الفهرس 
INTRODUCTION AND LITERATURE SURVEYOnly 14 pages are availabe for public view
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Abstract
SUMMARY
The chemical processing of the poly-mineralized carbonate-rich latosol ore material occurring at Abu Thor locality of Southwestern Sinai area has been studied for the recovery of uranium, copper, vanadium and lanthanides (Lns) metal values.
To realize this goal, a technological representative sample of Abu Thor mineralization was properly collected and which was found to assay 700 ppm U and up to 9.7% Cu, 789 ppm V and 456 Lns .
The present work has thus been firstly oriented towards the recovery of both uranium and copper from the study ore material by using two successive percolation leaching procedures. For the former, urea was used as a selective non-conventional and economic organic solvent followed by ammonium hydroxide solution for selective leaching of the associated copper. Beside the agitation leaching procedure , the characteristics of the percolation leaching technique have also been determined for both U and Cu metal values.
To optimize the agitation leaching relevant parameters for uranium by using urea, different series of agitation leaching experiments have been performed. In these experiments, a suitable weight of the working ore material, ground to -60 mesh size, was -I
Summary
mixed with a suitable volume of urea of various concentrations. The slurry was then agitated for different contact time periods and solid /liquid ratio ( S/L) at a certain temperature and the obtained slurry was then filtered and the residue left behind was thoroughly washed with distilled water and both filtrate and washings were made up to volume before analysis of the leached uranium.
The optimum leaching conditions have been determined using an ore ground to -60 mesh size and was found to include a urea solutions of concentration 60g/L at room temp. for 90 min. in a S/L ratio of 1/5. Under these conditions the uranium leaching efficiency was found to attain up to 97.5%. The uranium leached ore sample was then subjected to almost complete Cu dissolution 98.9% using the optimum conditions of 120 g/L ammonia solution and 35 g/L ammonium carbonate for 2 hours in a S/L of 1/5.
After determination of optimum agitation leaching conditions, the obtained results were then be applied to successive percolation leaching procedures for both metal values, the percolation leaching experiment has been carried out upon 3 Kg of the working latosol ore material which was mixed with 1.5 Kg of sand and packed in a column, of length 300cm, outer diameter is 6 cm and inner diameter is 5.5 cm and the height of the ore is 270 cm. selective uranium percolation leaching was then performed using 15L of 60 g/ L urea leaching solution with flow rate of 1.6 ml/min. at room temperature. The processed ore material was washed by 300 ml of distilled water and the collected overall uranium leaching efficiency reached
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Summary
90.27 %, while copper did not respond to the applied conditions.
The Collected uranium leach liquor was then subjected to the
precipitation step by increasing its pH up to the range of 12-13
using NH4OH solution where ammonium di- urinate was obtained.
After the filtration and drying the latter, analysis has indicting its
high purity with about 68% of uranium content.
The uranium leached ore sample of 3 Kg mixed with 1.5 Kg
sand and containing about 290 g Cu was then subjected to copper
leaching using the studied ammonium hydroxide / carbonate mixed
leach solution assaying 120 and 35 g/L respectively. The obtained
flow rate was as that used during uranium leaching i.e 96 ml / h
(1.6 ml/min.) and the total applied leaching solution was 3 liter.
The obtained leaching results indicate a total leaching efficiency of
about 93%. Then the pH of collected copper leach liquor has then
been decreased down to 5.5 using sulfuric acid solution to obtain the
interesting crystals of copper sulfate. The latter was then subjected
to ESEM-EDX qualitative analysis to insure high purity.
After the recovery of U and Cu, from the study raw material,
the dried residue was then subjected to another two successive
recovery procedures for V and Lns. These remaining two metal
values which were left behind in the spent residue were found to
assay 1050 and 649 ppm for V and Lns respectively. So, the present
work was then shifted to study the recovery of both V and Lns metal
values from the spent residual material. These involved V recovery
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Summary
by alkaline leaching followed by classical acidic leaching
procedures for Lns. from the latter a procedure for separating V
from the accompanying Lns can be achieved by breaking down the
study spent ore residue by using NaOH at the optimum conditions
of spent ore residue / NaOH ratio: 1/3, 3 h roasting time and 200
°C roasting temperature . To prepare a rich V leach liquor, 600 g
of the spent ore residue was carefully mixed with 1800 g NaOH
and roasted for 3 h at 200 °C. The roasted cake was then leached
with 600 ml distilled water to obtain the dissolved vanadate ions .
Under these condition, the dissolution efficiency of the vanadium
attained 98.8 %. The obtained vanadate solution may contain small
amounts of V(IV), which can be quickly oxidized to V(V). This was
accomplished by the addition of 30 % HR2ROR2R to the prepared solution
to oxidize V(IV) to V(V).The pH of the latter solution was then
adjusted to pH 8 by using 30 % HR2RSOR4R to obtain vanadium
precipitate . The latter precipitate has been subjected to
calcinations at 800°C to form VR2ROR5R.
For separating 96.2% of Lns from the spent ore residue free
from V, this can be achieved by using H2SO4 80 g/L concentration
,120 min. stirring time within S/L: 1/3 at room temp.
To prepare a rich Lns leach liquor, 3000 g of spent ore residue
free from V was subjected to sulfuric acid agitation leaching under
the obtained leaching conditions. A volume of 9 litter was obtained
assying 208 mg Lns / L. To increase the concentration of Lns, this
leach liquor was then subjected to evaporation process to reduce the
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Summary
volume to 3 liters in order to obtain a Lns- oxalate concentrate . The latter was accomplished by adjusting the pH of the solution to 4.5, then by the addition of 10% oxalic acid ,Lns oxalate was precipitated at pH 1. The obtained Lns oxalate precipitate was then calcined at 800 °C for 3 h and analyzed by using ESEM-EDEX technique to insure its high purity