Translate this page into:
Recovery of alumina and some heavy metals from sulfate liquor
⁎Corresponding author at: Nuclear Materials Authority, P.O. Box 530, El Maadi, Cairo, Egypt. monaelhazek@hotmail.com (M.N. El Hazek)
-
Received: ,
Accepted: ,
This article was originally published by Elsevier and was migrated to Scientific Scholar after the change of Publisher.
Peer review under responsibility of King Saud University.
Abstract
The gibbsite bearing shale occurrence in the Paleozoic sedimentary sequence of SW Sinai, Egypt, was found to be associated with several metal values. From sulfate liquor prepared by proper leaching, the recovery of these metal values has been studied. Alumina was first separated in the form of potash alum followed by Cu-selective extraction by hydroxyoxime LIX-973N solvent. Then U recovery using an anionic exchange resin Amberlite IRA-400 was achieved. For the associated heavy metal Zn, it was subsequently extracted using di-2-ethylhexyl phosphoric acid. The relevant factors affecting the extraction process were adequately studied.
Keywords
Cu
Zn
Al
U
Solvent extraction
Resin Amberlite
DEHPA
LIX-973
1 Introduction
After studying the proper leaching conditions of Abu Zeneima metalliferous gibbsite ore material in a previous work (El Hazek et al., 2008), a proper sulfate leach liquor of the contained metal values (Al, Cu, Zn, U, Co and Ni) has been prepared for studying their recovery in the present work. There are indeed several recovery procedures that can be applied, and the choice of the convenient procedure for each metal value would actually depend upon its concentration, the associated recoverable metal values, presence of other impurities such as iron besides the nature of the pregnant liquor. Taking these factors into consideration, it was found convenient to first recover Al from the acidic medium through potash alum crystallization. This has been performed by adding KOH which would be converted into K2SO4 by partial neutralization of the present acid. Starting by alum manufacture is greatly advantageous, as the density and acidity of the working leach liquor would be reduced. On the other hand, the concentration of the studied metal values will in turn be increased during alum preparation due to the required evaporation.
On the other hand, while the ion exchange procedure was found quite suitable for U due to its relatively low concentration, solvent extraction would be the preferable procedure for the studied heavy metals. Accordingly, the extractant LIX-973N would be used for Cu recovery whereas D2EHPA would be used for Zn recovery. It is worth mentioning herein that either of these solvents can also be used for Co and Ni extraction at proper working conditions. The relevant studied extraction and stripping factors of both Cu and Zn such as pH, concentration of the extracting and the stripping reagents as well as the contact time and the O/A ratio have been properly studied.
Several works have indeed been performed upon Abu Zeneima mineralized ore material for the recovery of the different contained metal values. Ritcey Ritcey (1991) has studied Cu recovery using LIX-973N and U by a tertiary amine from their sulfate liquor. Amer (1993) has also applied the ion exchange technique for U using Amberlite IRA-400 anion exchange resin. Mahdy (1995) has presented a number of flowsheets for the recovery of U, Cu and Mn using different methods such as precipitation, crystallization besides using the organic solvents LIX-64N and TBP as well as the anion exchange resin. On the other hand, Amer (1997) and Amer et al. (2000) have also studied Cu and U recovery from sulfate liquor. In addition, Abdel Fattah (2003) has studied the leaching and recovery of Al, Cu, Zn and U from acidic sulfate liquor as well as from caustic soda liquor. In this work, Al, Cu and Zn were recovered by crystallization and solvent extraction techniques while for U the Amberlite IRA-400 anion exchange resin has been used.
2 Experimental
2.1 Material and reagents
To study the applicability of the above-mentioned procedures in the present work for Al, U, Cu and Zn recovery, a suitable leach liquor of the working metal values of Abu Zeneima gibbsite ore was first prepared using mostly the studied optimum leaching conditions (400 g/L acid, S/L ratio of 1/3 at 80 °C for 4 h). Analysis of the leach liquor obtained under these conditions indicated the following assay: Al 28.2 g/L, Zn 4.725 g/L, Cu 400 ppm and U 75 ppm while Co and Ni assay amounted to 150 and 165 ppm respectively.
