rare earth precipitation and coprecipitation behavior. The limiting role of PO3− 4 on dissolved rare earth concentrations in seawater. Geochimica et Cosmochimica Acta 57, 519–526.
47
Cetiner, Z.S., Wood, S.A., Gammons, C.H., 2005. The aqueous geochemistry of the rare earth elements. Part XIV. The solubility of rare earth element phosphates from 23 to 150 °C. Chemical Geology 217, 147–169. El Moussaouiti, M., Boistelle, R., Bouhaouss, A., Klein, J.P., 1996. Agglomeration kinetics of calcium sulphate hemihydrate crystals in sulpho-phosphoric solutions. Journal of Crystal Growth 169, 118–123. El Moussaouiti, M., Boistelle, R., Bouhaouss, A., Klein, J.P., 1997. Crystallization of calcium sulphate hemihydrate in concentrated phosphoric acid solutions. Chemical Engineering Journal 68, 123–130. El-Shall, H., Rashad, M.M., Abdel-Aal, E.A., 2005. Effect of cetyl pyridinium additive on crystallization of gypsum in phosphoric and sulfuric acids medium. Crystallisation Research and Technology 40 (9), 860–866. Firsching, F.H., Brune, S.N., 1991. Solubility products of the trivalent rare-earth phosphates. Journal of Chemical Engineering 36, 93–95. Free, M.L., Zhu, J.S., Moudgil, B.M., 1999. The effect of PEO and organic sulfonates in enhancing phosphogypsum filtration. International Journal of Mineral Processing 57, 25–42. Hannachi, A., Habaili, D., Chtara, C., Ratel, A., 2007. Purification of wet process phosphoric acid by solvent extraction with TBP and MIBK mixtures. Separation and Purification Technology 55 (2), 212–216. Harvinderpal, S., Mishra, S.L., Vijayalakshmi, R., 2004. Uranium recovery from phosphoric acid by solvent extraction using a synergistic mixture of di-nonyl phenyl phosphoric acid and tri-n-butyl phosphate. Hydrometallurgy 73, 63–70. Hurst, F.J., Crouse, D.J., 1973. Reductive stripping process for the recovery of uranium from wet-process phosphoric acid. U.S. Patent 3711591. Hurst, F.J., Crouse, D.J., Brown, K.B., 1969. Solvent Extraction of Uranium From WetProcess Phosphoric Acid. Oak Ridge National Laboratory. Report ORNLTM-2522. Hurst, F.J., Crouse, D.J., Brown, K.B., 1972. Recovery of uranium from wet-process phosphoric acid. Industrial Engineering and Chemical Process Design Developments 11 (1), 122–128. Ilyin, Andrei V., 1998. Phosphorites of the Russian Craton. Earth-Science Reviews 45 (1–2), 89–101. Jarosinski, A., Kowalczyk, J., Cz., Mazanek, 1993. Development of the Polish wasteless technology of apatite phosphogypsum utilization with recovery of rare earths. Journal of Alloys and Compounds 200, 147–150. Jorjani, E., Bagherieh, A.H., Mesroghli, Sh., Chehreh Chelgani, S., 2008. Prediction of yttrium, lanthanum, cerium, and neodymium leaching recovery from apatite concentrate using artificial neural networks. Journal of University of Science and Technology Beijing 15 (4), 367–374. Koopman, C., Witkamp, G.J., 2000. Extraction of lanthanides from the phosphoric acid production process to gain a purified gypsum and a valuable lanthanide by-product. Hydrometallurgy 58, 51–60. Krea, M., Khalaf, H., 2000. Liquid–liquid extraction of uranium and lanthanides from phosphoric acid using a synergistic DOPPA–TOPO mixture. Hydrometallurgy 58, 215–225. Lededev, V.N., Lokshin, E.H.P., 2000. Method of isolation of rare earth elements from phosphogypsum. RU Patent 2158317 (in Russian). Lededev, V.N., Masloboev, V.A., Lokshin, E.H.P., Popova, L.A., Sergeeva, S.D., 2000. Method of extraction of rare earth element phosphates from apatite concentrate. RU Patent 2120408 (in Russian). Liu, X., Byrne, R.H., 1997. Rare earth and yttrium phosphate solubilities in aqueous solution. Geochimica et Cosmochimica Acta 61, 1625–1633. Lokshin, E.H.P., Vershkova, J.U.A., Kalinnikov, V.T., Tareeva, O.A., Ivlev, K.G., Fedorov, S.G., et al., 2004. Method of recovering rare earth minerals from phosphogypsum. RU Patent 2225892. (in Russian). Pessac, J.F., Lambert, A.P., Tognet, J.P., 1987. Essentially complete recovery of uranium, yttrium, thorium and rare earth values from phosphate rock during wet-process production of phosphoric acid there from. U.S. Patent 4636369. Preston, J.S., Cole, P.M., Craig, W.M., Feather, A.M., 1996. The recovery of rare earth oxides from a phosphoric acid by-product. Hydrometallurgy 41, 1–19. Rickelton, W.A., 1988. Recovery of uranium from wet process phosphoric acid using assymetrical phosphine oxides. U.S. Patent 4778663. Samir, I., Abu-Eishah, Nizar, M., Abu-Jabal, 2001. Parametric study on the production of phosphoric acid by the dihydrate process. Chemical Engineering Journal 81, 231–250. Sangwal, K., 1993. Effect of impurities on the processes of crystal growth. Journal of Crystal Growth 128 (1–4), 1236–1244. Sangwal, K., 1996. Effects of impurities on crystal growth processes. Progress in Crystal Growth and Characterization of Materials 32 (1–3), 3–43. Sangwal, K., 2002. On the nature of supersaturation barriers observed during the growth of crystals from aqueous solutions containing impurities. Journal of Crystal Growth 242 (1–2), 215–228. Sato, T., 1989. Liquid–liquid extraction of rare-earth elements from aqueous acid solutions by acid organophosphorus compounds. Hydrometallurgy 22 (1–2), 121–140. Smith, B.R., Alexander, A.E., 1970. The effect of additives on the process of crystallization. Journal of Colloid and Interface Science 341, 80–90. Tadros, M.E., Mayes, I., 1979. Linear growth rates of calcium sulfate dihydrate crystals in the presence of additives. Journal of Colloid and Interface Science 722, 245–254. Xu, G.X., 1995. Rare Earth, Second Edition. Metallurgical Industry Press, Beijing. (in Chinese). Zhang, Y.Q., Li, J.N., Huang, X.W., Wang, C.M., Zhu, Z.W., Zhang, G.C., 2008. Synergistic extraction of rare earths by mixture of HDEHP and HEH/EHP in sulfuric acid medium. Journal of Rare Earths 5, 688–692.
