CA1247372A - Pressure leaching of uranium ores in a chloride medium - Google Patents

Abstract

TITLE
PRESSURE LEACHING OF URANIUM ORES IN A CHLORIDE MEDIUM

INVENTORS
George P. Demopoulos Gordon M. Ritcey ABSTRACT OF THE DISCLOSURE
Concurrent dissolution of both uranium and radionuclides from ores thereof has been achieved by leaching under oxygen pressure with an alkaline earth metal chloride preferrably calcium chloride solution acidified with HCl. Sulphide ores lead to sulphate ion build-up, excess of which is removed by autogenous alkaline earth metal sulphate precipitation: this has been found to allow increased solubility and extraction of radlonuclides particularly radium. Leaching times are short e.g. 4 hrs, such that radium precipitation or adsorption after dissolution is minimized. Uranium recoveries of above 95% and radium recoveries above about 90% have been obtained. The extracted radionuclides can be disposed of separately, leaving the tailings environmentally inert and suitable for backfill.

Description

~L2~737~

, FIELD OF THE I~VENTION
Minerals containing uranium and other radionuclides are leached wlth acidified calcium or other alkaline earth metal chloride solllt,ion under pressure to extract uranium and other radionuclides with substantially complete extractions of uranium and radium being obtainable leaving relatively innocuous tailings for disposal.
BACKGROUND AND PRIOR ART
Present methods of extractlng uranium from its ores involve leaching with sulphuric acid or alkaline (carbonate) solutions, depending upon the ore mineralogy (1). In either process only uranium is recovered from the ore, while the rest of the radionuclides including radium-226 reports to the tailings (2). Weathering and/or bacterially promoted oxidation of sulphides contained in these tailings (i.e. at Elliot Lake, Ontario) result in environmentally unacceptable high levels of radium releases into neighbouring streams (3).
The prevention of these radium releases during the operating life of the mine-mill complex i9 effected in mo~st cases by adding limestone to the tailings slurries to counter bacterial acid generation and by adding barium chloride solution to liquid effluents to coprecipitate radium as radium barium sulphate. However, the long term containment of these radionuclides remains a serious and as yet unsolved problem. Many investigations aimed at alleviating the adverse environmental impact of these uranium mill tailings have been carried out. The principal objective of these investigations was to remove radium from the tailings (4,5).
In addition to the problem of tailings stability, sulphuric acid leaching (e.g. as practiced in Elliot Lake, Ontario), is associated with slow kinetics (>40 h leaching time) and losses of potentially use~ul byproducts such as thorium and rare earths. Thus an alternative leaching system should be developed to meet the requirements of an environmentally acceptable process with simultaneous maximization of uranium recovery, possible recovery of byproducts and a ~impli~ied flowsheet. A simple ore treatment process that would solubilize all or most of the radium together with uranium and possibly thorium and rare earths is believed most desirable (6,7).