The reagents used involve the specific Cu Chelating extractant LIX-973N belonging to the hydroxyoximes (product of M/S. Cognis) while the acidic extractant D2EHPA belongs to the organophosphorus compounds and (product of Merric Co.) was used for Zn extraction. Both solvents were diluted in local kerosene produced by Misr Petroleum Co. For U recovery, Amberlite IRA-400 anion exchange resin (product of Rohm and Haas Co., USA) has been used.
2.2 Extraction and stripping procedures
Extraction and stripping tests were performed in separatory funnels where the prepared organic phase and the aqueous leach liquor on the loaded organic and the stripping solution were shaken for a proper time. The two phases after equilibration were then allowed to separate and an aliquot sample of the aqueous phase was analyzed for its metal content while that in the organic phase was obtained by a difference.
2.3 Analytical procedures
For the analysis of Al, Cu and Zn, the atomic absorption technique was adopted using a Unicam atomic absorption spectrophotometer model 969 flame type, auto gas box at wavelengths 309.30, 222.60 and 307.70 nm respectively (Weltz and Sperling, 1999). On the other hand, for U determination, an oxidimetric titration method against ammonium metavanadate was used in the presence of diphenylamine sulfonate indicator. Prior to titration, proper reduction of U was performed using ammonium ferrous sulfate (Mahmoud et al., 2003).
3 Results and discussion
3.1 Alum crystallization
Potash alum is preferred than sodium alum for Al recovery in the present work as alum. This is due to the fact that the latter is extremely difficult to purify and is much more soluble than the former (www.google.com). On the other hand, from the solubilities of K2SO4, Al2(SO4)3·9H2O and potash alum in water as a function of temperature, it is evident that the solubility of all the three compounds increases with an increase in temperature. At high temperatures (>70 °C), potash alum is the most soluble among the three salts, while K2SO4 is the least soluble which will thus be the first to precipitate from solution. On the other hand, at low temperatures (<46 °C), alum is the least soluble of the three salts and will be the first to precipitate.
For potash alum preparation in the present work, 2 L of the prepared leach liquor assaying 28.2 g/L Al was mixed with about 0.5 L containing 116 g of KOH. This solution was then slowly evaporated until 1 L at a temperature not exceeding 45 °C. To ensure separation of the least soluble alum, the evaporated solution was left overnight and the obtained alum was filtered and properly washed with methanol, dried and weighed. From the obtained weight of 660 g, a crystallization efficiency of about 70% has been obtained (calculated after taking in consideration that potash alum solubility at 4 °C is 39 g alum/L). Leaving the solution to further cooling after 24 h, further alum crystallization has been occurred.
3.2 Copper recovery
The volume of the 2 L working pregnant liquor after alum crystallization has been decreased to 1 L due to evaporation and therefore the concentration of all the dissolved metal values has been doubled. The Cu content of the latter has thus been increased to 800 ppm and its pH was found to attain 0.5.
3.2.1 Copper extraction
3.2.1.1 Effect of aqueous phase pH
The effect of pH on Cu extraction with LIX-973N was carried out in the range from 0.2 to 2.0 as indicated by Calligaro et al. (1983) using NaOH or H2SO4 solution. The working extraction conditions involved 1/1 as organic/aqueous ratio, 5% solvent concentration in kerosene as a diluent and a shaking time of 5 min. It was observed from the obtained results plotted in Fig. 1, that Cu extraction does not greatly depend on the pH of the aqueous phase in the tested range (0.2–2.0). At pH 1, the Cu extraction efficiency attained 98.2% (D = 55). The latter could be considered as an optimum value in a manner to avoid increasing the pH and in turn possible interference from iron and/or its precipitation. The remaining Cu at this pH can be recovered by variation of other conditions.
Effect of pH on Cu extraction efficiency using LIX-973N.
3.2.1.2 Effect of LIX-973N concentration
The effect of LIX-973N concentration on Cu extraction efficiency has been studied in the range from 1% to 5% (v/v) in kerosene at pH 1 and using an O/A ratio of 1/1 for a shaking time of 5 min. From the obtained results plotted in Fig. 2, it is clear that Cu extraction efficiency increases directly with an increase in the extractant concentration. Thus increasing the solvent concentration from 1% to 3% has resulted in increasing the Cu extraction efficiency from 89.1% to 97.5% and the distribution coefficient from 8 to 39. By further increasing the extractant concentration from 3% to 5%, a slight increase was observed where 98.2% Cu extraction efficiency was obtained at a D value of 55.