Contents lists available at ScienceDirect
Hydrometallurgy j o u r n a l h o m e p a g e : w w w. e l s e v i e r. c o m / l o c a t e / h y d r o m e t
Recovery of rare earths from wet-process phosphoric acid Liangshi Wang, Zhiqi Long ⁎, Xiaowei Huang, Ying Yu, Dali Cui, Guocheng Zhang National Engineering Research Center for Rare Earth Materials, General Research Institute for Nonferrous Metals, and Grirem Advanced Materials Co. Ltd., Beijing, 100088, China
a r t i c l e
i n f o
Article history: Received 30 July 2009 Received in revised form 23 November 2009 Accepted 23 November 2009 Available online 1 December 2009 Keywords: Rare earths extraction Phosphorite Wet-process phosphoric acid D2EHPA
a b s t r a c t Phosphorite ores are a potential resource of rare earths (RE) as well as phosphate so the recovery of rare earths during the wet processing of phosphoric acid is important. This study investigates the influence of operating conditions and crystal modifiers on the leaching of RE and the solvent extraction of RE with organo-phosphorus reagents. The results indicate that lower temperature, higher concentration of phosphoric acid and larger liquid/solid ratio are beneficial to RE enrichment in the phosphoric acid. Surfactant additives which enhance the crystal growth of gypsum also enhanced RE leach recovery about 75% under optimized conditions. Studies on the solvent extraction of RE found that D2EHPA mixtures with neutral organo-phosphorus reagents were antagonistic and that Fe3+ competed strongly over RE. Higher D2EHPA concentration, larger phase ratio, lower temperature and lower phosphoric acid concentration increased the RE extraction efficiency. A negative enthalpy change was found indicating an exothermic extraction reaction. © 2009 Elsevier B.V. All rights reserved.
1. Introduction Phosphorite is a rare earths bearing mineral that may occur as the main component in individual deposits as well as associated with other minerals, such as apatite (Jorjani et al., 2008). The total reserves of phosphorite ores in the world is about 100 billion tons (Ilyin, 1998; Becker, 1983) which on average contain 0.5‰ RE. Table 1 showed the composition of rare earths in various phosphorite reserves. In particular, Russia has abundant reserves of high grade phosphorite ore at Kola which contain up to 1% rare earths and is considered as the primary rare earths resource in the country. Phosphorite is also relatively abundant in China where the Zhijin phosphorite deposit in Guizhou province has reserves of more than 1.6 billion tons containing more than 1.4 million tons. Another deposit is in Fanshan. Therefore it is important to study the recovery of the rare earths from phosphorite which is a phosphate mineral suitable for the manufacture of phosphoric acid. Wet processing, using the digestion of phosphate rocks by sulphuric acid, is the most common process used to produce phosphoric acid. The overall and simplified reaction between calcium phosphate minerals and sulphuric acid (Becker, 1983) can be written as: Ca5 FðPO4 Þ3 þ 5H2 SO4 þ xH2 O ¼ 3H3 PO4 þ 5CaSO4 ·x=5H2 O↓ þ HF↑ where x is 0 for calcium sulphate anhydrate, 0.5 for hemihydrate and 2 for dihydrate, i.e. gypsum. Studies show that the dominant species of
⁎ Corresponding author. Tel.: + 86 10 82241188; fax: + 86 10 62355405. E-mail address: [email protected] (Z. Long). 0304-386X/$ – see front matter © 2009 Elsevier B.V. All rights reserved. doi:10.1016/j.hydromet.2009.11.017
rare earths during the wet-process phosphoric acid (WPA) is phosphate, which has low solubility (Xu, 1995). Therefore, rare earths will be enriched in the phosphor-gypsum with only 20%–30% RE in the phosphoric acid leaching solution. In South Africa (Preston et al., 1996), it was reported that the phosphor-gypsum adsorbed 85% RE during the acidulation of Phalaborwa phosphorite with sulphuric acid. Similarly, in Poland (Jarosinski et al., 1993), it was found that 70% RE entered the calcium sulphate dihydrate (CaSO4·2H2O, DH or gypsum) during dihydrate process for the Kola phosphorite, whilst almost all the rare earth was adsorbed by the calcium sulphate hemihydrate (CaSO4·1/2H2O, HH) during the hemihydrate process. Hence, various recovery techniques of rare earths from the phosphogypsum have been developed. For instance, in Russia the rare earths were leached out from phosphor-gypsum with sulphuric acid and then the leaching solution was evaporated to crystallize rare earths or was neutralized to precipitate the rare earths (Lokshin et al., 2004; Lededev and Lokshin, 2000; Lededev et al., 2000). In Poland, it was reported that rare earths was precipitated with hydrofluoric acid or extracted with nonyl–phenyl phosphoric acid (NPPA) to obtain a 40% concentrate from the phosphor-gypsum leaching solution (Jarosinski et al., 1993). However, the process of the rare earths recovery from phosphor-gypsum is too complicated and uneconomical to be industrialized. On the other hand, rare earths can be recovered directly from phosphoric acid leaching solution and the key problem is to enhance their enrichment. Pessac et al. (1987) mentioned that during the acidulation of the phosphate rock, the leaching efficiency of RE was increased to 56% by introducing aluminum and/or iron and/or silica minerals into the slurry. However, this may lead to difficulties in subsequent stages of phosphoric acid production.