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Both nitric and hydrochloric acid systems have been tested as alternatives to sulphuric acid for leaching with the latter found to be the most promising one (8), ln addition to ambient pressure leaching systems, leaching under oxygen pressure could be an alternative system. In fact, this approach has been proposed for uranium extraction from sulphidic type ores ~ound in Canada (9), ~ustralia (10) and South Africa (11) as long ago as the mid 1950's. Further studies conducted on various ores from the Elliot Lake area at the Can. Govt, Mines Branch (now Canmet) (12,13), confirmed the effectiveness of the oxidation pressure leaching process.
Gow and Ritcey (14) and Dasher (15) have critically evaluated its applicability in the uranium industry and have indicated equivalent or substantially reduced capital and operating costs in comparison with the conventional practice. Among the advantages of pressure leaching are:
low or no H2S04 addition; fast kinetics and high uranium recoveries.
However all prevlous reports in the literature are concerned only with uranium recovery without any attention being given to the radium problem outlined above and the recovery of byproducts.
For the very high-grade uranium/nickel/arsenic ores of Northern Saskatchewan pressure leaching proved the preferred treatment process.
Extensive test work showed that thls method gave consistently high yields of uranium into the leach solution even when there was considerable variation of other major elements (e.g., arsenic and nickel). Thus the Key Lake Sask, uranium process was developed based on a mild pressure leaching in sulphuric acid in the second stage of a two-stage system (20,21). Leaching in autoclaves using sulphuric acid and oxygen was also considered for the treatment of the Midwest uranium deposit which is remarkably similar to the Key Lake deposits (22, 23).
While in conventional atmospheric leaching sulphuric acid is added to satisfy the acid requirements of the process, pressure leaching of sulphidic uranium ores makes little or no use of external sulphuric acid. The reason is the in situ generation of sulphuric acid during pressure leaching of the uranium ore. This exothermic reaction represents the oxidation of pyrite to produce the soluble iron and sulphuric acid necessary for the dissolution of uranium. The extent to 37~
~3-which this reaction would progress defines in a large measure the amount of oxygen consumed during pressure leaching and hence the economics of the process since oxygen is the most costly reagent used.
In view therefore of the increasing role played by acid pressure leaching in the uranium extraction Lndustry both in Canada and elsewhere and the urgent need for environmentally acceptable processes producing radium-~ree tailings, we proceeded to evaluate the possibility of simultaneously obtaining uranium and high radium dissolution during pressure leaching (evidently no other work has been conducted previously in this direction).
Experimental evidence from several investigations reported in the literature (2) and theoretical considerations (radium as a daughter of the radioactive decay of uranium is incorporated into the mineral structure of latter) lead to the conclusion that radium should be dissolved along with uranium during leaching. However, due to the very low solubilities of its sulphate salt (H2S04 acid leaching) and its carbonate salt (alkaline leaching) precipitation occurs resulting in most of radium reporting in the leach residue. Coprecipitation, adsorption or ion exchange phenomena further contribute to removal of radium from leach solutions.
We have found that through controlling a combination of experimental variables (temperature, oxygen pressùre, pulp density, retention time, catalysts) optimum oxidation conditions were established for the production of sufficient H2S04 and Fe3+ necessary for high uranium extraction with simultaneously minimal excess of sulphate ions thereby allowing high radium mobility. Subsequently we investigated the addition of various salt chlorides with the objective o~ maximizing radium extraction. Extraction levels of thorium and rare earths were also monitored.
SUMMARY OF THE INVENTION
According to the invention, a process has been developed for dissolving uranium and radionuclides from their ores comprising the steps of (a) leaching the comminuted ore with an aqueous metal chloride solution containing HCl while under oxygen pressure and at elevated temperatures above ambient, said metal being selected from the alkaline earth met,als calcium, strontium and barium and mixtures thereof with 7~'7~

ferric iron and cupric copper, the pressure, témperature and chloride concentrations being sufficient to obtain the desired co-extraction of uranium and radionuclides, (b) separatlng the loaded leach solution from the leach residue, and (c) recovering at least the uranium values from the pregnant leach solution.
Preferably the process is carried out on crushed and ground sulphidic uranlum ore which is leached for about 3-5 hrs with an aqueou~
solution containing calcium chloride and hydrogen chloride with amounts of approx. 500-1400 kg CaC12 2 H20 and approx 4-90 kg HCl being used per ton of ore at a temperature of about 50 - 150C under an oxygen pressure of about 100 - 500 kPa, to dissolve at least 95% of the uranium and at least 90~ of the radium present.
When the sulphate concentrations exceeds about 0.3 g/Q,gypsum is precipitated during the leaching thereby maintaining the sulphate ion concentration at levels insufficient to inhibit significantly the dissolution of radium, thereby obtaining high radium extraction.
A second stage leach in CaC12-HC1 solution (usually at atmospheric pressure) may be desirable in some cases to increase the recovery of radionuclides.
The resulting pregnant leach solution is processed by known techniques for stagewise removal of uranium, nickel and thorium if present in economic amounts, and radium.
DESCRIPTION OF DRAWINCS
Figure 1 i8 a preferred flowsheet of the process of the in~ention applied to conventional low grade uranium ores.
Figure 2 is a preferred flowsheet of the process of the invention applied to complex uranium ores.
Figure 3 is a graph showing the % extraction of uranium(from low grade ores)and radium variation with extraction time for four CaC12/HCl mixtures.
Figure 4 i~ a graph showing the % extraction of uranium and radium from complex ores with increasing extraction time at two temperatures 50C and 70C.
Figure 5 is a graph showing the % extraction of uranium and radium from complex ores with increasing extraction time at two partial oxygen pressures 172 and 517 kPa 2 Figure 6 is a graph showing the effect of various HCl additions on radium extraction from complex ores.
Figure 7 i8 a eraph showing the effect of various ~`7~

