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An assessment for the recovery of lanthanides and [P.sub.2][O.sub.5] from phosphate rocks.

INTRODUCTION

Phosphate rock is a raw material in the fertilizers used to supply food and feed for mankind and animals. Most of the world's phosphate rocks are of sedimentary origin and primarily composed of the apatite group in association with a wide assortment of accessory minerals, fluorides, carbonates, clays, quartz, silicates, and metal oxides.

Phosphoric acid is an important intermediate chemical product. It is mainly used for the manufacturing of fertilizers. Production capacity for phosphoric acid yielded about 33 million tons of [P.sub.2][O.sub.5]. About 90% of world [P.sub.2][O.sub.5] consumption involves the fertilizer industry. There is a steadily growing demand for phosphate fertilizers. The phosphoric acid production is directly linked to the phosphate fertilizer consumption. In spite of phosphate rocks are considered as a poor source of uranium and REE; it contains 100-1000 g REE per each ton of ore, depending on phosphate type and origin [1]. Attention has been directed to phosphate deposits as a potential source of these elements as separation of rare earth elements (REEs) namely, scandium, yttrium and lanthanides from phosphate rocks prior to production of phosphoric acid has a great importance from economic point of view due to the wide range of applications in high-tech applications as electronic industry, steel industry, household batteries, fluorescent lamps, permanent magnets, and lasers for surgical and nuclear technologies [25] and environmental aspects that is greatly recommended for production of phosphoric acid with lowering impurities [6,7].

Al Jalamid site comprises which located in the north of KSA act as a pioneer source of phosphate rocks for its huge content from phosphate and percentage of [P.sub.2][O.sub.5] ([approximately equal to]28%). The phosphate mine, beneficiation plant and supporting infrastructure and encompasses an area of approx. 50 sq. km. Mine production is around 11.6 Mtpy of ore and the beneficiation facilities can produce an estimated 5 Mtpy of flotation concentrate on a dry basis.

The wet process presents 90% of the world current phosphoric acid production which produces high volume, low purity phosphoric acid from phosphate ores and of low cost. In this process, the mineral acid that is used for the acidulation may be nitric, hydrochloric or sulfuric. Sulfuric acid is advantageous since it is the only acid which forms insoluble precipitate (CaS[O.sub.4] x 2[H.sub.2]O), which can be separated from the produced acid but the REE co precipitated with the phosphogypsum [8,9]

[Ca.sub.10][(P[O.sub.4]).sub.6][F.sub.2]+ 10[H.sub.2]S[O.sub.4] +10n [H.sub.2]O [left right arrow] 6[H.sub.3]P[O.sub.4] + 10CaS[O.sub.4] x n [H.sub.2]O + 2HF (1)

Hydrochloric acid is aggressive acid, it has very fast kinetic, the drawback arises in the difficulty the corrosion of equipments that may be arises.

[Ca.sub.10][(P[O.sub.4]).sub.6][F.sub.2] + 20HC1 [left right arrow] 6[H.sub.3]P[O.sub.4] + 10Ca[Cl.sub.2] + 2HF (2)

When phosphate rock is treated with nitric acid, phosphoric acid and soluble calcium nitrate are formed according to the following equation:

[Ca.sub.10][(P[O.sub.4]).sub.6][F.sub.2] + 20HN[O.sub.3] [left right arrow] 6[H.sub.3]P[O.sub.4]+ 10Ca[(N[O.sub.3]).sub.2] + 2HF (3)

in nitric acid way or nitrophosphate is used to obtain ternary NPK fertilizer or binary NP fertilizer [10-13]. The objective of the present paper is to study the recovery of REE, U and [P.sub.2][O.sub.5] from phosphate rock and the production of NPK fertilizer. Different experimental conditions were optimized.

Experimental:

Reagents:

The apatite used in this work was obtained from Hazm Al-jalamid area, northern Saudi Arabia. All chemicals and reagents used were analytical reagent grade and their solutions were prepared with distilled water. All rare earth elements were obtained from Sigma Aldrich (99.9%). The extractant HDEHP (di-2-ethylhexyl phosphoric acid) and TOPO (Tri-n- phosphine oxide) were purchased from Sigma Aldrich.

Instruments:

X-ray Fluorescence (XRF, Philips PW 2404) was employed to determine the overall chemical composition of apatite; the samples were analyzed using pressed powder pellets, with the calibration refined using both Standard Reference Materials (SRMs) and synthetic ones but failed to indicate any REE (due to their extremely low quantity). The individual REE and uranium were determined by inductively coupled plasma (ICP) and pH measurements were performed by Orion 920 pH meter. The nitrogen content in the leach solution was determined by ion chromatography while potassium was determined by ICP. [P.sub.2][O.sub.5] was spectrometrically by means of the vanadomolybdo phosphoric acid colorimetric method [14] using a Shimadzu, UV/visible recording spectrophotometer of type UV-160A. The concentration of the metal in the organic phase was calculated from the difference between its concentration in the aqueous phase before and after extraction. Hence, the distribution ratio (D) was calculated as the ratio of the concentration of the metal in the organic phase to its concentration in the aqueous phase.