Effect of LIX-973N concentration on Cu extraction efficiency.
3.2.1.3 Effect of contact time
A number of experiments have been studied to determine the effect of the shaking time from 1 to 5 min upon Cu extraction efficiency. The obtained results shown in Fig. 3 reveal that by increasing the contact time from 1 to 3 min, the Cu extraction efficiency has increased from 93.3% to 97.0%. Increasing the contact time to 5 min, the Cu extraction efficiency was only slightly increased from 97.0 to 97.5. It is interesting herein to refer the comparable work performed by Amer (1997) who has reported that almost complete Cu extraction (98.1%) was obtained in the first 30 s from an aqueous solution assaying 450 ppm Cu at pH 1.1 by 3% LIX-973N in kerosene.
Effect of contact time on Cu extraction efficiency.
3.2.1.4 Effect of O/A ratio and construction of McCabe–Thiele diagram
The effect of O/A ratio upon Cu extraction efficiency was studied in the range of 3/1 to 1/10 using 1% (v/v) LIX-973N in kerosene for a 5 min contact time after adjusting the pH of the aqueous phase at 1. From the obtained results shown in Table 1 and plotted in Fig. 4, it is clearly evident that varying the O/A ratio it would be possible to realize Cu loading of the organic phase with up to about 4 g/L; a matter which would realize a concentration factor of 5.
O/A ratio
Cu conc. (ppm)
Ext. coeff. (D)
Ext. eff. (%)
Aqueous
Organic
3/1
24.0
258.7
10.8
97.0
2/1
40.0
380.0
9.5
95.0
1/1
87.2
712.8
8.2
89.1
1/2
120.0
1360.0
11.3
85.0
1/3
141.6
1975.2
13.9
82.3
1/4
172.8
2508.8
14.5
78.4
1/5
258.4
2708.0
10.5
67.7
1/10
400.8
3992.0
10.0
49.9
McCabe–Thiele diagram for Cu extraction from Abu Zeneima gibbsite leach liquor.
On the other hand, for plotting the McCabe–Thiele diagram, the obtained equilibrium data at different O/A ratios would result in Cu equilibrium isotherm. To the latter, a suitable operating line was fitted with an A/O slope representing the counter current flow rate of A/O of about 6 and a final loading of about 4.8 g Cu/L. Stepping off the possible extraction stages starting from the input aqueous feed Cu concentration, it was found that four stages would realize a depletion of the aqueous phase down to 24 ppm Cu.
3.2.2 Copper stripping
To obtain Cu as a marketable copper sulfate product or for electrowinning, the available and cheap H2SO4 solution was herein used as a suitable stripping agent. In this regard, the relevant stripping factors including the stripping contact time, the concentration of the stripping agent, and the aqueous to organic volume ratio have been studied using a Cu loaded LIX-973N extractant assaying 776 ppm Cu.
3.2.2.1 Effect of contact time upon Cu stripping
The effect of stripping contact time on the Cu stripping efficiency from the loaded LIX-973N has been studied using 5 M acid in the range from 0.5 to 10 min at an A/O ratio of 1/1. From the results plotted in Fig. 5, it was shown that the stripping efficiency increased by increasing the contact time. Thus at 0.5 min, a stripping efficiency of 82.6% was obtained which was steadily increased until 93% at 10 min contact time. However, a contact time of 3 min could be considered sufficient and the stripping efficiency could be increased by varying the other stripping conditions.
Effect of contact time on Cu stripping efficiency.
3.2.2.2 Effect of H2SO4 acid concentration upon Cu stripping
The effect of H2SO4 concentration upon the stripping efficiency of Cu was studied in the range from 0.5 to 5 M. The other stripping conditions were fixed at an A/O ratio of 1/1 and a stripping contact time of 3 min. The obtained results illustrated in Fig. 6, reveal that the Cu stripping efficiency was increased from about 69% to 98% by increasing the acid concentration from 0.5 to 5 M. However, to economize acid consumption, 2.5 M acid which resulted in about 86% Cu stripping could be adequate and complete stripping could be achieved by varying the other conditions.
Effect of H2SO4 acid concentration on Cu stripping efficiency.