CaC12-2H20 levels on radium extraction from complex ores.
In the flowsheets, CCD is counter-current decantation.
DETAILED DESCRIPTION AND PREFERRED EMBODIMENTS
Both conventional uranium ores and more complex uranium ores (contalning Ni, As etc) can be leached according to this invention. A
conventional ore such as that from Elliot Lake Ontario contains about 0.1~ wt U, about 3% S about 3~ Fe and other radioactive elements including radium being about 0.05~. Higher Brade uranium ores containing nickel and arsenic and lower in iron and sulphur such as the Key Lake Sask. ore, can also be treated. The ore i9 comminuted to a suitable particle size for leaching, usually to about minus 50 mesh and preferably wlthin the range from about 30 to about 80% wt minus 200 mesh.
The leach medium comprises an acidified aqueous alkaline earth metal chloride solution. The alkaline earth metal may be selected from calcium, strontium or barium with calcium being preferred. These metals are selected to precipitate excess sulphate from solution. Strontium and barium tend to draw radium into the sulphate precipitate more than does calcium: thus the former are not preferred when high radium dissoluton is desired. Some ferric or cupric chlorides optionally may be present to enhance dissolution. The concentration of CaC12 may range from about 0.5 Molar up to saturation (i.e. about 7M at 25C) preferably about 1 to 5 M. This solution is acidified with HCl to a concentration of at least about 0.01 M preferably to within the range of about 0.03 - 1.0 M HCl.
The amounts of CaC12 and HCl found preferable per ton of ore leached are: at least about 500 and preferably from about 1000 to 1400 kg CaC12-2H20 per ton; and at least about 4, most preferably from about 60 to 90 kg HCl per ton.
The leaching can be carried out in any apparatus suitable for pressure leaching such as autoclaves and other pressure vessels. Due to the chloride leachant the vessels most suitably are made of a corrosion resistant material such as titanium or lined with a material such as polytetrafluoroethylene (Teflon - trademark) or glass or other corrosion-resistant coating.
The pulp density or solids content in the slurry during leaching usually should be within the range of about 10 to about 60% wt solids, preferably from about 25 to 40%. The solids content a~ects 3'~

the leaching efficiency: at hieher solids contents the leaching, especially of the radium is less effective and more leaching stages usually are necessary.
The temperature during leaching may range from ambient to about 150C depending upon the type of ores.
The pressure of oxygen applied usually will be within the range from about 130 to about 600 kPa, preferably from about 150 to about 500 kPa. Oxygen itself or oxygen-enriched atmospheres may be used (e.g, atmospheres of at least 40% wt. 2) Leaching contact ~ime normally need not exceed about 6 hours, with about 3 to 5 hours being preferable.
In most cases the ore will contain some sulphur (usually as pyrite or pyrihotite ) at least part of which will be oxidized and solubilized as sulphate ions. When the sulphate ion concentration builds up there will be an autogenous precipitation of calcium sulphate due to the presence of the calcium ions. This precipitation serves to maintain the sulphate ion concentration at a level which does not favor dissolved radionuclides precipitation. Radium in particular readily precipitates after being leached if the leach time exceeds about 6 hours and/or if the sulphate ion concentration becomes excessive. It has been found that the calcium sulphate autogenous precipitation maintains the sulphate level below that needed for radium precipitation.
An optional second stage leaching of the solids at either elevated or atmospheric pressures can be carried out using the CaCl2-HCl medium to increase extraction particularly of radium. For this second stage leach, pressures may range from atmospheric (No 2 overpressure) up to about 600 kPa oxygen pressure, temperatures from about ambient to 150C, and pulp densities from about 20 to 60 ~ wt.
solids.
Test results suggest that significant radium removal should be achieved during the first stage pressure leach if very low residual radium concentrations are to be realized durlng a second stage atmospheric leach. Radium concentration in the first stage leach residue preferably should be in the range of 100-120pCi/g in order to obtaln a final residue (i.e. after second stage atmospheric leaching) containing Ra 40 pCi/g solids.

