Rubidium and Potassium Extraction from Granitic Rubidium Ore

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Rubidium and Potassium Extraction from Granitic Rubidium Ore: Process Optimization and Mechanism Study Peng Xing, Chengyan Wang, Baozhong Ma, Ling Wang, Wenjuan Zhang, and Yongqiang Chen ACS Sustainable Chem. Eng., Just Accepted Manuscript • DOI: 10.1021/ acssuschemeng.7b04445 • Publication Date (Web): 27 Feb 2018 Downloaded from http://pubs.acs.org on March 1, 2018

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ACS Sustainable Chemistry & Engineering

Rubidium and Potassium Extraction from Granitic Rubidium Ore: Process Optimization and Mechanism Study Peng Xing, Chengyan Wang,* Baozhong Ma, Ling Wang, Wenjuan Zhang, and Yongqiang Chen* School of Metallurgical and Ecological Engineering, University of Science and Technology Beijing, Beijing 100083, China. Mailing address: 30 Xueyuan Road, Haidian District, Beijing 100083, China Corresponding Authors: [email protected] (C. Wang) and [email protected] (Y. Chen).

ABSTRACT: Thus far, little is known about the use of granitic rubidium ore for the extraction of rubidium (Rb). Herein, we examined the extractability of rubidium and potassium (K) from granitic rubidium ore via sulfuric acid baking, reductive decomposition, and alkaline leaching. In addition, the extraction mechanism was studied by using interdisciplinary approaches based on the mineralogy and thermodynamics. Under the optimum conditions, more than 94% and 92% of Rb and K were extracted, respectively. Mineralogical analysis suggests that the Rb is scatted in micas (biotite and muscovite) and potassium feldspar in the form of isomorphism. Micas were transformed into sulfates during the sulfuric acid baking process. Increasing the dosage of coal,

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decomposition temperature, and the time was conducive to the decomposition of sulfates. The sulfates were transformed into aluminum oxide and potassium sulfate during the reductive decomposition process. The basic structure of potassium feldspar was unchanged in the stages of baking and decomposition. The feldspar was finally altered to cancrinite and zeolite in alkaline leaching. These phase transformations reveals the release approach of Rb and K from micas and feldspar.

KEYWORDS: Granitic rubidium ore, Rubidium, Potassium, Extraction mechanism INTRODUCTION Rb is a soft, silvery-white metallic element of the alkali metal group. It has been used for thermoelectric generator, atomic clocks, photocell, laser application, medicine, etc.1-5 Especially in recent years, the new and more efficient techniques established in photocell and laser application with the usage of Rb have promoted the demand for it. Rb is the 16th most abundant metal element in the Earth’s crust, rather more abundant than some common metals, such as Cu, Pb, and Zn and the main group elements Cs and Li, but is always obtained only as a byproduct of the extraction of these two metals, the main reason is that the Rb forms no minerals of its own.6 Lepidolite ((K,Rb)Li2AlSi4O10F2) and pollucite (Cs2Al2Si4O12) are considered to be the known principal source of Rb. In addition, Rb is ubiquitous in potassium minerals, such as carnallite, mica, and potassium feldspar.7 Geologically, the latter two minerals are found often together in granitic rocks.8 Minor amounts of Rb are reported in brines, kaolin clay waste, boron clay waste, gold waste, and jarosite.9-15 The early process for the extraction of Rb from minerals was the prolonged leaching with sulfuric acid. Roasting the ore with calcium salts followed by water leaching was developed


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afterwards to improve extraction efficiency.6,12,13,16-18 Calcium chloride has been found to be the most effective roasting agent, especially in presence of silica. The Rb in leaching solution can be extracted by fractional crystallization, solvent extraction, ion exchange, and adsorption.19-22 The existing roasting process, however, has several issues: large consumption of roasting agent (with a dosage of more than 50% of the mass of feed), generation of waste gas (HCl) and waste water with high salinity (mainly the calcium chloride), and severe corrosion of equipment. The application potential and the extraction challenge of Rb have caused the concern about the development of new methods for the extraction of that from resources.23 A type of Rb-bearing granitic rock is widely distributed in southern China (e.g. Jiangxi, Guangxi, and Guangdong provinces), which is the complex, multi mineral aggregate and generally consists of micas, quartz, and potassium feldspar. However, this granitic rock has not been regarded as a rubidium resource, because the data on the extractability of Rb are insufficient. Although flotation is a commonly method for the enrichment of mica,24,25 it is not applicable to the treatment of such ore, as the valuable components Rb and K are scatted both in micas and feldspar. Geochemical studies on the micas weathering show that the micas could be slowly leached under mild acidic conditions.26-28 More recently, hydrothermal acid leaching of mica ore was proposed by Luo et al.29 Although the high acidity leaching was effective to break the structure of mica, due to the low utilization efficiency of sulfuric acid, a considerable portion of acid was lost in the leaching solution and could not be regenerated. In addition, there is a paucity of information available on the mechanism of interactions between micas and sulfuric acid in the literature. Similarly, the sulfuric acid was used for the leaching of aforementioned granitic rubidium ore by our group before. However, the structure of feldspar was hard to destroy,