3.2.2.3 Effect of A/O ratio and construction of McCabe–Thiele diagram
To study the effect of the A/O ratio on the copper stripping efficiency, a number of stripping experiments were performed at A/O ratios varying from 1/3 to 3/1. The other stripping factors were fixed at 3 min contact time while 2.5 M sulfuric acid was found convenient as mentioned above.
In Table 2, it is clear that by increasing the A/O ratio from 1/3 to 3/1, the stripping efficiency has increased from about 49% to complete Cu stripping respectively. To construct the corresponding McCabe–Thiele diagram, the obtained equilibrium data have been plotted in the form of an equilibrium isotherm and a proper operating line has been fitted. The slope of the latter which would thus represent the counter current aqueous and organic (A/O) flow rate was found to attain about 0.65 (Fig. 7). From the latter, it is clearly evident that three stripping stages would bring a Cu concentration in the stripping solution of about 1.2 g/L.
H2SO4 conc. (M)
Cu conc. (ppm)
Strip. coeff. (S)
Strip. eff. (%)
Aqueous
Organic
1/3
1131.4
398.9
2.8
48.6
1/2
939.0
306.5
3.1
60.5
1/1
668.9
107.1
6.2
86.2
2/1
383.0
10.0
38.3
98.7
3/1
258.6
0.08
>3000
100.0
McCabe–Thiele diagram for Cu stripping from Cu-loaded LIX-973N in kerosene.
The optimum conditions for Cu extraction would involve a pH of 1, LIX-973N concentration of 3% and 3 min contact time. Under these conditions, 97.0% of Cu extraction efficiency has been realized. The optimum conditions for Cu stripping would involve a contact time of 3 min and 2.5 M sulfuric acid as well as 1/1 A/O ratio resulted in about 86% Cu stripping.
3.3 Uranium recovery
After alum and Cu separation from the working pregnant leach liquor, its pH was increased to 1.8 by NaOH before U recovery by ion exchange resin. As previously mentioned, the latter was found to be the optimum choice due to the low U concentration which is only as low as about 150 ppm. However, in order to stimulate the expected U concentration in the leach liquor and to obtain the corresponding results, the U concentration was increased in the working liquor to attain 464 ppm by adding a calculated amount of UO2(NO3)2·6H2O.
3.3.1 Uranium adsorption
A volume of 5 ml wet settled Amberlite IRA-400 anion exchange resin was properly packed in a suitable glass column and thoroughly washed with distilled water. The pH of the working leach liquor after Cu extraction was adjusted to 1.8 and was then passed through the prepared column until the effluent U content became equal to that of the influent. Adjustment of the pH to 1.8 would decrease the
content, which has indeed a high affinity for the resin. A contact time of 3 min equivalent to a flow rate of 0.66 ml/min was used and the effluent sample fractions were collected every 50 ml. In the latter, U was oxidimetrically analyzed using NH4VO3 after its prior reduction by ammonium ferrous sulfate and using diphenylamine sulfonate as indicator and the obtained results are plotted in Fig. 8 in the form of the corresponding adsorption or loading curve.
Uranium adsorption curve of Abu Zeneima gibbsite leach liquor.
Calculation of the total U adsorbed after saturation of the resin volume used indicated that the total U adsorbed attained 316.8 mg. Referring to the theoretical capacity of Amberlite IRA-400 of 1.56 mequiv./ml, the U adsorption capacity would amount to 92.8 mg U/ml resin, i.e. 464 mg U/5 ml wsr (wet settled resin) sample. However, the total U adsorbed of 316.8 would represent an adsorption efficiency of 68%. This relatively low efficiency is primarily due to excessive sulfate concentration (∼100 g/L) which would compete with the uranyl sulfate complex for the available exchange sites. On the other hand, the relatively wide exchange zone between the U breakthrough at the 11th sample (96.5% adsorption) and U saturation at the 20th sample can also be due to this competition.