42
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
Table 1 Content of rare earth elements in various phosphate rocks (Krea and Khalaf, 2000; Koopman and Witkamp, 2000). Deposit origin
Florida Pebble
Morocco Khoribga
Algeria Djebel Onk
Russia Kola
Palfos
Kovdor
Jordan
Israel
Lanthanides%
0.29
0.14–0.16
0.13–0.18
0.84–0.88
0.48
0.14
0.01
0.01
Table 2 Content of rare earth elements in various phosphorite phosphate rocks (ppm). Deposit origin
Y2O3
La2O3
CeO2
Pr6O11
Zhijin Fanshan
434.2 248.9
273.2 1034
162.7 2170
45.8 285.0
198.2 1201
35.2 229.6
8.5 56.9
Deposit origin
Ho2O3
Er2O3
Yb2O3
Lu2O3
Sc2O3
∑ REO
LREO
HREO
6.8 –
1295.0 5513.2
723.7 4976.1
571.3 537.1
Zhijin Fanshan
7.4 9.1
20.4 19.8
Tm2O3 2.3 1.9
Nd2O3
12.0 9.7
Since 1950s, solvent extraction methods have been studied to recover trace elements from wet-process phosphoric acid leach solutions, focusing on the recovery of uranium. In 1967, Oak Ridge National Laboratory (ORNL) recovered uranium using a synergistic extractant of DEHPA-TOPO (tri-n-octyl-phosphine oxide), and carried out large scale production (Hurst et al., 1969, 1972; Hurst and Crouse, 1973). Afterwards, liquid Cyanex 923 was used to replace solid TOPO as this offers a materials handling advantage (Rickelton, 1988). Recently, Institute of Atomic Physics in Romania (Bunu and Dumitrescu, 1992; Bunu et al., 1994), University of Blida in Algeria (Krea and Khalaf, 2000) and Rare Earth Development Section of BARC in India (Harvinderpal et al., 2004) studied the liquid–liquid extraction of uranium and lanthanides from wet-process phosphoric acid using a synergistic acidic and neutral organo-phosphorus extractants. As indicated, all the research focused on the uranium recovery. Therefore, this study focuses on both the leaching and solvent extraction method to recover rare earths from phosphorite ores during wet-process phosphoric acid operation. 2. Experimental 2.1. Reagents and solutions The phosphorite ore (Fanshan Phosphorite) used in this study consisted of 33.8% P2O5, 51.0% CaO, and 0.55% REO, 0.8% MgO, 1.5% FeO, 0.5% Al2O3. The rare earth element content is shown in Table 2. The di-(2-ethylhexyl) phosphoric acid (D2EHPA) was purified by nantokite before use (purity N99%) (Zhang et al., 2008). Cyanex 923 was supplied by CYTEC Canada INC. The extractants were diluted with kerosene to the required concentration. Kerosene was sulfonated before use. Stock solutions of each rare earth were prepared by dissolving their oxide with a purity N99.9% in concentrated sulphuric acid and diluting with deionized water. Unless indicated otherwise, all reagents were reagent-grade.
