US2873165A

US2873165A - Uranium recovery process

Info

Publication number: US2873165A
Application number: US164566A
Authority: US
Inventors: Richard H Bailes; Ray S Long; Robert S Olson; Herbert O Kerlinger
Original assignee: Individual
Current assignee: Individual
Priority date: 1950-05-26
Filing date: 1950-05-26
Publication date: 1959-02-10
Anticipated expiration: 1976-02-10

Description

Feb. 10, 1959 R. H. BAlLEs ETAL 2,873,165

URANIUM RECOVERY PROCESS Filed May 26, 195o 3 sheets-sheet 1 Petz. 11o, 1959 R. H. BAILES ETAL URANIUM RECOVERY PROCESS Filed May 26, 1950 PER GENT URANIUM RECOVERY IOO 3 Sheets-Sheet 2 A O L53 EQU|VALENTS F/l. A L75 EQUIVALEN-l- S F/l.

CATION EQUIVALENTS FL`UORlDE EQUIVALENTS NVENTORS. RICHARD H. BA/LES ROBERT S. OLSON HERBERT 0. KERL/NGE? By RAY 5. LONG TTORNEK feb. 10, 1959 R. H. BAILEs rs1-A1.

URANIUM RECOVERY PROCESS Filed may 2e, 195o R E .3 BNG OMUN SIG .t TALLN.. NBOHO .n \l, E .EL S .m WHSK. Q- m 4 .t RRTY e O AE A e MRM E 3 m H Y m B O nlv O E 8 m T u o le N H250., V H2504 E D. C A

A T TORNEY.

F, IC

.- v Fei. 10, 1959 r as nal products evsiiitab-le for processing by the method` of the invention. l i `illus.traltively, such raffscilution-k may have the composi- `URANIUM RECOVERY PRocEss tion indicated in 'the .tellbwislftsbler Richard LNI. Bailes, Walnut Creek, Ray S. Longfvalleio, Egberts. Olsdm San ErancscoQnd HerbrtOfKerf linger, Berkeley, Calif., as'signorsx, `by"mesne assignments, toV `the .United-` States of America as represented by the United StatesAtomic Energy Commissioni Application May 26, 1950, Serial No. 164,566

14 crains. (c1. zel-14.5)

d This invention relates to processes for recovering uranium from solutions and, `more particularly, `relates to the recovery of uranium from phosphoric' 'acidlor other phosphatic solutions.

`The presence oflow concentrations of uranium in certain` phsphaticoreshas beenknown forlsome time but fthe direct recovery Aof uranium from this source has llQttben QC.Onlvlllsally` feasible dus t0 the 10W concentrationpresent, SinceA these ores are pres-V onlyv P Hlled inthe Preparation 0f Superphosphate and phpsphoric acid-l in `processes` wherein sulfuric acid` is etuplgyedtonuleaching the phosphatic materials `from the ore herginthe uraniumvis leached intothe re cfphosphaticnsolutions, there existsdfa posgctluvering. this low` grade` uram'lum from` the of uranium from th`e"phosphatic Vsolution complicated beth `bythe 19u concentration thereof and the high Cenu ccntration'y 'of phosphate in thel solution in `which'solution .the uranium is highly soluble due.- to complexing phenomena." Y l owfit lbeen iound thaty lowfconcentrationsof .isomer bsfesovslectfrem crude phosphoric acid er? other ias'fific phqsphatc Sfluies by reducing `.the uranium to the tetrai/alent state and precipitatingi .the as` awtdiuoride and, further, that 4the uranium' methods 1.1.1.0..` fully' described" hereinafter..

,Ascetdiesly it' is 211,1 Object ef the invention to provide a method for recovering uranium from phosphoric acid or` other phosphatic solutions.

AAnother object of the invention is to provide a fluoride CII However, solutions diteriug very considerably from that indicated i A relativedc nc'ent-rationsdorin the types of itms pre,Y y als be satisfactorily processed.