3.3.2 Uranium elution
After the resin sample was saturated with U, the loaded resin bed column was first washed with distilled water to drain off the influent liquor. The resin bed was then eluted by a proper eluant; namely 1 N NaCl solution acidified to 0.1 M with H2SO4 using 5 min as a contact time equivalent to a flow rate of 0.4 ml/min and the obtained eluate sample fractions were collected every 5 ml for U analysis. The obtained results are plotted in Fig. 9 in the form of the corresponding elution curve. From the latter, it is clearly evident that the total eluted U amount attained 298 mg which is almost equal to the adsorbed amount (316.8 mg) and the difference might be due to analytical error. On the other hand, a maximum U concentration of up to 14.26 g/L was obtained in the third eluate sample. In the meantime, the U assay in the mixed solution of sample Nos. 2 through 6 would assay 10.68 g/L; a concentration which is quite suitable for U precipitation. In these samples, the elution efficiency attained more than 86%. Therefore, U can be precipitated from only these five eluate samples while the other eluate samples which are of low U concentration can be recycled for elution of a next saturated U resin bed (split elution technique).
Uranium elution curve of 5 ml saturated Amberlite IRA-400 resin bed.
The optimum conditions for U loading would involve pH 1.8 of the feed solution, a contact time of 3 min equivalent to a flow rate of 0.66 ml/min. The optimum conditions for U elution would involve 1 N NaCl solution acidified to 0.1 M with H2SO4 using 5 min as a contact time equivalent to a flow rate of 0.4 ml/min.
3.4 Zinc recovery
The uranium effluent solution assaying 9.45 Zn besides 300 and 330 ppm Co and Ni, respectively and of pH 1.8 was then directed to solvent extraction for the recovery of Zn while the latter were left for later work due to their low assay besides lack of material. For this purpose, di-2-ethylhexyl phosphoric acid (D2EHPA) was used for Zn recovery by several authors (Ritcey, 1983; Thorsen, 1983; Cheng, 2000). According to Ritcey and Ashbrook (1979), the order of extraction in D2EHPA extractant of the interesting metals is given as follows:
Therefore, it was found necessary to eliminate iron (and most of the remaining Al) from the above treated Abu Zeneima gibbsite sulfate liquor by a prior precipitation step at pH 3.5
3.4.1 Zinc extraction
The obtained iron filtrate was thus ready for Zn extraction (as well as for Co and Ni) using D2EHPA in kerosene and the corresponding relevant factors have been studied. For this purpose, the contact time was first studied between 1 and 15 min using 30% D2EHPA in kerosene at an O/A ratio of 1/1 after pH re-adjustment to 2.0. The obtained results indicate that equilibrium is indeed rapidly attained at 1–4 min, the extraction efficiency amounted to about 46–50%. Increasing the contact time to 6, 8 and 10 min did not almost change the extraction efficiency and at 15 min, the latter increased to only 53%. Therefore, 4–5 min as contact time was chosen as an optimum contact time.
3.4.1.1 Effect of pH
The pH of different sample solutions of the working Abu Zeneima iron filtrate (pH 3.5) was adjusted to different values by H2SO4 acid solution from 1 to 3.5 to determine its effect upon Zn extraction. The other extraction conditions were fixed as 30% D2EHPA in kerosene, a contact time of 5 min and an O/A ratio of 1/1. From the obtained results plotted in Fig. 10, it is clearly evident that at pH 2.5 the extraction efficiency amounted to only 56.8 (D = 1.32) and remain unchanged at pH 3.5. This low value is actually most probably due to an increased H+ transfer from the organic phase to the aqueous phase by Zn extraction; a matter which would compete with Zn for extraction. Accordingly, it would be necessary to perform the extraction at a constant optimum final pH (equilibrium pH).
Effect of pH on Zn extraction efficiency using D2EHPA.
3.4.1.2 Effect of D2EHPA concentration
To study the effect of D2EHPA concentration in the organic phase upon Zn extraction, a number of shake out tests were performed using D2EHPA concentration varying from 10% to 70% in kerosene. The other extraction conditions were fixed at pH 2.0 of the aqueous phase, an O/A ratio of 1/1 and a contact time of 5 min. From the obtained results plotted in Fig. 11, it was found that only about 21% Zn extraction occurred at 10% D2EHPA (D = 0.26) while about 51% was realized at 30% D2EHPA. Increasing the latter concentration to 50% and 60% increased the extraction efficiency to about 64% and 70% with a slight decrease at 70% D2EHPA; a matter that might be due to solubility effects of the formed Zn complex in the organic phase. In addition, the latter is also most probably due to pH decrease of the aqueous phase as mentioned above. Therefore, 30% D2EHPA could be considered as an optimum concentration and at which the extraction efficiency could be improved by other extraction factors.