leading

to

the

low

extraction

efficiencies

of

Rb

and

K.

A

study


3


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within a geochemical context shows that the dissolution rate of feldspar increases monotonically with increasing aqueous fluoride concentration at acidic condition.30 Nevertheless, fluoride has disadvantages, such as high corrosiveness and complicated recycling. Moreover, Ciceri et al. studied the leaching of potassium feldspar ore with nitric acid solution.31 Xu and van Deventer reported the formation of geopolymeric gels from alkali-feldspars by alkaline dissolution.32 However, as the focus of the research was surface chemistry, mineral dissociation and potassium release were not involved in the literature. Using low temperature sulfuric acid baking to extract metals from secondary sources has attracted attention in recent years. Owing to the violent interactions between materials and concentrated sulfuric acid, the method has the merits of high extraction efficiency while low reagent and energy consumptions.33-36 By contrast, sulfuric acid baking is rarely reported in minerals processing. This study presents the extraction of Rb and K from granitic rubidium ore through sulfuric acid baking, reductive decomposition, and alkaline leaching. The sulfuric acid consumed in the baked ore could be regenerated via decomposing baked ore with the subsequent production of sulfuric acid. The Rb and K in alkaline leaching solution could be recovered by solvent extraction with 4-tert-butyl-2-(α-methylbenzyl)-phenol (t-BAMBP), which will be reported later. The process parameters, phase transformations, and the micro-structural changes of Rb and K bearing micas and feldspar during the sulfuric acid baking, decomposition and alkaline leaching are primarily studied in the present work to optimize the process and clarify the mechanism of Rb and K extraction from granitic rubidium ore. EXPERIMENTAL SECTION Materials. The chemical composition of the granitic rubidium ore is shown in Table 1. The sulfuric acid and sodium hydroxide used in this study were both of analytically pure grade. Low


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cost lignite (52.1% of fixed carbon content) was used as carbonaceous reducing agent for the decomposition of baked ore. Table 1. Chemical Composition of the Rubidium Ore

Component Content (wt %)

Al

K

Fe

Na

Ca

Mg

C

Rb

S

SiO2

8.77

5.67

3.17

1.21

0.70

0.29

0.15

0.09

0.06

62.2

Apparatuses and Procedure. The raw ore was preliminarily crushed into 0.5–5 mm and then ground below -0.074 mm in a rod mill (XZM-100, Wuhan Exploring Machinery Factory, China). The ground sample (20 g) was mixed with sulfuric acid in a porcelain boat and subsequently baked in a tube electric furnace (BLMT-1600G, Luoyangshi Bolaimante Test Electric Co. Ltd., China) at predefined temperature and time. The baked sample was ground in the rod mill for 2 min, mixed with lignite and then placed in the tube electric furnace for decomposition. The decomposition product was finally leached with NaOH solution in a pressure autoclave (GSHA, Weihai Xintai Chemical Machinery Co. Ltd., China). In order to determine the optimum conditions for baking and decomposition, the parameters (NaOH concentration, liquid/solid ratio (mL/g), temperature, and time of leaching) were held constant at 300 g/L, 20:1, 150 °C and 1 h, respectively. Analysis

Methods.