With refer ce t thetou hee't o f `Fig. Land in accordance 'invention yalpl'iosphafte or phosphoric acidV solution offfthe `cljra'rafct,er described is subjected to reduction, whereby all'of' the uranium contrzrtinedl therein vis reduced to the tetravalent state. The reduction of the uranium to the tetravalent state may be accomplished utilizing metallic reducing agents such asiron, zinc, zinc amalgam,` aluminum or lead, or byV using reducing agents such as hydrazine salts, hypophosphorous acid or hydroxylamine hydrochloride or'by electrolytic methods. The Yrequirements for the reducing'agent are no t critical and a majorityof'the'reducing agents having a suiciently high reducing potential appear to besuitable ,for the purpose. l.'l`-he..pro"gress` of the reduction' is easily followed by observing; the. electromotive potential developed between platinum and saturated calomel electrodesl immersed in the solution. Reduction is substantially connplif@v .when .the potential is lowered to about +O.1,.yolt.

l `Policy/ving reduction, the solution is treated to coprecipitate;the reduced Auranium as a fluoride together with other iIlSQlllble iuorides thereby achieving a subi stantially `complete removal of even very small quantipreipitation 'rne'thod for recovering uranium froripliosphoric acid or'other'phosphatic "solutionsl l "A furtherobject of'the'ivention is to provide uranium `ieovery processes vwherein a fluoride preciisita'te ontaining' uranium is deuor'inated andthe uranium isA- recovered fromthe defluorinated precipitate.

stillturther 4object of. the invention is to provide uranium .recovery processes wherein uranium is 'leached from. a.uoridefprecipitatefand the uranium is recovered from the `.leach solution. Y

kOther robjects and advantages of the invention will become .apparentffromlthe following description taken in conjunction -with the accompanying drawing `of which: Eigurefl is' a how sheet of the processes of the invention; i d d' fFig. 2is a graphical illustration of the ndependence of uraniumV recovery upon precipitant ratios and amounts; and

Fig. 3 is aigraphical illustration of the dependence of the leaching characteristics of deliuorinated ndcalcined uoride vprecipitates .upon time and temperaturefof icalcination. Relatively concentrated phosphoricacid or phosphatc solutions containing uranium in" low concentrations fs'uch as lare obtained in"v`arious.phosphatefertilizer and metallurgial y, ,progesses g. as byproducts; intermediate' :products ties of. uranium from the phosphatic solution. event :dthiat theI l tonis `d ei'itent in lcoprecipzitant materials, fa sulic .q offa solubley compound of the coprecipitant `isy supp Y to.fthe'solution and then a SolublF-fluoideger hydefl c acid .is added t0 .coprecipitats the reduced. ,grausam with. the 1Cor-recipient msterial. Calcium carbonate, calcium oxide, hydrated calciutn o aide, andbariuml carbonate have been found In the satisfactbryfftoE supply the required coprecipitant matef rials; however, other `materials@may also be used as will` become apparentffrom theff'llowingdiscussion. vVarious solubleuoriddes, such asf'sodiumhuoride, potassium fluoride, ,bariurntluoride and ammonium fluoride vand other fluorideswhich `fdurnislr'al Vconsiderable fluoride ion concentration tohefsolution, may be employed in the precipitation, also hydrouoric acid (hydrogen uoride) is completely satisfactory. "['he ratio of cation equivalents ,to ttuoride equivalents present in the acid and the total amount'of precipitants present determine to a critical` degree the amount of uranium which is recovered in the precipitate. `Fig. `2` graphically illustrates the close and critical relationship which these factors bear to the amount offuranium recoveredL` The mechanismof .the precipitation of uranium from d thephosphnjic,` acidrsolution is not completely understood. AIt has been noted that the presence of aluminum in the solution ygreatly increases the recovery of uranium. In the presence of; aluminum va cryolite type of compoundl precipitatesfwhich compound is preferentially precipitated rather than uranous i fluoride. The cryolite s compound which precipitates appears to have the general formula X (AIFBL, wherein X may represent various proportions and combinations of radicals such as uranous uranium, Mg, Ca, Na, K, NH4 and others. The existence of compounds such as `MgaUllF): and Ca3(AlF)2 in the precipitate has been indicated by X-ray diffraction studies. Moreover, a; mixture of variousV ,otherfinsoluble tluorides may be precipitated along with the insoluble uranium compound. Some of the iluoride originally present in the acid and thatadded during the precipitation may be recovered Vfrom the precipitate -andfrom the solution forn recyclingby methods more.fully..describedhereinafter.4