Effect of D2EHPA concentration on Zn extraction efficiency.
3.4.1.3 Effect of O/A ratio and construction of McCabe–Thiele diagram
The effect of the O/A ratio upon Zn extraction from Abu Zeneima sulfate leach liquor was studied in the range from 5/1 down to 1/5. The other extraction conditions were fixed at 30% D2EHPA concentration, pH 2.0 of the aqueous phase and a contact time of 5 min. The obtained results are shown in Table 3 and plotted in Fig. 12 as the corresponding equilibrium isotherm.
O/A ratio
Zn conc. (g/L)
Ext. coeff. (D)
Ext. eff. (%)
Aqueous
Organic
5/1
2.77
1.34
0.48
70.7
4/1
2.80
1.67
0.59
70.4
3/1
2.75
2.23
0.81
70.9
2/1
3.65
2.90
0.80
61.4
1/1
4.65
4.80
1.03
50.8
1/2
5.82
7.26
1.25
38.4
1/3
7.09
7.08
1.00
25.0
1/5
7.41
10.21
1.83
21.6
McCabe–Thiele diagram for Zn extraction by D2EHPA from the sulfate liquor of Abu Zeneima gibbsite-bearing shale.
To construct the corresponding McCabe–Thiele diagram, a proper operating line whose slope would represent the counter current flow rate of the organic and the aqueous phase has been plotted. The slope of the latter (O/A flow rate) was found to attain about 1.3 and would result in about four extraction stages.
The difficulty in decreasing the Zn assay in the raffinate below about 2 g/L is mainly due to the decrease in pH of the aqueous phase due to H+ transfer as previously mentioned. Also the same reason can interpret the relatively decreased Zn saturation in the organic phase which, stoichiometrically, could have been saturated with 18.9 g Zn/L.
3.4.2 Zinc stripping
To study Zn stripping from the loaded 30% D2EHPA phase, a proper loaded organic phase sample assaying 10.21 g Zn/L was first prepared (maximum obtained saturation at an O/A ratio of 1/5). Working with 0.2 M sulfuric acid solution, the effect of contact time upon Zn stripping efficiency was first studied between 1 and 15 min at an A/O ratio of 1/1. The obtained results indicated that Zn stripping is adequately rapid where 49% was obtained at 1–2 min and which was only increased to about 50% at 3–5 min and to 53% at 7–10 min and only to 53.3% at 15 min.
3.4.2.1 Effect of sulfuric acid concentration
A number of stripping experiments were performed at an A/O ratio of 1/1 and for a contact time of 5 min while varying the H2SO4 acid concentration from 0.1 to 1 M. From the obtained results illustrated in Fig. 13, it was revealed that the stripping efficiency increased steadily from 33.1% to 90.6% by increasing the acid concentration from 0.1 to 0.5 M. Increasing the latter to 0.7 increased the stripping efficiency to 96.7% and by using 1.0 M acid, a slight increase in the stripping efficiency to 97.0% was only obtained. Thus, 0.2 M acid can be considered as an adequate acidity for Zn stripping from D2EHPA solvent. Using the latter, the stripping efficiency of 50% could be improved by other stripping factors.
Effect of H2SO4 acid concentration on Zn stripping efficiency.
3.4.2.1.1 Effect of A/O ratio and construction of McCabe–Thiele diagram
To study the effect of A/O ratio upon the Zn stripping efficiency, the working loaded organic phase sample assaying 10.21 g Zn/L was stripped by different volumes of 0.2 M H2SO4 acid at a contact time of 5 min. The obtained results shown in Table 4 and plotted as the corresponding equilibrium isotherm in Fig. 14 indicate that at A/O ratios of 2/1 and 3/1, stripping efficiencies of 89.3% and 94.3% have been obtained respectively, while at 4/1 ratio complete stripping was achieved.
A/O ratio
Zn conc. (ppm)
Strip. coeff. (S)
Strip. eff. (%)
Aqueous
Organic
1/3
6307.3
8103.6
0.78
20.6
1/2
6531.8
6940.1
0.94
32.0
1/1
5103.0
5103.0
1.0
50.0
2/1
4556.9
1092.0
4.17
89.3
3/1
3208.1
581.7
5.51
94.3
4/1
2041.2
0
–
100.0
McCabe–Thiele diagram for Zn stripping from Zn-loaded D2EHPA in kerosene.