Thermogravimetric-differential

scanning

calorimetry

(TG-DSC)

experiment was performed on a NETZSCH STA 409C unit under a nitrogen atmosphere, at a heating rate of 10 K min-1. X-ray diffraction (XRD) experiments were performed with a Rigaku Ultima IV diffractometer (Cu, Kα). The microstructures and modes of occurrence of the main elements of the raw ore, baked ore, and the decomposition product were analyzed by scanning electron microscopy (SEM; FEI Quanta 600) equipped with energy dispersive X-ray spectroscopy (EDS; EDAX GENESIS). Metals in raw ore and leaching residue were analyzed by


5


inductively coupled plasma optical emission spectrometer (ICP-OES) (Optima 7000DV, PerkinElmer), carbon and sulfur in raw ore, baked ore, and decomposition product were analyzed using a carbon-sulfur analyzer (EMIA-820V, Horiba, Japan), and the silica in raw ore was analyzed by silicon molybdenum blue spectrophotometry. RESULTS AND DISCUSSION Characterization of the ore. The XRD pattern of the raw ore (Figure 1) indicates that quartz (SiO2), orthoclase (KAlSi3O8), and micas were the major phases, with microline (KAlSi3O8), albite (NaSi3AlO8), chlorite ((Mg,Al)6(Si,Al)4O10(OH)8), and kaolinite (Al2Si2O5(OH)4) as minor phases. The further SEM-EDS analysis (Figures S1) shows that the micas were composed of biotite (K(Mg,Fe)3(AlSi3O10)(OH,F)2) and muscovite (KAl2(AlSi3O10)(OH)2) and the Rb was scatted both in micas and feldspar in the form of isomorphism. Chlorite was derived from biotite alteration while the kaolinite was derived from feldspar alteration. 10400

Intensity (Counts)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

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7800

5200

2600

0

46-1045>Quartz (SiO2) 03-0058>Kaolinite (Al2Si2O5(OH)4) 31-0966>Orthoclase (KAlSi3O8) 07-0025>Muscovite (KAl2Si3AlO10(OH)2) 19-0932>Microline (KAlSi3O8) 52-1044>Chlorite ((Mg,Al)6(Si,Al)4O10(OH)8) 41-1480>Albite (Na(Si3Al)O8) 10

20

30

40

50

60

70

80

90

2θ (°)

Figure 1. XRD pattern of the granitic rubidium ore. The TG-DSC profile of the ore (Figure S2) shows mass losses and three endothermic peaks: the distinct peak at 78.6 °C likely corresponds to the loss of free water, the indistinct peak at


6

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261.1 °C likely corresponds to the loss of a small amount of crystalline water, whereas the third endothermic peak at 496.9 °C likely corresponds to the dehydration of chlorite and kaolinite.37,38 Hydrothermal Direct Leaching of Ore. The direct sulfuric acid leaching of ore was previously performed in the pressure autoclave under the conditions of liquid/solid ratio 20:1, H2SO4 concentration 200 g/L, leaching temperature 200 °C, and leaching time 1.5 h. Nevertheless, the leaching ratios for Rb and K were only 66.3% and 39.7%, respectively. In addition, the direct alkaline leaching was simultaneously performed under the conditions of liquid/solid ratio 20:1, NaOH concentration 300 g/L, leaching temperature 150 °C, and leaching time 1 h. The leaching ratios for Rb and K were merely 40.2% and 60.5%, respectively. The leaching results indicate that the direct acid/alkaline leaching is not very effective for the complete extraction of Rb and K. Figure 2 indicates that micas, chlorite, and kaolinite were leached while the quartz, potassium feldspar, and albite remained in the residue during the direct acid leaching process. On the contrary,

feldspar

reacted

with

alkali

and

was

altered

to

cancrinite

(Na6Ca2Al6Si6O24(CO3)2·2H2O) and zeolite (Na2Al2Si3O10·2H2O), nevertheless, most of the micas did not react with alkali during the alkaline leaching. These results explain the unsatisfactory hydrothermal acid/alkaline leaching process.