The uorideprecipitate Acontaining uranium is. separated from thegsolution, bylltering and the uranium is recovered' from the precipitate by one of several valternative methods and the ltrate is. treatedfor recovery of the uoride or/and is returned to .theplant for normal PfC6SS1Ug- A typical precipitate derived from such a solution using calcium carbonate and'hydrogenvuoride as the precipitants may have a composition determined by analysis approximately as follows: However, it will be api preciated that the composition 'of such a precipitate may vary considerably' from that indicated dependent upon the composition of the original acid 'and conditions -of precipitation. -v-

Percent U3O8 0.25 Phosphate 1.95 Sulfate v 0.85 Ca. A1 v e A 8.6 Mg 4.95 F 47.0

Emission spectrographic analysis discloses Y thatA other materials may be present as follows:

Percent Fe 0.5-1.0 Na '0.5-,1-10 Si "0.05-0Ql0" V 0.050.10 Cu f 01001-001 Cr l 0.001-001 Mn 0.00l-0.0l Ti 0.0014001 In general the methods of uranium recovery whichare employed may be classed in two groups; one, in which the uranium is recovered from the precipitate following defluorination of the precipitate; or, two, in which the uranium is recovered from the precipitate prior to defluorination. i

In the methods of the iirst classification, sulfuric acid is added to the precipitate in stoichiometric. excess relative to the quantity of fluoride present and hydrogen liuoride is vaporized from the precipitate and recovered in order that it may be recycled. `Alternatively, ammonium sulfate is added to the precipitate and ammo nium fluoride is sublimed'from the precipitate and recovered for recycling. Then the delluorinated precipitate is calcined and leached with water or aV solution such as dilute hydrochloric, nitric or sulfuric acid solutions and the titanium is finally recovered from the leach solution.

The temperature to which the precipitate is heated in driving oftl the iiuoride atects the conditions which must be employed during the leaching. If the'precipitate, following treatment with sulfuric acid, is heated to a temperature below about 800 C., for 1% to 2 hours the precipitate produces an acidic solution on contact with water since aluminum sulfate appears to be formed by this treatment which aluminum sulfate produces a sufliciently acid solution to extract the uranium. However, if the heating time is prolonged or the temperature is. increased to about SOO-1000" C., alkaline products prey 4 dominate in the heated precipitate and au acidic solution must be employed to leach the uranium. Moreover, the solubility of various components of the precipitate is markedlyv affected by the time and temperature of heat ing; Ilittle aluminum and/ or magnesium being leached from the more intensively heated materials.

The composition of a leach liquor typical of those which are derived by leaching a iluoride precipitate, which was heated to a temperature of 800 to 1000" C. for 1% to 3 hours with a slight excess vof sulfuric acid, employing 1% sulfuric acid is indicated below in Table 2., About 80-90% of the uranium, none of the aluminum, and only one-third of the magnesium was'leached. l I

' calcination and leach reagent are graphically indicated in e this type follows:

Fig. 3. The time-temperaturer curve (A) of'deuorination vare plotted with respectv to the leach'liquorpH (reading upwards) curve (B); percentage-ofuran1um leached,'curve (C); percentage of aluminum'leached, curve (D); and percentage of Mg leached, curve (E).

Optimum conditions for a sulfuric acidleach are as follows: y0.2 N or about 1% sulfuric acid as the leach reagentminus 40 mesh size of the precipitate; and a liquidto solid ratio of about 2.5: 1.

The uranium may berecovered from the leach liquor by two methods:

(l) Precipitation of uranyl phosphate by pH adjustment, yielding 99.9% uranium recoveries ofl a material which after ignition contains uraniumequivalent to about 10% U3O8, employing amounts of sodium hydroxide equivalent. to about 8 to 16 gramsv per gram of USOS precipitated.