1.4 1.22
Sm2O3
Eu2O3
Gd2O3
Tb4O7
42.9 166.6
5.9 16.9
Dy2O3 35.5 62.8 LREO/HREO 1.27 9.26
by mass balance. RE extraction efficiency was obtained from RE concentrations extracted. 3. Results and discussion 3.1. Leaching of phosphorite concentrate 3.1.1. Influence of WPA process conditions on the RE leaching efficiency Generally, the dominant species of rare earths in phosphoric acid is RE phosphate which has a low solubility, e.g. the solubility product (Ksp) of the LaPO4 is 4 × 10− 23 and that of CePO4 is 1.6 × 10− 23 at 298 K. Nevertheless, the solubility of rare earths in phosphoric acid can be increased by increasing the acid concentration which is a benefit for rare earths leaching (Cetiner et al., 2005 and references therein). A series of experiments performed with increasing concentrations of phosphoric acid ranging from 5% to 25% (by the weight of P2O5) found a significant increase in RE leaching efficiency. The leaching efficiency for the elements is in the order Lab Ceb Pr b Ndb Sm N Gd b Dyb Y which rises with increasing atomic number (except for Y) and is in reasonable agreement with the solubility products of the RE phosphates (Firsching and Brune, 1991; Byrne and Kim, 1993; Liu and Byrne, 1997). As shown in Fig. 1, the RE leaching efficiency increased sharply from 32% when the phosphoric acid contained 5% P2O5, to 71% with phosphoric acid containing 25% P2O5. In addition, as the concentration of phosphoric acid increased, the energy consumption reduced. However, substitution of SO2− ions by HPO2− ions in the gypsum crystal lattice also increases, 4 4
2.2. Leaching studies The wet-process phosphoric acid production was simulated in a batch reactor and the rare earth elements in the phosphoric acid and gypsum were determined by ICP (Inductively Coupled Plasma). The morphology of the gypsum was recorded by scanning electron microscopy, after washing by ethanol and dried at 338 K for 3 h. 2.3. Solvent extraction studies The organic and the aqueous phases were shaken in a separation funnel at constant temperature to reach equilibrium. After phase separation, take aliquot of aqueous solution to analyze rare earths with ICP. The RE concentrations in the organic phase were calculated
Fig. 1. Effect of concentration of phosphoric acid on RE leaching efficiency 343 K, 0.2% NS-2 with 4000 molecule weight, 4:1 ratio of liquid to solid, 1.05 stoichiometric ratio of H2SO4 to CaO.
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
43
Fig. 2. Effect of stoichiometry of H2SO4 to CaO on RE leaching efficiency 343 K, 0.2% NS-2 with 4000 molecule weight, 4:1 ratio of liquid to solid, 21% of phosphoric acid.
Fig. 4. Effect of ratio of liquid to solid on RE leaching efficiency 343 K, 21% P2O5 of phosphoric acid, 0.2% NS-2 with 4000 molecule weight, 1.05 stoichiometric ratio of H2SO4 to CaO.
which results in the loss of P2O5. Thus, about 20%–25% phosphoric acid is optimum for the wet-process phosphoric acid production. According to Becker (1983), sulphuric acid plays an essential role in the crystallization process of dehydrate. With better crystallization, there is less eutectic crystallization and adsorption between lanthanide phosphates and calcium sulphate. Therefore special attention was paid to the effect of the stoichiometric ratio of H2SO4 to CaO that is present in the phosphorite. Fig. 2 shows the effect of varying the mole ratio of H2SO4:CaO from 1.0:1.0 to 1.2:1.0 which is comparable to the industrial mole ratio (0.9:1.0–1.2:1.0), as dictated by the composition and nature of the phosphate rock used (Becker, 1983). The results reveal that RE leaching passes through a maximum when the stoichiometric ratio is about 1.10:1.0 and that excess sulphuric acid affects the solubility and supersaturation of gypsum (El Moussaouiti et al., 1997; Samir et al., 2001). Clearly, the RE leaching efficiency can be enhanced by controlling the crystallization rate and the crystal shape of the calcium sulphate. When the stoichiometric ratio is 1.1, the calcium sulphate is formed with large clusters as shown in Fig. 5(c), which is easily filtered and scrubbed. This observation is consistent with the results reported by El Moussaouiti et al. (1996) who found that calcium sulphate
dihydrate crystals agglomerate more or less depending on the excess of sulphuric acid in sulpho-phosphoric acid solutions. Generally, during the wet-process phosphoric acid production, high sulphuric acid concentration leads to encapsulation and passivation of phosphorite (Becker, 1983). Therefore, the appropriate concentration of sulphuric acid increases the crystallinity of the calcium sulphate to enhance the RE leaching efficiency and at the same time, reduces the substitution of SO2− by HPO2− in the calcium sulphate crystal to 4 4 enhance the productivity of phosphoric acid. A series of experiments was carried out in the temperature range 343 K–358 K to study the effect of temperature on RE leaching efficiency. Lower temperatures lead to stabilized foams, resulting in substantially increased consumption of defoamer. While higher temperatures lead to the corrosion of equipment and to higher solubility of impurities which determine the quality of the phosphoric acid. The results obtained reveal that the RE leaching efficiency rapidly decreases with a small increase of temperature (Fig. 3). It reached 61.4% at 343 K and decreased to 25.1% at 348 K and reduced further with increasing temperature. This is attributed to the sharp decrease in solubility of rare earths phosphate in the phosphoric acid with the increasing temperature (Cetiner et al., 2005). However, in industry, the crystallization conditions would deteriorate and the yield of the phosphoric acid would be lower as the temperature drops below 343 K. Thus, 343 K is considered to be the optimum temperature. A series of experiments was then performed at liquid/solid ratios ranging from 2:1 to 5:1 L/kg at 343 K, with 21% of phosphoric acid and 1.05 stoichiometric ratio of sulphuric acid. The results in Fig. 4 indicate that an increase in the liquid/solid ratio leads to a significant increase in RE leaching efficiency. However, a higher liquid/solid ratio leads to a Table 3 Effect of additives on RE leaching efficiency.
Fig. 3. Effect of temperature on RE leaching efficiency 0.2% NS-2 with 4000 molecule weight, 4:1 ratio of liquid to solid, 21% P2O5 of phosphoric acid, 1.05 stoichiometric ratio of H2SO4 to CaO.
No.