7


O

Q-quartz (SiO2) O-orthoclase (KAlSi3O8) M-microline (KAlSi3O8) A-albite (Na(Si3Al)O8) ∗-muscovite (KAl2Si3AlO10(OH)2)

S2-potassium alum (KAl(SO4)2⋅12H2O)

Intensity

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

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Z-zeolite (Na2Al2Si3O10⋅2H2O)

M A Ca-cancrinite (Na Ca Al Si O (CO ) ⋅2H O) 6 2 6 6 24 3 2 2 A O direct acid leaching residue O O S2 O Q OO O Q Q Q Q O Q S2 Q Q Z Ca QZ Ca Z Z Z direct alkaline leaching residue Z Ca Ca Ca Ca O Ca Z Ca Ca * Q

O

Z * C

10

20

30

40

50

60

70

2θ (°)

Figure 2. XRD patterns of the direct acid leaching residue and alkaline leaching residue. Sulfuric Acid Baking of Ore. Three variables were studied for the sulfuric acid baking of ore: baking temperature, dosage of sulfuric acid, and baking time. Other constant parameters of decomposition included a mass ratio of coal/baked ore of 5%, a decomposition temperature of 750 °C, and a decomposition time of 8 min. The baking temperature positively influenced the extraction of Rb and K (Figure 3a). But the higher temperature (above 300 °C) would aggravate the volatilization and decomposition of sulfuric acid39 and thus is not applicable. It was noted that when mixing a kilogram of rubidium ore sample with sulfuric acid in a mass ratio of sulfuric acid/ore of 55%, the highest temperature of mixture was up to 142 °C. This indicates that when processing at large-scale, the additional heat would be greatly reduced due to the exothermic reaction. The amount of sulfuric acid was also an important factor influencing the extraction of Rb and K. The mass ratio of sulfuric acid/ore was varied in the range from 10 to 55%. As shown in Figure 3b, the extraction of K increased slowly while the extraction of Rb increased apparently with increasing dosage of sulfuric acid.


8

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Prolonging the baking duration was favorable to the extraction, as the extraction efficiencies of K and Rb increased with increasing baking time in the range from 5 to 20 min (Figure 3c). The further prolonging of baking time did not significantly improve metals extraction. The interaction between sulfuric acid and rubidium ore was fast: more than 80% of K and 75% of Rb were extracted only in 5 min. On the basis of the above batch tests results, the optimum conditions for the sulfuric acid baking were determined to be a temperature of 300 °C, a mass ratio of sulfuric acid/rubidium ore of 55%, and a baking time of 20 min. The phase transformations occurring during sulfuric acid baking are shown in Figure 4. The diffraction peaks of micas, chlorite, and kaolinite completely disappeared, meanwhile, the new phases (S1–S4) peaks appeared after sulfuric acid baking. Moreover, the diffraction peak intensities of quartz, orthoclase, microline, and albite somehow decreased. Figure S3 shows the microstructure of baked ore and the mode of occurrence of main elements. The microcrystalline particles of the newly generated sulfates and SiO2 gather together and scatter around the quartz and feldspar. As the new phases were extremely fine, it was difficult to obtain the spectrum of a single particle. The EDS of point 4 in Figure S3 indicates that followed by the K, the Rb in micas migrated to the potassium alum phase after the interaction of micas with sulfuric acid. Nevertheless, the structure of potassium feldspar was not destroyed during the sulfuric acid baking.


9


100

100

b

a 90

Metal extraction (%)


90

80

K Rb 20

Mass ratio of sulfuric acid/ore: 55%; Baking time: 20 min

10

0

80

K Rb

20

Baking temperature: 300 °C; Baking time: 20 min

10

0 150

200

250

300

10

20

30

40

50

60

Mass ratio of sulfuric acid/rubidium ore (%)

Backing temperature (°C) 100

c 90


1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

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80

K Rb 20

Baking temperature: 300 °C; Mass ratio of sulfuric acid/ore: 55%

10

0 5

10

15

20

25

30

Baking time (min)

Figure 3. Effects of (a) baking temperature, (b) dosage of sulfuric acid, and (c) baking time on the extraction of Rb and K.


10

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M O Q

7000

Q-quartz (SiO2) O-orthoclase (KAlSi3O8)

6000

M-microline (KAlSi3O8)

Intensity (Counts)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60


5000

A-albite (Na(Si3Al)O8) P-K2SO4

4000

S1-(Fe,Al)2(SO4)3 S3-Al2(OH)4(SO4)⋅7H2O

S2 M O S1AP AS3M O A S2

2000 1000

S2-potassium alum (KAl(SO4)2⋅12H2O)

S1 A O Q

3000

S4-Al4(SO4)(OH)10

S1 S4 S3 O

S3

OP

10

20

30

Q O Q Q P

Q Q

O

O

O

0 40

50

60

70

2θ (°)