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Selected mixtures of CaC12/HCl (preferably 1-2 M/0.037-0.1M) were found the most effective in achieving high radium extraction from low grade ores, this being due to sulphate removal from the leach solution by GaS04 precipitation. The CaC12 to solids ratio was found to control the final radium level in the pressure leach residue. Thus one stage pressure leaching at 125C, and 345 kPa 2 with 25% solids and 2 M CaC12/0.1 M HCl addition resulted in a very low residual radium concentration (36 pCi/g). Tailings can then be disposed of with minimiæed environmental impact. The requirement for operating at lower pulp densities than conventionally, is not expected to affect the production capacity of a given plant employing the present process since much shorter retention times are used. Operation at conventional pulp densities (50-60% solids) during pressure leaching is also possible but in this case a second leach stage at 70C under atmospheric pressure with CaC12/HCl as leaching medium should follow in order to render tailings almost free of radionuclides (-40 pCi/g Ra226). This two-stage scheme would allow full utilizaton of the working capacity of the pressure reactors: thus savings in capital expenditures may be anticipated.
After leaching, the loaded leach solution is separated from the residual solids by any suitable means such as decantation, filtration, centrifugation or cyclones or any combination thereof. Desirably the residual solids are washed and the wash ]iquid combined with the leach solution. A counter-current decantation is very suitable.
The second stage leach solution (and wash soluton) is combined with that from the first stage and processed for recovery of dissolved metals by known techniques.
Uranium can be removed from the pregnant leach solution very conveniently by solvent extraction e.g. using amine reagents or trioctyl phosphine oxide (TOPO) for example.
Ion exchange, adsorption or precipitation steps may also be used in the removal. For instance for the stagewise removal of U, Th and radium one suitable procedure would be to extract the uranium by solvent extraction followed by thorium removal by solvent extraction. The radium could be isolated by adsorption on an ion exchange resin.

~7372 If nickel and lead (Pb2'0) are present they also would be dissolved to a large extent and may be recovered by using solvent extraction for nickel and ion exchange isolation of the Pb2l0.
Rare earths if present tend to remain with the solid residue (separate recovery steps on the residue may be carried out if desired).
For example in order to recover rare earths and yttrium from the solid residue the following steps can be carried out:
a) leaching of the residue in hot H2S04, and b) stagewise solvent extraction separation of the rare earths.
The tailings will have a uranium content below about 0.05% wt, a Th content below about 0.02%, a radium content below about 70 pCi/g, and a very low sulphur content. These tailings can be disposed of without the need for neutralization and without significant effect on the environment e.g. as mine backfill or on the surface with eventual revegetation.
The chloride leach medium can be recycled thus reducing leaching reagent requirements and reducing liquid effluent discharge requiring neutralization. Some "make-up" CaClz and HCl will be required to regenerate the leach solution.
The following examples are illustrative.
Example 1 A uranium ore was obtained from the Elliot Lake area (Quirke mine9 Rio Algom Ltd) and consisted of a pyritic quartz pebble conglomerate. Typical chemical analysis of the ore used in this work is shown in Table 1. The ore samples were ground to 63% minus 200 mesh.
Brannerite is the ma30r uranium bearing mineral, but uraninite and monazite are also present in the oreO Most oI the thorium and rare earth content~ are a~sociated with monazite which is a very refractory ore. Pyrite i~ the maln sulphide mineral. According to the chemical analysis shown in Table 1, the iron content is much higher than the amount expected lf it i9 a~sumed that all sulphide sulphur of the ore is associated with 7 3 ~ 2 pyrite only. Most likely therefore, some pyrrhotite is al~o present:
Iron oxides as well as iron introduced during grinding might al.so be the reason for the hlgher iron content.
Table 1. Chemical Analysis of Elliot Lake Uranium Ore :I~N.~IIU~NI WElGHT i CONSTIIUENT WEICWr ; .
_ U 0.11 SiO2 83.34 Ra 381 pCi/g Al 3.47 Th 0.037 Mg 0.14 _ _ ~