Active additives
RE leaching efficiency (%)
0419-1 0423-1 0423-2 0424-1 0802-2 0427-1 0427-2 0621-1 0716-2
– 0.5%IS-1 0.2%IS-2 0.2%IS-3 0.2%NS-1 0.1%CS-1 0.2%NS-2 0.2%NS-3 1%IA-1
19.8 17.4 23.5 21.4 53.0 29.2 61.4 40.7 60.5
343 K, 4:1 liquid to solid, 21% P2O5 of phosphoric acid and 1.05 stoichiometric ratio H2SO4 to CaO, the content of additives is measured with phosphorite weight percent. NS-2 with 4000 molecule weight.
44
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
Table 4 Effect of NS-2 with 4000 molecule weight on RE leaching efficiency. No.
Amount of NS-2/%
RE leaching efficiency (%)
0626-1 0427-2 0802-1 0626-2
0.1 0.2 0.3 0.5
46.6 61.4 75.2 62.4
343 K, 4:1 liquid to solid, 21% P2O5 of phosphoric acid and 1.05 stoichiometric ratio H2SO4 to CaO.
Table 5 Effect of molecule weight of NS-2 on RE leaching efficiency. No.
Molecule weight of NS-2
RE leaching efficiency (%)
0705-1 0802-1 0705-2
400 4000 20,000
61.2 75.2 73.6
343 K, 4:1 liquid to solid, 21% P2O5 of phosphoric acid, 1.05 stoichiometric ratio H2SO4 to CaO and 0.3% NS-2.
lower RE and P205 content in the pregnant liquor, which subsequently influences further processing. Therefore the optimal liquid/solid ratio is 4:1 L/kg, which gives a 61.4% RE leaching efficiency. 3.1.2. Effect of additives on RE leaching efficiency Most of the fundamental information in the literature regarding the use of additives is directed towards the inhibition of gypsum crystal growth (Smith and Alexander, 1970; Tadros and Mayes, 1979; Amjad, 1987). The additives affected the crystal growth mainly by entering into the crystal lattice, being selectively adsorbed on the interface or changing the crystal surface energy (Sangwal, 1993, 1996, 2002). In this study, the rare earths in the leach solution were enriched by adding non-ionic surfactant NS-1, NS-2 and inorganic additive IA-1, as shown in Table 3. This was mainly because the surfactant increases the aspect ratio of the formed crystals, decreases the induction time, decreases the nucleation rate, increases surface
energy and decreases the percentage of fine crystals (El-Shall et al., 2005). Based on the significant impact of NS-2 on the RE enrichment in phosphoric acid during the production process, the effect of concentration and molecular weight of NS-2 on RE leaching efficiency was studied. The results in Tables 4 and 5 show that the RE leached increases with increasing concentration until a maximum around 75% is reached with 0.3% NS-2; and was optimum with molecular weight of 4000. It appears that NS-2 delayed the onset of nucleation, as evidenced by turbidity, and inhibits nucleation, leading to fewer nuclei. Reduced numbers of nuclei lead to fewer, but larger crystals, as shown in Fig. 5(b), and higher RE leaching efficiency. This observation is consistent with the results in the gypsum growth inhibition literature which shows that the ability of polymers to inhibit crystal growth decreases with increasing molecular weight (Amjad, 1987; Free et al., 1999). 3.2. Solvent extraction for RE recovery 3.2.1. Selection of extractant Recovery of rare earths from wet-process phosphoric acid by solvent extraction is carried out in highly acidic systems. Generally, acid and neutral organo-phosphorus extractants were used with formula of (RO)3PO or R3PO with the extraction ability order of R3PON R2(RO)PON R (RO)2PON (RO)3PO. In this work, Cyanex 923 (a trialkyl-phosphine oxide) was first chosen, but for different phosphoric acid concentrations less than 10% extraction of rare earths was achieved by Cyanex 923. The extraction efficiency of acid organo-phosphorus extractants decreases with the increase of C–P bond energy and the existence of the RO branched chain, while the conjugative effect of a benzene ring is beneficial for extraction. Furthermore, acid organo-phosphorus extractants which have a low pKa value can be used in highly acidic systems, such as D2EHPA, P215, P406, DOPPA and DNPPA. However, uranium and lanthanides are difficult to strip from DOPPA, DNPPA, and need high concentration acid with HF to be used (Krea and Khalaf, 2000; Harvinderpal et al., 2004).
Fig. 5. SEM images of calcium sulphate prepared in different conditions. 343 K, 4:1 ratio of liquid to solid, 21% P2O5 of phosphoric acid and 1.05 stoichiometric ratio of H2SO4 to CaO. a. Blank sample, b. 0.3%NS-2 with 4000 molecule weight, c. 0.2% NS-2 with 4000 molecule weight, 1.1 stoichiometry of H2SO4 to CaO, d. 1%IA-1.
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
45
Table 6 Effect of different extractants on degree extraction of RE. Extractants
0.6 M D2EHPA + 0.2 M
25% P2O5 10% P2O5
1.5 M D2EHPA + 0.2 M
–
TBP
Cyanex 923
TOA
–
TBP
Cyanex 923
– 32.5%
– 23.1%
– 21.9%
– 7.8%
8.4% 65. 5%
6.0% –
6.6% –
298K, phosphoric acid with 1 g/L REO, O/W = 1.