(2) By reduction of the leach liquor with sodium hydrosultite, or other suitable reducing agents, 99+% of the uranium is precipitated in a material containing about 10 to 20% UaOg. An analysis of a typical precipitate of Percent 13.8 Phosphate 30.3 Sulfate 1 Ca n Y Mg In the methods of the second classication, hot sodium carbonate is employed to leach the uranium from the fluoride precipitate with great' etiiciency While ammonium carbonate may also be employed if somewhat less leaching eiiciency can be tolerated. The leaching eiciency of sodium carbonate solutions drops off very rapidly below a critical pH value between about 8.5.and '9.0.

o A decrease in eiiiciency isnoted with increasing solid to liquid ratios and a marked increase if the temperature at which the leaching is performed is iucreasedjtonear the boilingA point. ACarbonate leaches all oftheV sulfate; about 0.1% of the phosphate, 0.8% of the aluminum and 2% of the fluoride from the precipitate in addition to the'd uranium. Y

. sodiumcarbonate leach solution preferably by neutralizing the solution' with hydrochloric acid and rprecipitating the, uraniumy in ja purified' form with carbonate-free S arnrnonia as ammonium vdiuranate. If there is present a considerable concentration of phosphate in the solution, uranyl phosphate or a mixture of uranyl phosphate and ammonium diuranate is precipitated. Precipitated ammonium diuranate is then calcined to yield high grade U30. The leached vprecipitate may then be treated to recover the uorine or is discarded.

Alternatively, in the second classication of methods, acidic leaching-of the fluoride precipitate may be employed without washing the precipitate. Dilute aqueous solutions of hydrochloric, nitric, citric acid and less effectively tartaric, acetic and-boric acids are used to leach the uranium from the precipitate with the uranium concentration being enhanced by successively leaching several portions of precipitate with the same solution and uraniumv is recovered from the acid leach solution by precipitation methods which further increase the purity of the material.

Acid leaching of the precipitate has been found to be influenced by the following variables as exempliiied when dilute hydrochloric acid is employed; concentration of acid, time of leaching, temperature, solid to liquid ratio and phosphate concentration in the uoride precipitate. Optimum conditions for leaching more than 95% of the uranium from a dry fluoride precipitate require `a 30 to 35 minute leach with agitation at 95 C. with 5 to l0 ml. of 5% hydrochloric acid per gram of uoride precipitate. Lowering the leaching temperature to 25 C. decreases the uranium recovery to 75% and a live minute leach yields nly"30% uranium recovery. Leach times beyond 30 minutesshow'no improvement in uranium recovery. When acid concentrations above about are employed the loss of fluoride from the precipitate becomes large and lower concentrations decrease the uranium recovery. Markedly decreased uranium recovery is noted with lower solid to liquid ratios while higher liquid to solid ratios show only a'slight improvement.

`Reduction of the uranium inthe leach solution with variousreducing agents causesthe formation of a uranous kphosphate precipate with the phosphate which is leached into the hydrochloric `acid solution. Sodium hydrosulfte, added to the solution as a reductant, precipitates a material containing uranium equivalent to about 55 to`65% U3O8 and about 30% of phosphate.

Particular details of the processesA of the invention will become apparent from a consideration of the following illustrativeexamples of the operation of processes in accordance with the invention:

EXAMPLE IA 17:1 `liters of a` solution having the` composition indicated in Table I was reduced bycontact with metallic ironfto a potential of +0.08 V. as measured with a platinum-standardcalomel electrode system.

24.7 grams per 'liter of calcium carbonate and 25 grams per liter of aqueous hydrogen uoride added to the sointionV produced a .precipitate Weighing 694 grams,v sub statially free of vphosphate after washing, andcontaining 0.21% uranium and about 78% of the recoverable tluorine. 'lhisuranium recovery amounts to abo'ut94`% bfglfon the content of the precipitates and tiltrates,

. of 900 C.

esta, 1ers ing an excess of concentrated sulfuric acid and heating the mixture for 130 minutes to a maximum temperature was leached with 325 cc. of 0.2 N sulfuric acid at a temperature `of 90C. for 15 minutes with a resultant removal of 68% of the uranium bythe leach liquor. A

second Vleaching of the same material with 325 cc. of 0.4 N sulfuric acid for minutes at 95 C. raised the overall amount of uranium leached by the two successive operations to 98% of the uranium originally present in the material being treated.

EXAMPLE `B 20.7 liters of this acid was reduced by circulation through a metallic iron packed column until the electromotive potential of the solution as measured between platinum and standard calomel electrodes was -|-0.033 V. The reduced solution was then treated with 360 g. of calcium carbonate and 476 g. of uorine as an aqueous solution of HF to precipitate the uranium. After washing with water and drying at 120 C. for 24 hours the precipitate weighed 805 g. and had the composition indicated in Table Ill.

