الفهرس 
Literature Survey and Theoretical AspectsOnly 14 pages are availabe for public view
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Abstract
Summary and Conclusions
S i n ai Peninsula is considered as one of the most important
mineralized areas in Egypt. It contains various mineral deposits e.g. Mn,
Fe, Cu, S and C beside kaolin, glass sands, beach sands and limestone,
which are widely distributed along the Gulf of Suez especially at East
Abu Zeneima area.
A bu Hamata sandy claystone rocks of Um Bogma Formation of
southwestern Sinai are considered as one of the most promising
occurrences for U beside the associated economic metal values e.g. Ln,
V, Zn, Cu,…. etc. as well as B which is recorded for the first time in the
present work.
T he main target of the present work; is to study the processing of
this ore material to recover its content of U, Ln, B and V, which are
important for many technological and nuclear applications.
For this purpose, a technological representative sample from Abu
Hamata sandy claystone ore material was collected. The mineralogical
analysis of this sample revealed the presence of montmorillonite as the
main clay mineral (K,Ca)Al2SiO4 together with quartz (SiO2) and
hematite (Fe2O3). The secondary U mineral renardite Pb(UO2)2 (PO4)2.
9H2O was recorded for the first time in the studied area by the present
work. On the other hand, it is worthy to be mentioned herein that there is
no identified minerals for Ln, B and V which are probably adsorbed upon
the clay mineral, montmorillonite.
C omplete chemical analysis of the technological sample shows
that, it is mainly composed of SiO2 which represents the main constituent
(55.6 %) beside Al2O3 (12.70 %) and Fe2O3 (9.81 %). K2O and MgO
represent the much lesser constituents of the study ore material.
Moreover, the sample was found to contain trace amounts of CaO, TiO2,
P2O5 and Na2O. It is worthy to be mentioned herein that the high contents
of SiO2, Al2O3 together with loss of ignition (11.1 %) and K2O reflect
the sandy clay nature of the studied ore material. Also, the chemical
composition of the study ore material shows the presence of 0.183 % of
U, 0. 137 % of Ln, 0.11 % of B beside 0.175 % of V.
W i th respect to recovery study, H2SO4 acid agitation leaching
process was applied upon the raw ore material for dissolving both of U
and Ln with dissolution efficiencies of 98.8 and 93.1%, respectively. The
applied optimum leaching conditions were 5 % H2SO4 acid concentration,
S/L ratio of 1/2 with stirring time for 1 h at leaching temperature of 75
oC. On the other hand, the alkaline NaOH roasting process followed by
distilled H2O leaching was applied upon the spent ore residue to recover
both of B and V contents. The applied roasting conditions were sample/
NaOH ratio of 1/1, 1.5 h roasting time at 140 oC roasting temperature.
The roasted matrix was then leached with distilled H2O for 30 min for
dissolving 94.1 % of B and 97 % of V.
Two different leach liquors were prepared to perform the suitable
extraction technique for the recovery of the metals of interest. These
leach liquors include sulfuric acid leach liquor for U and Ln recovery and
alkaline solution for the recovery of B and V.
The prepared sulfuric acid leach liquor was obtained by processing
1kg of the raw ore sample, assays 0.45 g/L of U, 0.32 g/L of Ln, 26 g/L
of free SO4
2- anions and 2.3 g/L of Cl- anions. This solution was firstly
treated with solid CaCO3 to minimize the concentration of the highly
Chapter V Summary and Conclusion
120
interfering anions, especially SO4
2- via precipitation as CaSO4.2H2O.
However, U was recovered via anion exchange resin by using Amberlite
IRA400 from the treated sulfate solution. More than 94 % of the U content
was extracted at optimum loading conditions of pH 1.8 and contact time
of 4 min. The saturated resin was directed to the elution process by using
an eluant solution of 1M NaCl acidified with 0.1M H2SO4 at 13.3 min
contact time to eluate about 99 % of the loaded U. Almost complete U
precipitation (98.9 %) takes place by gradual addition of H2O2 solution
together with continues stirring for 3 h at the optimum pH value of 1.
After filtration, the prepared UO2(O)2.2H2O was washed, dried and
ignited at 850 oC for 1 h to obtain high pure U3O8..
The effluent solution free from U was subjected to Ln recovery via
their precipitation as Ln2(C2O4)3 by using 10 % H2C2O4 acid. Almost
complete precipitation of Ln was attained at pH 1.1 and after stirring for
45 min . with precipitation efficiency of 99 %. The obtained Ln-oxalate
cake of high Ca content was re-dissolved in 5 % HCl and then reprecipitated
with NH4OH solution at pH 9.6 to obtain pure Ln(OH)3.
However, Y from the prepared Ln-cake was separated individually via
selective dissolution by using (NH4)2CO3. Almost Y dissolution (95.1%)
was achieved by using 200 g/L (NH4)2CO3 at solid/liquid ratio of 1/5
with stirring for 40 min at 70 oC. Also, Ce (IV) was then separated from
the Ln-cake almost free from Y by dissolving the other Ln in 5 % HCl at
the optimum pH value of 3.5 leaving Ce (IV) behind as un-dissolved
material.
With respect to B and V recovery, the prepared alkaline leach
liquor assaying 0.116 g/L of B, 0.18 g/L of V, 10.2 g/L of Si and pH >12
was subjected to Amberlite IRA743 anion exchange resin in its hydroxide
form to extract both of B and V after adjusting the pH value of the
prepared alkali leach liquor via conc. H2SO4 to 8.5. The latter was found
to be suitable for extraction of the metals of interest from one hand and to
remove the associated SiO2 gel, from the other hand. This realized about
87.5 % loading efficiency of B and 74.4 % efficiency of V at flow rate of
2.5 mL/min. Also, during elution process separation between co-loaded
elements takes place where B was first eluate by using hot water at flow
rate of 1mL/min realizing elution efficiency of 94 % . This followed by
elution of V by using 4 % H2SO4 at flow rate of 1.5 mL/min with elution
efficiency of about 97 %. After that, B was precipitated from the eluate
solution through gently thermal evaporation to obtain boric acid (H3BO3).
On the other hand, The eluant sulfate solution rich in V was subjected to
V oxidation to its pentavalent state via the addition of KClO3. The red
cake V2O5 was then precipitated by partial neutralization of solution to a
pH value of about 2.5 with ammonia solution and heating to 75 oC
for 2 h.
Finally, a technological flow sheet for processing of Abu Hamata
sandy claystone ore material was designed for the recovery of U3O8,
Y2O3,CeO2, H3BO3 and V2O5.