Procedure:

The working sample was collected from the experimental mine of Al Jalamid site which located in The north of KSA. The chemical analysis of the phosphate rock is as shown in Table 1.Four different size fractions (0.063-0.15, 0.25-0.315, 0.5-0.6, and 0.7-0.75 mm) were obtained by sieving through ASTM standard sieves, 0.5-0.6 mm was used in this investigation. The samples of the ground phosphate rock were dried in an electric oven at about 105 [degrees]C, cooled to room temperature and stored in closed desiccators. The chemical composition of JALAMIDE phosphate concentrate is given in Table 1. The content of uranium and REE in various phosphate rocks was given in Table 2 [15-18]. Al JALAMIDE phosphate concentrate sample under investigation contains 600 ppm of REE that is near to that obtained by Jordanian phosphate rocks as literature by Al-Thyabat and Zhang [19]

RESULTS AND DISCUSSION

Leaching process:

Preliminary experiments were conducted to investigate the optimum conditions for the recovery of Ln, U and [P.sub.2][O.sub.5] using 0.5, 1 and 2 M nitric acid as leaching agent at phase ratio L/S = 4 and stirring time of 5, 10, 20 and 30 min and 25 [degrees]C. Table 3 showed that 0.5 M HN[O.sub.3], phase ratio L/S = 4 and 5 min at 25 [degrees]C gave excellent recovery for lanthanides with little contamination of U and [P.sub.2][O.sub.5], hence the following experiments were carried out the above mentioned conditions.

Leaching of lanthanides:

Based on the above investigations, nitric acid pre-leach were optimized to maximize rare earth dissolution and minimize calcium and phosphate dissolution, the lanthanides were selectively leached from phosphate rock with a little uranium and [P.sub.2][O.sub.5] by conducting 12.5 g of the phosphate rock (0.5-0.6 mm) with 50 mL of 0.5 M nitric acid in thermo- static vessel and 400 rpm,, in two extraction stages each within 5 min. the distribution of lanthanides were given in Table 4. 96% of lanthanides were leached within two extraction stages, the weight loss in the sample were 16 and 27%, respectively. The leached solution was adjusted to pH 1 then agitated with 9% oxalic acid in three successive portions each within for 10 min. after the precipitation of lanthanides as [(REE).sub.2][([C.sub.2][O.sub.4]).sub.3], n[H.sub.2]O(s), the precipitate was filtered, washed with water, dried in air, the composition ware analyzed by XRF, the distribution of REE. In oxalate leach solution, the concentration of [P.sub.2][O.sub.5] was 3.5 g/L. Finally this oxalate was calcinied at 950 [degrees]C to obtain [Ln.sub.2][O.sub.3].

Leaching of uranium:

After leaching of lanthanides, The residue (10.5 g) was agitated with 3 M nitric acid at L/S =4 for 30 min.; the leached nitrate solution that is rich in uranium, [P.sub.2][O.sub.5] among with other impurities mainly calcium, the weight loss in the sample reached 90%,. Uranium was extracted by synergistic mixture of 0.1 mol/L D2EHPA/0.1 mol/L TOPO in kerosene with a ratio 4:1 at (O/A) = 1 that was conducted with the nitrate solution for 30 min. at 25 [degrees]C. The concentration of uranium in the organic phase was calculated from the difference between its concentration in the aqueous phase before and after extraction. The extraction percent was found to be 20% which is some extent agreed with the preliminary investigations, the stripping was carried out by 2 M [Na.sub.2]C[O.sub.3] solution.

Preparation of NPKfertilizer:

The raffinate that is rich in [P.sub.2][O.sub.5] among with that results from lanthanides route were treated with KOH until the pH 4 to obtain NPK fertilizer solution with high potassium content. The nitrogen and phosphorus content were determined by Ion Chromatography while potassium was determined by ICP. The composition of NPK sample is N[H.sub.4.sup.+], N[O.sub.3.sup.-],P[O.sub.4.sup.-3] and K are 1.314, 70, 2.32 and 21.5 g/L, respectively. No fluoride was detected due to the role of adding potassium salt in addition to the present silica that formed potassium fluorosilicate [20]. The concentration of uranium was 0.02 ppm. The white precipitate that resulted during the neutralization of acidic nitrphosphate solution was analyzed is found to be mainly calcium hydroxide in addition to potassium fluorosilicate that was free of uranium and REE. The composition of that precipitate was analyzed by XRF and found to be (Ca, 83%, P, 0.67%, Fe, 0.45%, Cr 0.09%, Si 0.66% and Al, 0.38%). The residue (1.25 g; 10% of the original mass) that was not leached by nitric acid was analyzed by XRF and found to be mainly calcium fluoride that is contaminated with silica (Ca, 86.5%, P, 3.78%, Fe, 2%, Cr, 0.2%, Si, 2.%45, Al, 0.2%, F, 2.6%, Y, 0.18%, Zr,0.2%). Figure 2 represented conceptual treatment of apatite ore using nitric acid.