Figure 4. XRD pattern of baked ore. The results of mineralogy in Figures 4 and S3 indicate that the Rb-bearing biotite (K(Mg,Fe)3(AlSi3O10)(OH,F)2) and muscovite (KAl2(AlSi3O10)(OH)2) react with sulfuric acid to form iron aluminum sulfate ((Fe,Al)2(SO4)3), potassium alum (KAl(SO4)2·12H2O), hydroxylaluminum sulfate (Al2(OH)4(SO4)·7H2O and Al4(SO4)(OH)10), SiO2 and potassium sulfate (K2SO4) in sulfuric acid baking: K(Mg,Fe)3(AlSi3O10)(OH,F)2+aH2SO4=b(Fe,Al)2(SO4)3+(1-2f)KAl(SO4)2·12H2O+d Al2(OH)4(SO4)·7H2O+eAl4(SO4)(OH)10+fK2SO4+gMgSO4+3SiO2+iH2O+jH2F

(1)

KAl2(AlSi3O10)(OH)2+0.5(5+5t-5x+v)H2SO4=tAl2(SO4)3+(1-2x)KAl(SO4)2·12H2O+v Al2(OH)4(SO4)·7H2O+0.5(1+x-t-v) Al4(SO4)(OH)10+xK2SO4+3SiO2+zH2O

(2)

Reductive Decomposition of Baked Ore. Industrial-scale decomposition of calcium sulfate and ferrous sulfate to produce acid has been reported in some countries, such as the USA and Germany.40 Previous studies have shown that the decomposition process is influenced by several factors (i.e. temperature, atmosphere, and retention time). The main sulfur-bearing phases in baked ore are iron aluminum sulfate, potassium alum and hydroxyl-aluminum sulfate (Figures 4 and S3), indicating that the sulfur recovery might be achieved by means of decomposition. The


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SO2 generated by decomposition could be used for sulfuric acid production, further, to achieve the recycling of sulfuric acid. Because of the shrinkage and hardening of ore sample in the process of sulfuric acid baking, the baked ore sample was ground for 2 min in the rod mill for decomposition. Particle diameter of ground baked sample was measured using a laser particle size distribution analyzer (LMS-30, Seishin Enterprise Co. Ltd., Japan), as shown in Figure S4. The following reactions are predicted to take place during the decomposition of iron aluminum sulfate: Al2(SO4)3 = Al2O3+3SO3(g) ∆GƟ=412.31-0.535TkJ/mol (3) Fe2(SO4)3 = Fe2O3+3SO3(g) ∆GƟ=422.12-0.538TkJ/mol (4) The existing results of TG and DTG of alunite ore41-43 suggest that the potassium alum decomposes mainly in two steps: dehydration (500~600 °C) and desulfation (~800 °C), as the following reactions: KAl(SO4)2·12H2O = KAl(SO4)2+12H2O (5) 2KAl(SO4)2 = K2SO4+Al2O3+3SO3(g) ∆GƟ=474.69-0.553TkJ/mol (6) Kodama and Singh44 proposed that the main reaction paths for decomposition of hydroxylaluminum sulfate may be as follows: firstly the loss of crystalline water, secondly the dehydroxylation of interlayer hydroxide, and finally the decomposition of sulfate at temperatures above 750 °C, as shown in eqs 7, 8, and 3. 3Al2(OH)4(SO4)·7H2O = Al2(SO4)3+2Al2O3+27H2O (7) 3Al4(SO4)(OH)10 = Al2(SO4)3+5Al2O3+15H2O (8) Specially, during the reductive decomposition (with addition of lignite) the carbon contributes to the decomposition of sulfates, as the following reactions:


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C+CO2(g)=2CO(g) ∆GƟ=124.13-0.177TkJ/mol (9) Al2(SO4)3+3CO(g) = Al2O3+3SO2(g)+3CO2(g) ∆GƟ=-146.47-0.553TkJ/mol (10) Fe2(SO4)3+3CO(g) = Fe2O3+3SO2(g)+3CO2(g) ∆GƟ=-136.66-0.556TkJ/mol (11) 2KAl(SO4)2+3CO(g)=K2SO4+Al2O3+3SO2(g)+3CO2(g) ∆GƟ=-84.09-0.571TkJ/mol (12) The diagram of Gibbs free energy change (∆GƟ) of above reactions versus temperature is drawn, as shown in Figure 5. Decomposition temperatures of aluminum sulfate and potassium alum are calculated to be 770 and 858 °C, respectively, quite close to the temperatures reported in Ref.45,46 Adding coal in the decomposition process reduces the ∆GƟ of the decomposition reaction of sulfates, and thus would be beneficial to the decomposition. In addition, the ∆GƟ decreases with the increase of temperature, indicating that the elevated temperature increases the possibility of decomposition. The ∆GƟ of the eqs 10–12 are negative even at room temperature, nonetheless, according to the Boudol reaction of solid carbon reduction (eq 9), the temperature required to produce carbon monoxide is calculated to be 700 °C, which means that the operating temperature for reductive decomposition is required to be above 700 °C. 400 200 0 -200