RE'~ (Total)* 0.15 Ca 0.19 Fe 3.37 K 1.88 S (Total)3.05 TiO2 0.46 S (S04) 0.26 r------ _ 0.024 3 S** 0-01 C (Total) 0.07 * RE = rare earth~
** S = elemental sulphur ~9L'7~7~
--1 o--Leaching tests were conducted in a 2-liter titanium autoclare supplied by Parr Instrumenk Company, Moline, Illinois. An aqueous slurry of the feed was heated to the required temperature under slight nitrogen overpressure (i.e. air was excluded). Chemical addition~ ~if applicable) were introduced at reaction temperature through a specifically adapted pressur~zed chamber. Oxygen was also admltted at reaction temperature (time zero) under the slurry surface to ensure good gas dispersion. The oxygen pressure reported here is the difference between the total pressure measured after oxygen introduction and that before. The autoclave was connected throughout the course of a leaching test to the oxygen cylinder thereby allowing contlnuou.s supply of Oxygen at the pre-set pressure level. Agitation was provided by two 6-bladed axial turbine impellers~at a constant rate of 450 rpm. Temperature control was maintained at + 2C throughout the test except at the initial stage of the reaction when the introduction of the cool mass of oxygen gas caused a temperature drop of 5-10C. The temperature drop was quickly compensated by the heat released from the exothermic oxidation of pyrite which resulted in equivalent temperature excess over the set point. In most of the tests stable temperature control was established after the Eirst half hour since oxygen introduction.
The progress of the leaching reactions was monitored by withdrawing slurry samples at appropriate intervals through a sampling port equipped with a ball valve. The samples were filtered immediately after collection and the residues were washed repeatedly: twice with acidified distilled water (0.025 vol% H2S04 or 1.0 vol% HCl the latter used for those tests where chloride additives were involved) and finally once with distilled water at room temperature. All the washings were conducted with approximately the same solid to liquid ratio like the one used for each test. Repulping of some residues with more acidified water showed no significant change in uranium and radium contents.
Pregnant, combined wash solutions, and dried (at 90C for at least 24 h) re~idues were sent for appropriate analysis.
Leaching efficiencies were determined based on the dried solids assays. Correction factor3 were applied to account for the solids weight losses during leaching (3~ after 6 h leach). Solution concentrations for uranium were 31ightly higher than those expected from 37~

mass balance calculations due to evaporation losses during sampling a3 well as due to some solution retained in residue arter filtration. In contrast solution concentrations for radium and sulphate ion were systematically lower, the former believed to be due to precipitation of RaS04 by the time analysis was conducted and the latter due to gross interferences Or the complex solution composition with the applled barium sulphate method.
Extraction curves for uranium using 1.0 M CaCl2 solutions acidified with various amounts of HCl are shown in Figure 3. The uranium picture is clear; increased uranium extractions were obtained with increased acid additions. Radium extractions also increased initially with increasing acid concentrations but this beneficial effect was almost cancelled out with longer leach times (6 h). Best results ~95% U
and 69% Ra) were obtained at 4 h with 1.0 M CaCl2/0.44 M HCl.
Significantly more iron was found to report in solution in the chloride system than in the case of the sulphate system.
The result3 from some additional tests conducted at different conditions are shown in Table 2. Comparison of runs No. 18 and 14 reveals that using a relatively low pulp density (25% solids) results in significantly improved radium extraction (90~). Also, comparison o~ runs No. 17 and 24 as well as No. 18 and 25 show that lower CaCl2 concentration results in lower radium extraction. Thus the results of these tests indicate that a high ratio of CaCl2 concentration to solids concentration is required for achieving high radium extraction. The effectiveness of CaCl2/HCl mixtures in enhancing radium extraction during pressure leaching of uranium ores is further illustrated with runs No. 19 and 19A of Table 2.

