As an example of the acid-leaching, solvent extraction class of uranium-concentration processes, a description will be given of the process used in the large uranium mill of the Kerr-McGee Corporation at Grants, New Mexico. This has been condensed from a 1960 paper by Hazen

tYellow cake is the name conventionally used for uranium ore concentrates.

[H4] and from Merritt’s [М3] account of operations in 1971. At that time the mill’s capacity was 5000 short tons ore (4500 Mg) per day. In 1978 its capacity was 6200 short tons per day. Figure 5.5 shows this mill.

Leaching operations in the Kerr-McGee mill are described in this section, with reference to Fig. 5.6. Recovery of uranium from leach liquor by solvent extraction with organic amines in the Amex process is to be described in Sec. 8.6.

The ore processed in the Kerr-McGee mill is primarily a sandstone, with uranium minerals in the material bonding the sand grains. The ore contains about 0.2 w/o U308, 0.01 to 0.03 w/o Mo03, and 0.05 to nearly 0.20 w/o V205. Uranium and molybdenum are leached and recovered, with uranium recovery exceeding 97 percent. Some vanadium is also leached, but was not being recovered in 1971. The ore also contains acid-soluble calcium minerals, equivalent to from 2 to 5 w/o CaO, which are the principal consumers of sulfuric acid.

Crushing and grinding. The ore is first crushed dry to particles smaller than 1 in. Crushed ore is then processed in two parallel, identical systems, of which circuit A is shown in Fig. 5.6. In each circuit 2500 t per day of crushed ore is ground with heated water in rod mills until 97 to 98 percent passes 28 mesh, with 70 percent coarser than 150 mesh.

Leaching. Slurry from the rod mills flows in series by gravity through 14 rubber-lined steel leaching tanks 13 ft (4 m) in diameter and 14 ft (4.25 m) high, equipped with turbine agitators. The holding time in the 14 tanks is around 4.5 h. The rod-mill slurry fed to the first tank is mixed with recycle water containing slimes, and steam and sulfuric acid sufficient to bring the temperature in the first tank to 43 to 54°C and the pH to 0.6 to 0.7. Here the most readily dissolved minerals react, and gaseous reaction products such as C02, H2, and H2S are liberated.

To leach the more acid-resistant minerals containing tetravalent uranium, steam is fed to the second tank to bring the temperature to 49 to 60°C, and sodium chlorate NaQ03 is added to bring the oxidation-reduction potential e, measured relative to the calomel electrode, to from —0.47 to —0.51 V. At —0.51 V, the equilibrium ratio of ferric iron to ferrous iron in the solution is 0.52.^ Ferric iron catalyzes the oxidation of insoluble tetravalent uranium to the soluble hexavalent uranyl form:

^Oxidation-reduction potentials and their effect on the valence state of materials being processed are explained in Chap. 9. Because the emf of the saturated KCl-calomel electrode relative to the standard hydrogen electrode is -0.244 eV at 25°C ([М3], p. 70), the relation between the emf e relative to the saturated calomel electrode and the emf E relative to the standard hydrogen electrode used elsewhere in this text is E° = e — 0.244. Thus, the emf in the second tank relative to the standard hydrogen electrode is from -0.714 to -0.754 V.

U4* + 2Fe3+ + 2H20 -*■ UVI02Jt + 2FeJ+ + 4H+

Addition of sodium chlorate to the second tank instead of to the first avoids consumption of this relatively expensive material by metallic iron introduced in grinding or by reducing gases, such as Hj or Ц* S, which are vented from the first tank.

As the hot, acid, oxidizing slurry flows through the remaining 13 tanks, dissolution of the more resistant uranium and molybdenum minerals is completed. After 4.5 h, when the slurry leaves the fourteenth tank, its temperature has dropped to 43 to 54°C, acid has been consumed with pH increased to 0.9 to 1.2, and sodium chlorate has been used up in oxidizing uranium, with the oxidation-reduction potential relative to the calomel electrode in the range —0.41 to —0.43 V. This is sufficiently negative to convert substantially all soluble uranium to the hexavalent state, while leaving only 2 percent of the iron oxidized to ferric, thus minimizing consumption of sodium chlorate.

In some other U. S. mills and in South Africa and Australia, manganese dioxide Mn02, syn­thetic or in the form of the mineral pyrolusite, is used as oxidant. Typical oxidant consumption is 3 lb (1.5 kg) NaC103 or from 3 to 6 lb (1.5 to 3 kg) Mn02 per short ton of ore.

Sulfuric acid consumption depends on the amount of reactive minerals present in the ore. In the United States, acid requirements range from 40 to 120 lb (20 to 60 kg) H2SC>4 per short ton of ore.

liquid-solid separation. Separation of the slurry leaving leach tank #14 into (1) tailings relatively free of uranium-containing liquid and (2) uranium-bearing leach liquor free of suspended solids is shown at the bottom and right of Fig. 5.6. A rough separation of slurry from leach tank #14 at about 150-mesh particle size is made in two 20-in (0.5-m) diameter cyclone separators in parallel. The coarse fraction from the cyclones passes through five rake classifiers in series, where the sand is washed countercurrently with acid-bearing aqueous raffinate from the solvent extraction system, Fig. 5.9. The fines fraction from the cyclone separators and the overflow from #1 rake classifier are combined and washed countercurrently with additional raffinate in six large countercurrent decantation thickeners 120 ft (36 m) in diameter and 17 ft (5 m) deep. Wash ratios are 2.5 to 3.0 in the classifiers and 3.0 to 4.0 in the thickeners. This recycle of raffinate returns H2S04 from the solvent extraction system to the washing circuits and maintains the pH in them at 1.5 or less, thus preventing precipitation of uranium during washing. Tailings from #5 rake classifer are about 75 v/o (volume percent) solids; from #6 thickener, about 31 v/o. The thickeners handle about 25 percent of total solids. Soluble uranium losses in the washed tailings represent only about 0.2 percent of the uranium in mill feed.

Overflow from #1 thickener contains 150 to 200 ppm of solids. This is reduced to 50 to 75 ppm by addition of flocculant and settling in a 60-ft (18-m) diameter by 20-ft (6-m) deep reactor-clarifier. Fine slimes from the clarifier are recycled to leach tank #1.

Overflow from the clarifier is combined with a similar stream from leaching circuit В and passed through six 600-ft5 (58-m2) U. S. pressure filters in parallel, each operated at a feed rate of 300 to 500 gal/min (1.1 to 1.9 m3/min). Filters are precoated with about 0.1 lb (0.05 kg) of precoat per short ton of ore, and 0.35 lb (0.17 kg) of filter aid per ton is added to the filter feed. The filter loading cycle lasts from 4 to 24 h, depending on the solids contents of clarifier overflow. Filter cake is returned to #2 or #3 thickener. Filtered leach liquor containing about 1 g U308/liter is the product of the leaching system and the feed to the solvent extraction system.