* Proposed process:

** On basis of the present work, the flow sheet in Figure 1 is proposed for the treatment of phosphate rock with nitric acid for the recovery of uranium and lanthanides as by-products of NPK fertilizer production. The process involves the following steps:

** The extraction of lanthanides from phosphate rock by 0.5 M nitric acid at (L/S =4) and 400 rpm, in two extraction stages each within 5 min the recovered ax oxalate. 96% of lanthanides were leached within two extraction stages, the weight loss in the sample was 27%.

The residue was leached by 3 M nitric acid to leach [P.sub.2][O.sub.5] and a little of uranium.

** The uranium was selectively extracted by 0.1 mol/L D2EHPA/0.1 mol/L TOPO in kerosene with a ratio 4:1 at (O/A) = 1 that was conducted with the nitrate solution for 30 min. at 25 [degrees]C in one stage. The extracted uranium was stripped by 2 M [Na.sub.2]C[O.sub.3].

** The leached solution was treated with KOH to attain the pH 4 to obtain NPK fertilizer that was confirmed to be almost free of uranium and REE.

** The precipitate obtained after KOH treatment was mainly calcium hydroxide.

** The unreacted apatite was 10% of the original sample, the concentration of Y was relatively high (0.18%) may be due to the presence of fluoride that form insoluble potassium Yttrium Fluoride Phosphate KYFP[O.sub.4].

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(1) M.S. Alshammari, (2) I.M. Ahmed, (2) A.A Nayla, (2) H.F. Aly, (3) Gehad G. Mohamed, (3) S. Abdel Rahman Mostafa

(1) Department of Chemistry, Faculty of Science, Al Jouf University, Saudi Arabia.

(2) Hot Laboratories and Waste Management Center, Egyptian Atomic Energy Authority, P. O. Box 13759, Inshas, Cairo, Egypt.

(3) Department of Chemistry, Faculty of Science, Cairo University, P. O. Box Egypt.

Address For Correspondence:

M.S. Alshammari, Department of Chemistry, Faculty of Science, Al Jouf University, Saudi Arabia E-mail: j_s_ksu@hotmail.com_

This work is licensed under the Creative Commons Attribution International License (CC BY). http://creativecommons.org/licenses/by/4.0/

Received 12 July 2016; Accepted 18 September 2016; Available online 22 September 2016

Table 1: Chemical analysis of Al JALAMIDE phosphate
concentrate by X-ray fluorescence.

Main Constituents        (wt,%)

Si[O.sub.2]                2.26
[Al.sub.2][O.sub.3]        0.10
[Fe.sub.2][O.sub.3]        0.12
MgO                        0.63
CaO                       58.29
[Na.sub.2]O                0.18
[K.sub.2]O                 0.02

[P.sub.2][O.sub.5]        27.30
S[O.sub.3]                 1.10
F                          0.74
L.O.I *                    9.21
SrO                        0.05
Zr[O.sub.2]                0.01
[Y.sub.2][O.sub.3]         0.01

L.O.I = Loss of ignition

Table 2: Lanthanide and uranium content in selected
phosphate rock

                                        [U.sub.2]
Phosphate                [Ln.sub.2]     [O.sub.4]
Rock Source             [O.sub.3] (%)    (g/ton)

Russia Kola                0.8-1.0          5
USA Florida-Pebble        0.06-0.29        150
Algeria Djebel Onk        0.13-0.18        110
Morocco Khoribga          0.14-0.16        130
Tunisia                     0.14          40-50
Egypt                       0.028         50-200
Vietnam                     0.031          n.d

Table 3: Recovery of Ln, U and P2O5 after leaching of apatite
using nitric acid, phase ratio L/S 4 at 25 [degrees]C.

                  0.5 M HN[O.sub.3]

Recovery, %       Ln     U      [P.sub.2]
Time                            [O.sub.5]

5 min             61     11.1   4.4
10 min            65.3   9      4.4
20 min            62     10     8.7
30 min            62     12.5   8.7

                  1 M HN[O.sub.3]

Recovery, %       Ln     U      [P.sub.2]
Time                            [O.sub.5]

5 min             82     15.8   13.1
10 min            82     18.8   8.7
20 min            72.6   16.3   8.7
30 min            61     14.9   8.8

                  2 M HN[O.sub.3]

Recovery, %       Ln     U      [P.sub.2]
Time                            [O.sub.5]

5 min             96     27     35.5
10 min            96     24     35.5
20 min            97     30     28.2
30 min            97     22     30

Table 4 Recovery of Ln, U and P2O5 after Leaching of apatite using
0.5 M HN[O.sub.3], phase ratio L/S = 4, 5 min. at 25[degrees]C.

Sample            [Ln]   [U]   [[P.sub.2]   Wt. loss
                               [O.sub.5]]

1st extraction    61%    11%      4.4%         2g
stage

2nd extraction    96%    18%      10%         3.4g
stage
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Author:Alshammari, M.S.; Ahmed, I.M.; Nayla, A.A.; Aly, H.F.; Mohamed, Gehad G.; Mostafa, S. Abdel Rahman
Publication:Advances in Environmental Biology
Article Type:Report
Date:Sep 1, 2016
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