In this work, D2EHPA was chosen as the best extractant and the effect of different mixtures of D2EHPA with TBP, TOA and Cyanex 923 on the extraction of rare earths was examined as shown in Table 6. With all the mixtures examined, the extraction of RE was less than that with D2EHPA alone, which indicates that there was an anti synergistic effect between the neutral organo-phosphorus extractants and D2EHPA. 3.2.2. Influence of D2EHPA concentration In the case of extraction by acidic extractants the reaction may be written as M
n+
P
P
+ q HL⇌ MLn ðHLÞq−n + nH P
þ
ð1Þ P
where HL is the extractant in the organic phase; and MLn ðHLÞq−n represents the metal-extractant complex in the organic phase. The expression for distribution coefficient D is P
D=
MLn ðHLÞq−n ½Mn+
P
= Kex
½ HLq ½H þ n
ð2Þ
where P
Kex =
þ n
½ MLn ðHLÞq−n ½H P
½Mn+ ½ HLq
ð3Þ
Therefore, in general, D increases with increasing extractant concentration and it is preferable to have a high extractant concentration for better RE extraction. However, high concentrations of D2EHPA increase the viscosity of the organic phase and therefore decrease the coefficient of mass transfer. Fig. 6 shows the rare earth extraction (RE%) increases as a function of D2EHPA concentration and reaches 80% with 2.0 M D2EHPA. It can be seen that the extraction efficiency for the elements by D2EHPA is in the order La b Ce b Pr b Nd b Sm b Gd b Dy b Y which rises with increasing atomic number (except for Y) and is consistent with the
Fig. 6. Effect of extractant concentration on degree of extraction of RE 298 K, phosphoric acid with 10% P2O5 and 1 g/L REO, O/W = 1.
Fig. 7. Effect of phosphoric acid concentration on the distribution coefficient of RE 298 K, phosphoric acid with 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
results reported by Sato (1989). However, high extractant concentrations are also favour impurities entering into the organic phase which lowers the separation between the lanthanides and impurities. 3.2.3. Effect of phosphoric acid concentration The effect of phosphoric acid concentration on the extraction of rare earths by 1.5 mol/L D2EHPA was then studied. The acidity of the feed was varied and the rare earths concentration was maintained at 1.0 g/L by spiking with a standard rare earths solution. The log–log plot of D vs. acid concentration (Fig. 7) shows that the distribution coefficient and extraction of RE decreases with increasing phosphoric acid concentration. This is in accordance with the results obtained by Krea using the DOPPA (Krea and Khalaf, 2000). Therefore relatively low concentrations of phosphoric acid are required for efficient RE extraction. Generally, after the purification of wet process phosphoric acid by solvent extraction, the content of P2O5 in raffinate varies from 5% to 10% (Becker, 1983; Hannachi et al., 2007). Hence, the concentration of phosphoric acid was chosen as 10% P2O5 for further solvent extraction studies. 3.2.4. Effect of O/A phase ratio The O/A phase ratio was varied between 1/5 and 5/1 under the following operating conditions: [D2EHPA] = 1.5 mol/L; phosphoric
Fig. 8. Effect of phase ratio on degree extraction of RE 298K, phosphoric acid with 10% P2O5 and 1 g/L REO, 1.5 mol/L D2EHPA.
46
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
ing the stoichiometry of the extracted species. The distribution coefficient of Eq. (2) can be written in a linear form as: P
þ
log D = log Kex + q log½HL−n log½H
Fig. 9. Effect of impurity ions on degree extraction of RE 298 K, phosphoric acid with 10% P2O5 and 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
acid with 10% P2O5 and 1 g/L REO at 298 K and the results are shown in Fig. 8. Increasing the O/A phase ratio from 1/5 to 2/1 led to an increase in the percentage of extracted RE from 37% to 82%, but beyond 2/1, the phase ratio had less effect on the RE extraction. 3.2.5. Effect of impurities The impurity ions in wet-process phosphoric acid mainly include Al3+, Fe3+, Ca2+, Mg2+. During the extraction of lanthanides from phosphoric acid, the impurities are also extracted into the organic phase. In this work, the influence of Al3+, Fe3+, Ca2+, Mg2+ concentration on RE extraction was studied with the concentration range between 0 and 3.0 g/L. The results in Fig. 9 show that the degree of extraction of rare earths decreases very little with increasing Al3+, Ca2+ and Mg2+ concentration but it is evident that Fe3+ has a great effect on RE extraction and there is significant competitive extraction between Fe3+ and rare earth metal ions. 3.3. Mechanism of extraction reaction 3.3.1. Nature of extracted RE-organic complexes Since the species extracted cannot be isolated, an indirect approximation method based on slope-analysis is used for determin-
Fig. 10. Effect of extractant concentration on the distribution coefficient of RE 298 K, phosphoric acid with 10% P2O5 and 1 g/L REO, O/W = 1. (In Figs. 10–12, the extraction efficiency for elements by D2EHPA is in order La b Ce b Pr b Nd b Sm b Gd b Dy b Y, which rises with increasing atomic number of elements (except for Y)).