3o 4737~

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Second-Stage Leach -Best results from the one-stage leaches of Example 1 were obtained by using 25~ solids slurries combining also at least 1.0 M
CaC12/0.037 M HCl at 125C and 345 kPa oxygen pressure. Thus uranium extraction was 295% while the final leach residues (6 h leach tlme) contained -40 pCi/g Ra226 i.e., 90% radium extraction into solution was achieved. To lower even further the residual radium content of the uranium tailin~s as well as to allow higher pulp densities to be employed, a two stage leaching scheme was explored.
Samples of residues from pressure leaches at high pulp densities (50-60%) with the addition of CaC12/HCl mixtures of Example 1 were treated separately either by employing pressure leaching at 125'C
with 103 kPa oxygen overpressure or by atmospheric leaching at 70C.
Both leaches were conducted with 0.1 M HCl in the presence of CaC12 at reduced pulp densities (25-33% solids). Atmospheric leaching proved more ei'ficient ror radium removal apparently due to a lesser tendency for sulphate production than in the case of the pressure leach. The second stage atmospheric leaching results along w1th the leaching conditions employed and the composition of the first stage residues used are summarized in Table 3.

3o 3~

Table 3 Second Stage Atmospheric Leaching Tests Conducted Using Residues ~rom first Stage Pressure Leaches SECOND STAGE LEACH - ATMOSPHERIC Overall Extraction Feed' RESIDUE (Both Stages) _ LEACHING CONDITIONS _ _ U Ra U Ra U Ra ~_ pci/g ~ pci/g %

0.012 334 70C; 25% solids; 6h 0.008 109 93 72 l.OM CaCl2/0.1 M HCl 15 _ _ _ _ 0.009 298 70C; 25% solids, 6h 0.006 98 94.5 75 _ 1.0 M CaCl2/0.1 M HCl ~ _ _ 0.006 112 70C; 25% solid~, 6h 0.006 41 94.5 89.5 . 2.0 M CaCl2/0.1 M HCl . .
.
1 Re~idues ~rom First Stage Pressure Leach 3o 37~

Ex ple 3 Midwest U ore Table 4 ~hows a summary of a number of experimental tests on the complex ore at various conditions of pulp density, HCl addition, CaCl2 add~tion, temperature, pressure and retentLon time. Uranium recovery of >99g was achieved in a number of the tests at various condltions.
Solubility of >99~ Ra226 was achieved after a 6-hour leach at 40~
solids, 82.5 kg HCl/ton, 796 kg CaCl2~2H20/ton, 70C, 75 psi 2- The resldue contained 60 pCi/g Ra226. These conditions also resulted in the highest nickel extraction of 81~.
Example 4 In a series of leaching tests on the complex ores, the temperature of the leach was varied at 50C and 70C. The slurry density was 25~ solids. To each test was added HCl equivalent to 55 Kg/ton~and 1333 Kg/ton CaCl2-2H20. The leach tests were at 517 kPa 2- Increased temperature has a definite positive effect on the solubility of the uranium and radium in the leach liquor, and 4 hours resulted in maximum recovery, The results are summarized in Fig. 4.
Example 5 In another series of leaching tests conducted on the complex ore, the partial oxygen pressure was varied at 172 and 517 kPa 2- The slurry density was 25~ solids, the temperature was 70C, and additions of 55 Kg/tonne HCl and 1333Kg~ton CaCl2-2H20 were made. The data show no improvement in solubility of U and Ra by increasing the pressure and 4 hours leach time was optimum. (see Fig. 5).
Example 6 -In another series of tests on the complex ore the amount of HCl ~as varied ln the range 40-110 Kg/ton. The leach at 25% solids was at 4 hours at 517 kPa 2 at 70C with the addition of 800 kg/ton CaCl2-2H20.
The results indicated about 80 kg HCl/ton was necessary for maximum radium extraction. The final concentration in the residue was 100 pCi/g Ra225. Th results are shown in Fig. 6.