Reductive decomposition zone of sulfates

θ

∆ G (kJ)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60


Reaction (3) Reaction (4) Reaction (6) Reaction (10) Reaction (11) Reaction (12) Reaction (9)

-400 -600 -800 0

200

400

600

800

1000

Temperature (°C)

Figure 5. Diagram of Gibbs free energy change of decomposition reactions versus temperature.


13


In order to study the effect of coal on the extraction of Rb and K and the recovery of sulfur, a series of decomposition tests about dosage of coal were conducted. Coal was found to have a little effect on the extraction of Rb and K but a significantly positive effect on the recovery of sulfur. As the above-mentioned sulfates could be dissolved in strong basic solution47, the extraction of Rb and K were basically not affected. Recovery of sulfur without addition of coal was only 20.73%, because the operating temperature was slightly lower than the theoretical decomposition temperature of sulfates. The results in Figure 6a are consistent to the conclusions in Figure 5. When the mass ratio of coal/baked ore was increased above 5%, the sulfur recovery reached a plateau. Thus, the optimum mass ratio of coal/baked ore was determined to be 5%.

a

100

80

80

60 60

K extraction Rb extraction Sulfur recovery

40

40

Decomposition temperature: 750 °C; Decomposition time: 8 min

20

M etal extraction (% )

100

b

100

80 80 60

K extraction Rb extraction sulfur recovery

40

60

40

Mass ratio of coal/baked ore: 5%; Decomposition time: 20 min

20

R ecovery of su lfur (% )


100


20

0.0

2.5

5.0

650

7.5

c

100

700

750

800

Decomposition temperature (°C)

Mass ratio of coal/baked ore (%)


20

0

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80 80 60

K extraction Rb extraction Sulfur recovery

40

60

40

Decomposition temperature: 750 °C; Mass ratio of coal/baked ore: 5%

20


1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

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Decomposition time (min)


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Figure 6. Effects of (a) dosage of coal, (b) decomposition temperature, and (c) decomposition time on the extraction of Rb and K and sulfur recovery. According to the Figure 5, temperature is also a key factor to decomposition. Hence, the effect of temperature was investigated in the range from 650 to 800 °C. As shown in Figure 6b, sulfur recovery at 700 °C was a leap compared with that at 650 °C, which corroborates the conclusions of Figure 5. Nevertheless, the recovery efficiency of sulfur increased slowly above 700 °C, which was mainly attributed to the stability of potassium sulfate. The maximum extraction of Rb and K was obtained at 750 °C, whereas a higher temperature (800 °C) led to the decrease of the metals extraction, an explanation is that the slight sintering might occur at that temperature, which was negative for leaching. Increasing decomposition temperature was apparently beneficial to the recovery of sulfur over the range, but in order to ensure the extraction of Rb and K, a decomposition temperature of 750 °C was appropriate. The variations in metals extraction and sulfur recovery with decomposition time (2–20 min) were investigated. It was observed that more than 80% of sulfur could be recovered in only 8 min and the maximum extraction of Rb and K was obtained in 10 min (Figure 6c). The further prolonging of decomposition time did not significantly improve sulfur recovery but resulted in the decrease of metals extraction, which could be attributed to the possible occurrence of slight sintering. Thus, 10 min was chosen as the preferable decomposition time. On the basis of the above batch tests results, the optimum conditions for the reductive decomposition were determined to be a temperature of 750 °C, a mass ratio of coal/baked ore of 5%, and a decomposition time of 20 min. An 84.8% sulfur recovery was obtained under such conditions.