For constructing the corresponding McCabe–Thiele diagram, a proper operating line whose slope represents the counter current flow rate of the aqueous and organic phases is then fitted to the plotted equilibrium isotherm. From the latter, it is evident that the A/O flow rate would be about 1.2 besides indicating three stripping stages.
The optimum conditions for Zn extraction would involve 30% D2EHPA concentration, pH 2.0 of the aqueous phase and a contact time of 5 min. The optimum conditions for Zn stripping would involve a contact time of 3 min and 0.2 M sulfuric acid.
4 Conclusion
The proper procedures studied for the recovery of Al, Cu, U and Zn have shown to be greatly applicable upon the sulfate leach liquor of Abu Zeneima gibbsite-bearing shale. Potash alum was first crystallized followed by Cu extraction using the LIX-973N extractant while U was recovered by ion exchange resin and Zn was finally recovered by D2EHPA. The relevant recovery factors have indeed been studied for most of these metal values. The overall required operations are properly summarized in a proposed flowsheet shown in Fig. 15.
Recovery of alumina and some heavy metals from sulfate liquor Abu Zeneima.
References
- Sulfuric acid leaching of polymetallic Abu Zeneima gibbsite shale. Hydrometallurgy. 2008;90:34-39.
- [Google Scholar]
- Ritcey, G.M., 1991. Potential for Yellow Cake Production, Report No. Ru 2780 on the Project EGY/3/010 IAEA, Vienna.
- Amer, T.E., 1993. Physical and Chemical Studies on the U–Cu Mineralization of Uraniferous Paleozoic Sediments, West Central Sinai; M.Sc. Thesis, Fac. Sc., Cairo Univ., Cairo.
- Mahdy, M.A., 1995. Uranium, Copper and Manganese Recovery from sulphate Leach Liquor of West Central Sinai Uraniferous Siltstone. Egypt. J. Anal. Chem. 4.
- Amer, T.E., 1997. Geochemistry and Extraction of U, Cu and Mn from the Ore Materials of the Uraniferous Paleozoic Sedimentary Rock, West Central Sinai, Egypt, Ph.D. Thesis. Fac. Sc., Cairo Univ., Cairo.
- Amer, T.E., Mahdy, M.A., El Hazek, N.T., El Bayoumi. R., Hassanien, S., 2000. Extraction of Uranium and Copper from sulphate leach liquors of West Central Sinai Uraniferous Siltstone of Egypt. In: Proc. Int. Symp. Process Metall. of Uranium, Saskatoon, Saskatchewan, Canada.
- Abdel Fattah, N.A., 2003. Uranium Extraction from Gibbsite-Bearing Sediments and Other Economic Elements by Using Solvents and Organic Compounds (Abu Zeneima Area, Sinai, Egypt). M.Sc. Thesis, Fac. Sc., Zagazig Univ., Egypt.
- Atomic Absorption Spectrometry (third ed.). Weinheim, New York: Wiley-VCH; 1999.
- Mahmoud, K.F., 2003. Simulated Heap Leaching studies of Qatar Uranium Occurrences. Int. Rep. N.M.A., Cairo.
- Preparation of Alum (SUSB-030). Available at: <www.google.com>.
- Calligaro, L., Mantovani, A., Belluco, U., 1983. Solvent extraction of copper II, nickel II, cobalt II, zinc II, and iron III by high molecular weight hydroxyoximes. Polyhedron 2.
- Commercial processes for Nickel and Cobalt. In: Lo Teh C., Baird Malcolm H.I., Hanson Carl, eds. Handbook of Solvent Extraction. New York: J. Wiley and Sons; 1983.
- [Google Scholar]
- Commercial processes for Cadmium and Zinc. In: Lo Teh C., Baird Malcolm H.I., Hanson Carl, eds. Handbook of Solvent Extraction. New York: J. Wiley and Sons; 1983.
- [Google Scholar]
- Purification of synthetic laterite leach solution by solvent extraction using DEHPA. Hydrometallurgy. 2000;56:369-386.
- [Google Scholar]
- Solvent Extraction: Principles and Application to Processes Metallurgy: Part II. Amsterdam, Oxford, New York: Elsevier Sc. Publ. Co.; 1979.