This percentage of uranium amounts to `a recovery of 76% of the uranium originally present in the acid and is slightly lower than that normally obtained due to incomplete reduction of the acid. Portions of this precipitate were deiluorinated and calcined by heating with` concentrated sulfuric acid using the conditions indicated in Table lV. Portions of these deuorinated materials were then leached employing the conditions indicated under the appropriate heading of Table IV. The water and 0.092 N sulfuric acid leach being done on the material heated for minutes at 800 C. and the 0.18 N sulfuric acid leach being done on the material heated for 17 0 minutes at 900 C.

Table IV DEFLUORINATION H1804 percent excess 6. 9 '5 Defluorination time, 100 170 Max. Temp. of dsuornatlon, "O 800 900 LEACH [50 g. of deuorinated material leached with 125 ce. of reagent.]

H1804 0. 092 N B leach grams of this Vdeuorinated material;

Uranium contained in these leach liquors was recovered by treating 25ml. portions with 0.5 g. of -sodium hydro sulte resulting in the precipitation of 8l and 99.9%,V respectively, of the uranium from leachliquors A and `C asa precipitate containing 13.4 and 19.1% of U.

' Also, the uranium was recovered as uranyl 4phosphate by adding alkali to the solutions. 25 ml. of leach solution B was treated with 26ml. of 0.1 N NaOH resulting in the precipitation of 41%y of the uranium present 'ein a precipitate containing 5.1% U. 25 ml.rv portiens'of solu tion C were treated with 18.8 and 39.2 ml. of 0.1 N NaOH, respectively, resulting in the precipitation of 99.7 and 99.9% of the uranium in a precipitate containing 10.3 and 9.7% U.

EXAMPLE C 2 liters of 35% commercial phosphoric acid' containing 108 Ing/liter of UsO was reduced with iron and the uranium precipitated by the addition of 36 g. of calcium carbonate and 52.8 g. of .hydrogen fluoride. About 90% of the uranium was found in the precipitate. This precipitate containing 190 mg. of U3O8 was successively leached with sodium carbonate solution as follows:

Extracted Solution:

A 30 g. NazCO; iu 500 m1. H30 at 25 C 3. 6 B,.-. 40 g. NazCOa in 500 m1. H20 211:80 C... 69.0 0...- 25 g. NazCOa in 500 m1. H00 at 80 C.-. 68. 4 D 25 g. NazCOs in 500 ml. H20 at 80 0..-. 3G. 5

Total (mg. Usos) 177.5

120.7 g. of the solid remaining after leaching had the following composition:

The four solutions obtained from above were then employed in succession to leach a second similarprecipitate weighing 68 g. and containing 136 mg. U30?, by heating for 20 min. at 80 C. with the following results.

Usos content Usos Content. 'ofLeachng .After Leach- Solution, mg. Y ing, mg.

In the same manner another precipitatey was treated with these four solutions resulting in the U3O8 content of the combined solutions being raised to 500 mg. The combined solutions were heated with hydrochloric acid to remove the carbonate and ammonia was adde'd to precipitate the uranium. There was obtained a vprecipitate weighing 0.7338 g. whichY contained 66.3% U3O8, .23.8% P04, 9.7% SOQ, and 0.55% vanadium, thus indicating that the product was fairly pure uranyl phosphate.

EXAMPLE D Both leached (with Na2CO3) and unleached precipitated material was deiiuorinat'ed by heating-with ammonium sulphate and sulfuric acid a'sf indicated in Table V.

Table V E Fluorine (NHm'SOiy Time and Temp. Fluoride Wt. ofppt g Content, used, g. of heating sublim ed,

f .percent percent n.06 1 30. a 10. 02 24 lin/4 sa 2 5.00 49; 2 8. 20 1.5 hu] 97. 1 10.00 49. 2 16. 00 1.5 hr. 97. 9 10.00 44. 9 20. 00 1.5 hr. 99. 1 15.00 44. 0 10. 01 1.5 hr. 08. 9 t1.98 44. 9 8. 01 v1.5 b'rx/ 98. 5 5.06 2 44. 9 10. 00 1.5 hr. 08. 8 5.04 2 44. 9 18. 00 1.5 lin/450 90. 0 7.05 2 44. 9 3. 83 1.5,hr.l450 C-.-. 70. 8 5.01 2 44. 9 6. 02 1.5 IIL/450 90. 1