15


Despite the addition of coal, significant phase transformation was not observed at low decomposition temperature (650 °C) (Figure 7a). The intensity of diffraction peaks of (S1) significantly increased, while the intensity of (S3) and (S4) diffraction peaks decreased during the decomposition without addition of coal (Figure 7b). This trend indicates that in the absence of carbonaceous reducing agent the sulfates tend to dehydrate firstly, but the desulfation of which is ineffective. In the presence of carbonaceous reducing agent (mass ratio of coal/baked ore 5%), as temperature was increased to 750 °C, the above sulfates diffraction peaks almost completely disappeared while the intensity of diffraction peaks of potassium sulfate (P) increased (Figure 7c). This finding is in good accordance with the previous results in Figures 5 and 6 that the operating temperature for decomposition must be above 700 °C and the carbonaceous reducing agent could obviously promote the decomposition of sulfates in baked ore. However, the feldspar peaks were unchanged during decomposition process. Q,O

Q-quartz (SiO2) O-orthoclase (KAlSi3O8) M-microline (KAlSi3O8) A-albite (Na(Si3Al)O8) S1-(Fe,Al)2(SO4)3 S3-Al2(OH)4(SO4)⋅7H2O

P A Q O

Intensity

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

Page 16 of 27

S4-Al4(SO4)(OH)10 S5-KAl(SO4)2 P-K2SO4 A O O PO S1

A,M O P P OO

(c) O Q

Q Q PQ

Q

Q,O

Q S1 S1 A O S5 A,M AS3 O S1 S1,S5 Q M O S5

S1 QQ Q

S1 Q

O Q

S1 QQ Q

S1 Q

OO,S1

Q

Q

Q,O,M

S5 S4 S3

S1

(b) M

S1 Q

S1 Q

Q,O,M

S5 S4 S1 S3 S3

10

Q O A S1 O S1 A,M A O S3S5 S1S5S1 S5

20

30

(a) Q

40

Q,O M

50

Q,S1 Q,S1

60

2θ ( ° )

Figure 7. XRD patterns of decomposition products ((a) decomposing at 650 °C with addition of coal; (b) decomposing at 750 °C without addition of coal; (c) decomposing at 750 °C with addition of coal).


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The structure of potassium feldspar was not destroyed during the reductive decomposition (Figure S5). Although the peak of aluminum oxide was absent in the XRD pattern of decomposition product (Figure 7c), the EDS of point 4 in Figure S5 shows that one of the major decomposition product, Al2O3 really existed in amorphous state. Alkaline Leaching. The decomposition product obtained under the optimal conditions was leached by sodium hydroxide to study the effects of temperature, liquid/solid ratio, NaOH concentration, and time on the leaching of Rb and K. The leaching results are shown in Table S1. It was found that these parameters had positive effect on the leaching of Rb and K. When considering the production practices and economic benefits, optimum conditions for leaching were determined to be a temperature of 150 °C, a liquid/solid ratio of 15:1, a NaOH concentration of 250 g/L, and a leaching time of 1 h with the extraction efficiencies of 94.7% and 92.2% for Rb and K, respectively. After the desilication and solvent extraction of Rb and K, the alkaline solution could be sent to the leaching step again to close the loop, which has been thoroughly studied but is not included in this article. In order to investigate the mechanism of alkaline leaching, the XRD pattern of alkaline leaching residue was analyzed. The main minerals, which are evident from the XRD pattern, are cancrinite and zeolite (Figure 8). They were derived from the hydrothermal alteration of feldspar, aluminum oxide, and silica in decomposition product. The diffraction peaks of potassium sulfate (K2SO4) and residual sulfates ((Fe,Al)2(SO4)3) completely disappeared and the quartz and orthoclase peaks dramatically decreased. These phase transformations account for the leaching of Rb and K from potassium sulfate and feldspar. Considering the uses of zeolite and cancrinite48,49, the alkaline leaching residue is suitable for recycling as adsorbent or building stone.