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Example 7 In another series of tests on the complex ore the amount of CaCl2-2H20 addltion was varied between O and 1400 kg/ton. The results, after a 4 hour leach at 70C at 517 kPa 0~ with 55 kg/ton HCl, indicated nearly 1400 k~ton CaCli-~H20 wa~ necQs~ary for maximum radl~m extraction. The radium in the residue analysed 220 pCi/g Ra22fi. The results are shown in Fig. 7.
Referring to Figure 1, this flowsheet summarizes a preferred process for low grade ores such as the Elllot Lake ore, The ore is crushed and ground and then leached ln concentrated CaCl2 solution containing 4 ke HCl/tonne ore at 125-150C for 4-6 hrs under an 2 overpressure of 15 - 50 psi. After the main solid/liquid separation, suitably by decantation, the solids are washed (suitably by counter-current decantation CCD) and the wash liquid added to the leach liquor.
Radium is removed from the liquor (suitably by ion exchange resins) either before or after uranium and thorium recovery giving a radium concentrate for burial or other disposal. Uranium is then recovered from the liquor (suitably by a tertiary amine or trioctyl Qhosphine oxide) followed if desired by thorium recovery (e.g. by solvent extraction using a primary or secondary amine.) Excess water is evaporated from the residual CaCl2 medium and the medium recycled to the leach stage. The ~olld tailings have less than 0.005~ U, 40 pCl/g Ra, 0.017~Th and about 1.7% S and about 2~ Fe.
Referring to Figure 2, this flowsheet summarizes a preferred process for complex U ores such as found in Saskatchewan. The crushing-grinding, pressure leaching, solid/liquid separation and CCD
wash are similar to Figure 1. The combined leach and wash liquors are sol~ent extracted, suitably with amlnes or TOPO to extact the U into the organlc phase, from which the U is strlpped with H2SO4 ~ H202 mixture, precipitated from the strip solution wlth H202, flltered and recovered as a uranlum peroxlde product. The leach liquor raffinate is treated for Ra removal similarly to Figure 1, and usually followed by Ni recovery, suitably by precipitation and subsequent purification by solvent extractlon and/or hydrogen reduction. This recovered nickel can be readily marketed to improve the economics of this process. The residual CaCl2 medlum ls recycled as in Figure 1.

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REFERENCES
1) Technical Reports Series No. 196; International Atomic Energy Agency: Vienna; "Signlficance of mineralogy ln the development of flowsheets for processing uranium ores"; 23-26; 1980.

10 2) Itzkovitch, I.J. and Ritcey, G.M. "Removal of radionuclides from process streams - a review"; CANMET Report 79-21, CANMET, Energy, Mines and Resources Canada; 1979.

15 3) Rltcey, G.M. and Silver, M. "Lysimeter investigations on uranium tailings at CANMET"; Can Min Metall Bull 75:846:134; 1982.

4) Proc OECD/NEA Sem on Management, Stabilization and Environmental Impact of Uranium Mill Tailings: Albuquerque, New Mexico; July 1978.

5) Proc Int Symp on Managing Waste from Uranium Mining and Milling:
Albuquerque, New Mexico; May 1982.

6) Ritcey, G.M. "Treatment of radioactive ores at CANMET"; Division Report ERP/MSL 77-139(0P); CANMET, Energy, Mines & Resources Canada;

3o 7) Phillips, C.R. and Poon, Y.C. "Status and future possibLlities for recovery of uranium, thorium, and rare earths from Canadian ores, with emphasis on the problem of radium", Miner Sci Eng 12:2:53; 1980.

i7~

8) Haque, K.E., Lucas, B.H. and Ritcey, G.M. "Hydrochloric acid leaching of Elliot Lake uranium ore"; Can Min Metall Bull 73:819:141;
, 1980.

9) Forward, F.A. and Halpern, J. "Acid pressure leaching of uranium ores"; J Met 463; March 1955.

10) Gray, P.M.J. 7'The extraction of uranium from a pyritic ore by acid pressure leaching"; Trans Inst Min Metall 65:55; 1955/1956.
.

11) Robinson, R.E., James, B.S., vanZyl, P.C.N., Marsden, D.D. and Bosman, D.J. "Preliminary a3sessment of the important econo~ic factors in the pressure leaching of low-grade South African uranium ore3"; Proc. of the 2nd Int Conf on the Peaceful Uses of Atomic _.
Energy; Geneva, tJnited Nations; 3:318; 1958.

12) Harrison, V.F. and Gow, W.A. "Air oxidation acid pressure leach investigations of uranium bearing ores from Elliot Lake, Ontario";
Tech Bull 3; Mines Branch; Energy, Mines and Resources Canada; 1959.
13) Vezina, J.A. and Cow, W.A. "Some design aspects of the pressure oxidation acid leaching of a Canadian uranium ore"; Tech Bull 110;
Mines Branch; Energy, Mines and Resources Canada; 1969.

14) Gow, W.A. and Ritcey, G.M. "The treatment of Canadian uranium ores -a review"; Can Min Metall Bull 62:1330; 1969.