4e 1 135 /i hr /s75jor 07.1y 46 7 135 %hr.'/ 0 0..-- 99:0 46 7 135 l v%h1./500G..'.; 94. 4

I 'Lcached Unleached. v

EXAMPLE E Variousreducing agents were employedto lreduce ltwo literportions of `industrial (35%) phosphoric acid containing "1'08 mg. {130s/liter and the uranium was precipitated under the conditions'and with the results 'indicated inl Table VI.

Table lVl UsOain Recovery Reducing Agent Preeltltatiug ppt. Wt., ppt., perof UOs Agents g. cent fromacid,

g percent Y say' 0; 14" 72 s2 0. 2a s1 Uranium was recovered from several of the precipitates listed in Table VI by the following methods:

Precipitate (9) weighing 82 g. and containing 0.22% or 180 mg. of U30B was heated with 400 m1. of afsolution containing 50 g. of Na2CO3 for 2 hours with a resultant leaching of 128 mg. or 71.3% of the uranium.

Precipitate (10) a well washed precipitate weighing 187.8 g. wet which is equivalent to 53.3 g. dry and'containing 0.27% or 144 mg. of U3O8 was 'heated' for 11/2v hours with 400 ml. of a solution containing 50 g. NazCOa resulted inthe leaching of of'theuranium from;

the precipitate. This leach solution which now amounted to 595 ml. due to wash water inclusion was lthen ernployed 4to leach the entire wet precipitate l(l1)','after an additional 25 g. of Na2CO3 was added and' with heating for lil/ahours. The leach filtrate contained 259 mg. of UBOB amounting to a 57% extractionv of uranium from the .second precipitate in addition to thatextract'ed from thefrs't.

Precipitate (12) fromabove was subjected tosuccessive` leachings with "fresh portions of carbonate solution as follows:

4(1)"1`l1fey entire Yprecipitater containing 190 mg'. of U'BCg was -stirred'for l1/2" hours with afcold'solution of 50 g. of Ninco, in 50o mi. ofwater. The leaehura contained 3.6 mg. U30 K y (2) The residual precipitate from (l) was stirred for 45 minutes with heating withr a solution of 40 g. of Na2CO3 in 500 ml. of water and theprecipitate was separated from the solution by filtering. This filtrate con-V tained 69 mg. of USOS.

(3) The residual precipitate from 2, precedingwas heated with a 500 ml. of a solution containing 25 g. NagCO?,V for 1% hours andwas then separated from the solution by filtering. The iiltrate contained 68.4 mg.

EXAMPLE F A variety of reducing agents were employed to reduce 100 ml. portions of 35% Ycommercial vphosphoric acid containing 10.6 mg. U3Oa usingthe conditions and with the results ndicatedin Table VII. The reduced uranium was then precipitated by the addition of 1 g. CaO

i and 2.64 g. of HF to the reduced solution while con-- tained in a plastic beaker. In some cases HF was addedA with the reducing agent with a` consequent improvement in the amount of uranium precipitated. Increased temperature also improves the recovery.

Table VII t'llme, Tempera- Recovery Reducing Agent hr. ture, C. UaOx,

percent 30 g. Pb l0 mesh 95 8l 10 g. Pb 10 mesl1 2 95 67 l g. Pb l0 mesh.. 3 95 78 2 g. Bn 10 mesh..- )i 95 63 g. Sn 10 mes B l5 67 30 g. Cu Turnings-.. 1 95 64 g. Cu Turniugs 1-- 1 95 88 2 g. Cu 80 mesh shot. 1 95 76 2 g. Cu 8O mesh shot 95 86 2 g. Fe Turnings...-- 2 25 94 2 g. Fe Turnings 1 95 83 5 g. Fe Turnings-. 1 25 S9 4 g. Zu 20 mesh--.. 1 25 81 2 g. Zn 20 mesh 1 1 25 94 1 g. Zn Amalgarn 1 25 75 10 g. Zn Amalgam- 4 25 89 5 g. Zn Amalgam.. 1 25 8S 2 g. Zn Amalgam 1 25 88 1 2.47 g. HF added.