ð4Þ
According to this relationship, the coefficient q gives the number of extractant molecules, which are bound to the metallic cation. Graphically this represents the slope of a straight line plot of log [D] vs. log ½ HL for constant [H+]. It must be taken into account that organophosphoric acids form dimers at low acidity in non-polar solvents (Xu, 1995). Sato (1989) extracted RE3+ from hydrochloric or nitric acid solutions with D2EHPA and EHEHPA and found that at higher acid concentrations the slope of the straight line plot of log D vs. pH varied between 0 and 3, which suggested that at high acidity these reagents extracted acid by a solvating mechanism. In this work, the plot of log [D] vs. log [D2EHPA] at a phosphoric acid content of 10% P2O5 gave a straight line with an average slope of q close to 2 as shown in Fig. 10. The log–log plot of D and phosphoric acid concentration, shown in Fig. 11, also gives a linear relationship with slope of −3 which indicates that 3 mol of acid are liberated for extraction of 1 mol of rare earths. Based on the above findings, the most plausible extraction equilibrium can be written as RE
3+
P
P
+ H3 PO4 + 2ð HLÞ2 ⇌ REðH 2 PO4 ÞL2 ðHLÞ2 + 3H
þ
ð5Þ
where HL = D2EHPA. It is known that D2EHPA at high acidity extracts H3PO4, by a solvation extraction mechanism. Whilst extraction at low D2EHPA concentration occurs by a cation exchange reaction and releases H+. On the other hand, the phosphoric acid is a moderately strong acid (Ka1 =7.52×10− 3) and 3+ the partial dissociation of H2PO− 4 has a strong binding capacity with RE and is extracted into the organic phase as a complex. 3.3.2. Effect of temperature Fig. 12 shows that the extraction of rare earths by 1.5 mol/L D2EHPA from phosphoric acid containing 1 g/L REO decreases with temperature due to the increasing pKa value of D2EHPA. According to the Van't Hoff equation Δ logD ΔH =− Δð1 = TÞ 2:303R
ð6Þ
The enthalpy change was determined between −4.7 kJ/mol and −9.0 kJ/mol, which indicates that it is an exothermic reaction.
Fig. 11. Effect of phosphoric acid concentration on the distribution coefficient of RE 298 K, phosphoric acid with 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
L. Wang et al. / Hydrometallurgy 101 (2010) 41–47
Fig. 12. Effect of temperature on the distribution coefficient of RE phosphoric acid with 10% P2O5 and 1 g/L REO, 1.5 mol/L D2EHPA, O/W = 1.
4. Conclusions More effective recovery of rare earths from phosphorite ores and RE enrichment in phosphoric acid was obtained by leaching at lower temperatures, with larger liquid/solid ratio and higher concentration of phosphoric acid. Non-ionic surfactant additives led to a significant increase in RE leaching efficiency up to 75% due to enhanced gypsum crystal growth and less adsorption of RE. D2EHPA in kerosene extracted rare earths from the phosphoric acid solution when there was a high concentration of extractant, large phase ratio, lower temperature and relatively low concentration of phosphoric acid. When the amount of Fe3+ ions increased, the degree of RE extraction decreased due to competitive extraction between Fe3+ and rare earth cations, but the decrease was insignificant for the addition of Ca2+, Mg2+ and Al3+ ions. The extraction reaction is exothermic and may be written as: RE
3+
P
P
+ H3 PO4 + 2ð HLÞ2 ⇌ REðH2 PO4 ÞL2 ðHLÞ2 + 3H
þ
where HL is D2EHPA. Acknowledgements The National High Technology Research and Development Program of China (863 Program: 2006AA06Z126) is thanked for the financial aid. Thanks also to Zhaowu Zhu, (CSIRO Minerals, Australia) for his valuable suggestions. References Amjad, Z., 1987. Kinetics of crystal growth of calcium sulfate dihydrate. The influence of polymer composition, molecular weight, and solution pH. Canadian Journal of Chemistry 66, 1529–1536. Becker, P., 1983. Phosphates and Phosphoric Acid: Raw Materials, Technology, and Economics of the Wet Process. Marcel Dekker, NewYork. Bunu, F., Dumitrescu, R., 1992. Simultaneous extraction of rare earth elements and uranium from phosphoric acid. Hydrometallurgy 28 (3), 331–338. Bunu, F., Miu, I., Dumitrescut, R., 1994. Simultaneous recovery and separation of uranium and rare earths from phosphoric acid in a one-cycle extraction-stripping process. Hydrometallurgy 35, 375–389. Byrne, R.H., Kim, K.H., 1993. Rare earth precipitation and coprecipitation behavior. The limiting role of PO3− 4 on dissolved rare earth concentrations in seawater. Geochimica et Cosmochimica Acta 57, 519–526.