17


6000 Z,Ca

5000

Ca-cancrinite (Na6Ca2Al6Si6O24(CO3)2⋅2H2O) Z,Ca

Intensity (Counts)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

Page 18 of 27

4000

Z-zeolite (Na2Al2Si3O10⋅2H2O) Q-quartz (SiO2)

Z,Ca

O-orthoclase (KAlSi3O8)

3000

Q Z Z,Ca

2000

Ca Ca Ca Z

1000

Q Z

Z

O

Ca Ca Ca

Ca CaCa Ca Ca Ca Z Ca

ZZ

0 10

20

30

40

50

60

70

80

2θ (°)

Figure 8. XRD pattern of the alkaline leaching residue. Figure S6 shows that the short columnar cancrinite crystals are ubiquitous while the fibrous zeolites tend to aggregate. The main chemical reactions that occurred in the new phase formation are as eqs. 13 and 14. A small amount of plagioclase in ore provided a source of calcium for cancrinite formation. Due to the interaction between potassium feldspar and alkali, K and associated Rb in feldspar are finally released. 6KAlSi3O8

+

2CaCO3

+24NaOH

=

Na6Ca2(Al6Si6O24)(CO3)2·2H2O↓

+3K2SiO3+9Na2SiO3+10H2O (13) 6KAlSi3O8 + 18NaOH = 3Na2(Al2Si3O10)·2H2O↓ +3K2SiO3+6Na2SiO3+3H2O (14) According to the above analyses, the mechanism of Rb and K extraction from granitic rubidium ore is summarized. Mineralogical analysis suggests that the Rb is scatted in micas (biotite and muscovite) and potassium feldspar in the form of isomorphism. During the acid baking, the structure of the micas was completely destroyed while the structure of feldspar was intact. The micas were transformed into iron aluminum sulfate, potassium alum, hydroxylaluminum sulfate, potassium sulfate, and SiO2. The baked ore were treated by reductive decomposition to release the SO2, so as to achieve the regeneration of sulfuric acid. In the third


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stage, namely alkaline leaching, feldspar in decomposition product was altered to cancrinite and zeolite, thus leading to the maximum extraction of Rb and K. CONCLUSIONS The extraction of Rb and K increased with increasing baking temperature, dosage of sulfuric acid, baking time, and decomposition temperature and the recovery of sulfur from baked ore increased with the increase of the dosage of coal, decomposition temperature, and time. In addition, leaching temperature, liquid/solid ratio, NaOH concentration, and leaching time positively influenced the leaching of Rb and K. Under the optimum conditions, more than 94% and 92% of Rb and K were extracted from granitic rubidium ore, respectively. These results suggest that Rb-rich granitic ore is a highly promising resource for Rb and K. The mechanism of Rb and K extraction is clarified. Sulfuric acid could promote the phase transformation of inert micas into active sulfates (iron aluminum sulfate, potassium alum, potassium sulfate and hydroxyl-aluminum sulfate) and SiO2. These sulfates decomposed in two steps: dehydration followed by desulfation and finally transformed into aluminum oxide and potassium sulfate during the reductive decomposition process and the sulfur dioxide was simultaneously released, in this way the regeneration of sulfuric acid could be achieved. The SEM-EDS analyses of baked ore indicate that followed by the K, the Rb in micas migrated to the potassium alum phase after the interaction of micas with sulfuric acid. The SEM analyses of baked ore and decomposition product show that the basic structure of potassium feldspar was unchanged in the above two stages. The potassium feldspar was altered to cancrinite and zeolite in the following hydrothermal alkaline leaching. As a result, Rb and K were released from potassium sulfate and feldspar. ASSOCIATED CONTENT


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Supporting Information Additional results as shown in Figures S1−S6, and Tables S1 (PDF) AUTHOR INFORMATION Corresponding Authors *E-mail: [email protected]. Tel.: +86-10-62332271. *E-mail: [email protected]. Notes The authors declare no competing financial interest. ACKNOWLEDGMENTS This research was funded by the Beijing Natural Science Foundation (2182040), National Natural Science Foundation of China (51674026 and U1302274), Fundamental Research Funds for the Central Universities (230201606500078), and the Beijing Science & Technology Program (Z171100002217063). REFERENCES (1) Saliba, M.; Matsui, T.; Domanski, K.; Seo, J.Y.; Ummadisingu, A.; Zakeeruddin, S.M.; Correa-Baena, J.P.; Tress, W.R.; Abate, A.; Hagfeldt, A.; Grätzel, M. Incorporation of rubidium cations into perovskite solar cells improves photovoltaic performance. Science 2016, 6309, 206– 209.


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Table of Contents graphic

Synopsis A feasible strategy to extract rubidium and potassium from granitic rubidium ore was developed, and the extraction mechanism was clarified.


27

Rubidium and Potassium Extraction from Granitic Rubidium Ore

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