In the experiments using copper, only small amounts of copper appeared to dissolve. Zinc and zinc amalgam were effective reducing agents with only small amounts although zinc dissolved with evolution of hydrogen. Hydrogen is not evolved when lead is used.

In the foregoing, the uranium content is indicated in amounts or percentages of USOS. However, it is not intended that this terminology is to be construed literally but is intended to indicate that the equivalent quantity of uranium is present in the appropriate form as indicated by context. v

While there has been described in the foregoing what may be considered preferred embodiments of the invention, itis believed that various modifications may be made therein without departing from the spirit and scope of the invention and it is intended to cover all such as fall within the scope of the appended claims.

What is claimed is:

1. The method of recovering uranium values from a phosphatic solution containing impurity materials the iluorides of which are insoluble in said phosphatic solution comprising reducing the uranium to the uranous state in the solution, adding a soluble calcium coprecipitant compound to said solution, adding a soluble uoride to said solution, whereby said coprecipitant and some of said impurities are coprecipitated with theuranium to form complex' uoride precipitate, leaching uranium from the complex tluoride precipitate, and recovering uranium from the leach solution.

2. The method of recoving uranium values from a crude phosphatic solution containing materials the fluoridesA of which are insoluble in said phosphatic solution, comprising reducing the uranium to the uranous state in the solution, adding -a soluble calcium coprecipitant compound to said solution, adding a soluble uoride to said solution, whereby said coprecipitant and the uranium are coprecipitated to form a complex uoride precipitate, de-

liuorinating and calcining Vsaid precipitate, leaching uranium from said delluorinated and calcined precipitate, and recovering uranium from the leach solution.

3. The method of recovering uranium values from a crude phosphatic solution comprising reducing the uranium to theluranous state in the solution, adding a soluble calcium coprecipitant compound to said solution, adding a soluble uoride to said 'solution whereby the uranium is coprecipitated'with said coprecipitant to form a com plex lluoride precipitate, leaching uranium from said complex fluoride precipitate under acidic conditions, and

`recovering uranium from the leach solution.

precipitating said uranous uranium from the solution by adding soluble uorides and at least one coprecipitant selected from the group consisting of calcium, barium, magnesium, aluminum, sodium, potassium and ammonia which lform insoluble lfluoride insaid solution thereto, leaching uranium from said precipitate With an acid solution, whereby some phosphate is also leached, and treating said acidic solution with a reducing agent to precipitate said uranium from the leach solution as uranous phosphate.

5. The process for recovering uranium from acidic. phosphatic solutions comprising reducing said uranium to the uranous state in the solution, precipitating said uranous uranium from the solution by adding thereto soluble uorides and coprecipitants selected from the group consisting of calcium, barium, magnesium, aluminum, sodium, potassium and ammonia which form in soluble fluorides in said solution, deuorinating and calcining said precipitate by heating with a material selected from the group consisting of sulfuric acid and ammonium sulfate, leaching uranium from said deuorinated and calcined precipitate with an acidic aqueous solution, and recovering uranium from said acidic leach solution.

6. The method of recovering uranium from an acidic phosphatic solution comprising reducing said uranium to the uranous state in solution, precipitating said uranous uranium by adding to the solution soluble lluorides and coprecipitants -selected from the group consisting of calcium, barium, magnesium, aluminum, sodium, potassium and ammonia which form insoluble tluorides in said solution, leaching uranium from said precipitate with a carbonate solution, acidifying said leach solution, and precipitating the uranium from the acidited leach solution as uranium diuranate by the addition of ammonia.

7. The meth-od ofI recovering uranium from a phosphatic solution comprising reducing said uranium to the uranous state in the solution, precipitating said uranous uranium by adding to said solution a soluble uoride and coprecipitants selected from the group consisting of calcium, barium, magnesium, aluminum, sodium, potassium and ammonia whchform insoluble lluorides in said solution, leaching uranium from said precipitate with an acid solution, and precipitating the uranium from said solution as uranyl phosphate by adjusting the pH of the solution with an alkaline material.

8. The process as defined in claim 7, wherein said acid is a material selected from the group consisting of hydrochloric, nitric, citric, tartaric, acetic and boric acids.