15) Dasher, J. "A conceptual design for the leaching of sulphidic uranium ores without reagents"; Min Sci Eng 4:28; 1972.
16) Weir, D.R. and Masters, I.M. "Key Lake project process development -Part 1 - uranium extraction"; Paper presented at the 19th Ann Con Metallurgists, CIM; Halii`ax, N.S.; Aug. 1980.

17) Neven, M. and Gormely, L. "Design of the Key Lake leaching process";
Uranium'82, 12th Ann Hydrometallurgical Meeting, CIM; Toronto, Ontario, Aug. 29 - Sept. 2, 1982.

3~;2 18) Melis, L.A., Fraser, K. and Fisher, J.W. "Midwe~t project - Canada Wide Miné~ Limited"; Proc of the Meeting of the Canadian Uranium Producers' Metallurgical Committee; Sa3katoon; May 19-20, 1981, G.M.
Ritcey, ed.; Division Report ERP/MSL 82-42(TR); CANMET, Energy, Mine3 and Resource3 Canada; 1982.

19) Melis, L.A., Fra~er, K. and Lakshmanan, V.I. "The Midwe3t uranium project - development of the milling proce~"; Uran_um '82, ibid ref 17.

3o

Claims (17)

Claims:

1. A process for dissolving uranium and radionuclides from comminuted ores containing same comprising:
(a) leaching the comminuted ore with an aqueous metal chloride solution containing HCl while under oxygen pressure and at elevated temperatures above ambient, said metal being selected from the alkaline earth metals calcium, strontium and barium and mixtures thereof with at least one of ferric iron and cupric copper, the pressure, temperature and chloride concentrations being sufficient to obtain the desired co-extraction of uranium and radionuclides, (b) separating the loaded leach solution from the leach residue, and (c) recovering at least the uranium values from the pregnant leach solution.

2. The process of claim 1 wherein the contact time during leaching is not longer than about 6 hours.

3. The process of claim 1 wherein the metal chloride is calcium chloride in concentrations of about 1 to 5 M in the leach medium.

4. The process of claim 1, 2 or 3 wherein at least about 500 kg of CaCl2?2H2O is used per ton of ore.

5. The process of claim 1, 2 or 3 wherein the HCl concentration is about 0.03 to 1.0 M in the leach medium.

6. The process of claims 1, 2 or 3 wherein about 4 to 110 kg HCl is used per ton of ore.

7. The process of claims 1, 2, or 3 wherein the oxygen overpressure is from about 130 to about 600 kPa.

Claims (continued

8. The process of claims 1, 2 or 3 wherein the temperature during the leaching is within the range of about 40°C - 150°C.

9. The process of claims 1, 2, or 3 wherein the solids content of the slurry during leaching is within the range of about 10 - 60 %
solids.

10. The process of claim 2 wherein the leaching time is about 4 hours.

11. The process of claims 1, 2 or 3 wherein a second stage leaching in CaCl2-HCl solution is carried out to increase the radionuclide recovery.

12. The process of claims 1, 2, or 3 wherein the leach solution, after removal of the leached metal values therefrom, is recycled.

13. The process of claim 1 wherein crushed and ground complex uranium ore containing nickel and arsenic is leached for about 3-5 hours with an aqueous solution containing calcium chloride and hydrogen chloride with amounts of approximately 1000-1400 kg CaCl2?2H2O and approximately 70-90 kg HCl being used per ton of ore at a temperature of about 70 - 100°C under an oxygen pressure of about 200-500 kPa, to dissolve at least 95% of the uranium and at least 90% of the radium present.

14. The process of claim 13 wherein gypsum is precipitated during the leaching thereby maintaining the sulphate ion concentration at levels insufficient to inhibit significantly the dissolution of radium, thereby obtaining high radium extraction.

15. The process of claims 1,2 or 3 wherein uranium is recovered from the pregnant leach solution by solvent extraction.

16. The process of claims 1, 2 or 3 wherein radium is recovered from the pregnant leach solution by adsorption on an ion exchange resin or by precipitation.

Claims (continued)

17. The process of claims 1, 2 or 3 wherein the chloride leach medium includes ferric or cupric chloride.