47
Cetiner, Z.S., Wood, S.A., Gammons, C.H., 2005. The aqueous geochemistry of the rare earth elements. Part XIV. The solubility of rare earth element phosphates from 23 to 150 °C. Chemical Geology 217, 147–169. El Moussaouiti, M., Boistelle, R., Bouhaouss, A., Klein, J.P., 1996. Agglomeration kinetics of calcium sulphate hemihydrate crystals in sulpho-phosphoric solutions. Journal of Crystal Growth 169, 118–123. El Moussaouiti, M., Boistelle, R., Bouhaouss, A., Klein, J.P., 1997. Crystallization of calcium sulphate hemihydrate in concentrated phosphoric acid solutions. Chemical Engineering Journal 68, 123–130. El-Shall, H., Rashad, M.M., Abdel-Aal, E.A., 2005. Effect of cetyl pyridinium additive on crystallization of gypsum in phosphoric and sulfuric acids medium. Crystallisation Research and Technology 40 (9), 860–866. Firsching, F.H., Brune, S.N., 1991. Solubility products of the trivalent rare-earth phosphates. Journal of Chemical Engineering 36, 93–95. Free, M.L., Zhu, J.S., Moudgil, B.M., 1999. The effect of PEO and organic sulfonates in enhancing phosphogypsum filtration. International Journal of Mineral Processing 57, 25–42. Hannachi, A., Habaili, D., Chtara, C., Ratel, A., 2007. Purification of wet process phosphoric acid by solvent extraction with TBP and MIBK mixtures. Separation and Purification Technology 55 (2), 212–216. Harvinderpal, S., Mishra, S.L., Vijayalakshmi, R., 2004. Uranium recovery from phosphoric acid by solvent extraction using a synergistic mixture of di-nonyl phenyl phosphoric acid and tri-n-butyl phosphate. Hydrometallurgy 73, 63–70. Hurst, F.J., Crouse, D.J., 1973. Reductive stripping process for the recovery of uranium from wet-process phosphoric acid. U.S. Patent 3711591. Hurst, F.J., Crouse, D.J., Brown, K.B., 1969. Solvent Extraction of Uranium From WetProcess Phosphoric Acid. Oak Ridge National Laboratory. Report ORNLTM-2522. Hurst, F.J., Crouse, D.J., Brown, K.B., 1972. Recovery of uranium from wet-process phosphoric acid. Industrial Engineering and Chemical Process Design Developments 11 (1), 122–128. Ilyin, Andrei V., 1998. Phosphorites of the Russian Craton. Earth-Science Reviews 45 (1–2), 89–101. Jarosinski, A., Kowalczyk, J., Cz., Mazanek, 1993. Development of the Polish wasteless technology of apatite phosphogypsum utilization with recovery of rare earths. Journal of Alloys and Compounds 200, 147–150. Jorjani, E., Bagherieh, A.H., Mesroghli, Sh., Chehreh Chelgani, S., 2008. Prediction of yttrium, lanthanum, cerium, and neodymium leaching recovery from apatite concentrate using artificial neural networks. Journal of University of Science and Technology Beijing 15 (4), 367–374. Koopman, C., Witkamp, G.J., 2000. Extraction of lanthanides from the phosphoric acid production process to gain a purified gypsum and a valuable lanthanide by-product. Hydrometallurgy 58, 51–60. Krea, M., Khalaf, H., 2000. Liquid–liquid extraction of uranium and lanthanides from phosphoric acid using a synergistic DOPPA–TOPO mixture. Hydrometallurgy 58, 215–225. Lededev, V.N., Lokshin, E.H.P., 2000. Method of isolation of rare earth elements from phosphogypsum. RU Patent 2158317 (in Russian). Lededev, V.N., Masloboev, V.A., Lokshin, E.H.P., Popova, L.A., Sergeeva, S.D., 2000. Method of extraction of rare earth element phosphates from apatite concentrate. RU Patent 2120408 (in Russian). Liu, X., Byrne, R.H., 1997. Rare earth and yttrium phosphate solubilities in aqueous solution. Geochimica et Cosmochimica Acta 61, 1625–1633. Lokshin, E.H.P., Vershkova, J.U.A., Kalinnikov, V.T., Tareeva, O.A., Ivlev, K.G., Fedorov, S.G., et al., 2004. Method of recovering rare earth minerals from phosphogypsum. RU Patent 2225892. (in Russian). Pessac, J.F., Lambert, A.P., Tognet, J.P., 1987. Essentially complete recovery of uranium, yttrium, thorium and rare earth values from phosphate rock during wet-process production of phosphoric acid there from. U.S. Patent 4636369. Preston, J.S., Cole, P.M., Craig, W.M., Feather, A.M., 1996. The recovery of rare earth oxides from a phosphoric acid by-product. Hydrometallurgy 41, 1–19. Rickelton, W.A., 1988. Recovery of uranium from wet process phosphoric acid using assymetrical phosphine oxides. U.S. Patent 4778663. Samir, I., Abu-Eishah, Nizar, M., Abu-Jabal, 2001. Parametric study on the production of phosphoric acid by the dihydrate process. Chemical Engineering Journal 81, 231–250. Sangwal, K., 1993. Effect of impurities on the processes of crystal growth. Journal of Crystal Growth 128 (1–4), 1236–1244. Sangwal, K., 1996. Effects of impurities on crystal growth processes. Progress in Crystal Growth and Characterization of Materials 32 (1–3), 3–43. Sangwal, K., 2002. On the nature of supersaturation barriers observed during the growth of crystals from aqueous solutions containing impurities. Journal of Crystal Growth 242 (1–2), 215–228. Sato, T., 1989. Liquid–liquid extraction of rare-earth elements from aqueous acid solutions by acid organophosphorus compounds. Hydrometallurgy 22 (1–2), 121–140. Smith, B.R., Alexander, A.E., 1970. The effect of additives on the process of crystallization. Journal of Colloid and Interface Science 341, 80–90. Tadros, M.E., Mayes, I., 1979. Linear growth rates of calcium sulfate dihydrate crystals in the presence of additives. Journal of Colloid and Interface Science 722, 245–254. Xu, G.X., 1995. Rare Earth, Second Edition. Metallurgical Industry Press, Beijing. (in Chinese). Zhang, Y.Q., Li, J.N., Huang, X.W., Wang, C.M., Zhu, Z.W., Zhang, G.C., 2008. Synergistic extraction of rare earths by mixture of HDEHP and HEH/EHP in sulfuric acid medium. Journal of Rare Earths 5, 688–692.