9. The process of recovering uranium from an acidic phosphatic Solution containing. impurities. including phosphate, sulfate, iron,l silicon, vanadiumand titaniunncomprising reducing the uranium to the uranous state in thev solution, adding a soluble calcium coprecipitant compound` to the solution,lt reatiug said solution with a soluble uoride to precipitate said; material and at least some of said impurities from the solutionwhereby the uranium is coprecipitated with the. precipitate, leaching uranium from said precipitate with a material selectedfrom thef group consisting of sodiunieearbonate and ammonium carbonate, acidifying the leach solution, and precipitating the uranium from the acidiiied solution withy ammonia.

l0. The process fortrecovering uranium from anacdic phosphatic solution containing impuritiesfincluding phosphate, sulfate, iron,` silicon, vanadium, and titanium, com prising reducing the uranium to the-uranous state in the solution, adding a soluble calcium coprecipitant compound to the solution, treatingI Said solution with a soluble uoride to precipitate said material and at least some of said impurities from thesolution, whereby the uranium is coprecipitated4 with the precipitate, detluorinating and calcining said precipitateby heating with a material selectedfrom theA group consisting of. sulfuric acid and ammonium sulfate, leaching uranium from the deuorinated and caleined precipitate with amaterial selected from the group consisting of water, sulfuric acid and nitric acid, whereby some phosphate is also leached, andn precipitating uranyl` phosphate by making. the solution alkaline.

1l. The processfor recovering uranium from an acidic phosphatic solution containing impurities 'including phosphate, sulfate, iron, silicon, vanadium, and titanium, com.- prising reducing the uranium to the uranous statein the solution, adding aI soluble calcium coprecipitant com-l pound to the solution, treating saidsolutionwith a soluble uorideto precipitate.said'material andat least some of said impurities from the solution, whereby the uranium isl eopnecipitated withtheflprecipitate, defluorinating and calcining said precipitate by heatiugvvwith a material selectedfrom the; group-consisting of sulfuric acid and ammonium sulfate, leaching, uranium froml thedeuorinated and calcined precipitate with a material selected from theA group consisting, of water, sulfuric acid and nitric acid, whereby some phosphate is also leached, and precipitating` uranous4 phosphate from said solution by reducingthe uranium tothe uranousstate.

12. The process for recoveringuranium from an acidic phosphatic-v solution comprising; reducing the uranium to the uranous state in solution, adding to said solution copriecipitaut. materials selectedv from the group consisting of -calcium,vbarium, I naguesium, aluminum, sodium, potassiurn and ammoniaV which form. insoluble fluorides in said solution,treatingsaidl solution with soluble uorides to, precipitate said coprecipitantmaterials, whereby the uranium is coprecipitated thenewithJcachingthe uranium from the precipitate with an, acidic solution, whereby some phosphate is also leached, and precipitating uranous phosphate from the, solution; by. reducing the uranium to the uranous state.

1.3. The process as defined in claim l2, wherein said acidic solution which is, employed toleach't'he precipitate is amaterial selected-fromme.- group consisting of hydrochloric, nitric;` citric, tartaric, acetic andboric acid solutions; t

1,4. ThevProcoSs as.A deinedf inf claim l0 wherein said soluble .lluQride is ai material; selected, from the group consisting of sodium, potassium, barium and, ammonia fluorides.

References Citediny the le of this patent Mellor: Inorganicy andTheoretical.Chemistry, vol. 12, pageA 74 (1932). Publ. by Longmans, Green & Co., London;

Claims

1. THE METHOD OF RECOVERING URANIUM VALUE FROM A PHOSPHATIC SOLUTION CONTAINING IMPURITY MATERIALS THE FLUORIDES OF WHICH ARE INSOLUBLE IN SAID PHOSPHATIC SOLUTION COMPRISING REDUCING THE URANIUM TO THE URANOUS STATE IN THE SOLUTION, ADDING A SOLUBL E CALCIUM COPRECIPITANT COMPOUND TO SAID SOLUTION, ADDING A SOLUBLE FLUORIDE TO SAID SOLUTION,WHERBY SAID COPRECIPITATED AND SOME OF SAID IMPURITIES ARE COPRECIPITATED WITH THE URANIUM TO FORM A COMPLEX FLUORIDE PRECIPITATE, LEACHING URANIUM FROM THE