Leaching yield

The residue after the water wash is leached at an elevated temperature by sulfuric acid, part of which is the recycled raffinate from the vanadium extraction. The leaching yield of vanadium (mainly IV-valent) is about 55% and of nickel about 95%. A final (post) leach with sodium hydroxide dissolves the remaining vanadium (mainly V-valent). The resulting leach solution, containing practically all the vanadium ( 25gdm ) and nickel ( 12gdm ) is fed to the solvent extraction circuit. 

Three of the four resins yield extracts of the functionality of Equation 4. The slope of the exponential decay allows for the evaluation of x. The resin, see Table II, initiated by benzyl dimethyl amine (BDMA) at the stated cure cycle, when subjected to leaching yields an extract of low solubility and a distribution of oligomeric molecules of low number average molecular weight. 

Catalyst preparations. Y zeolites with overall Si/Al ratio equal to 2.7, 15 and 45 were used as supports. The Y2>7 sample was a stabilized zeolite available from Union Carbide (LZY 82). This material when submitted to selfsteaming and acid leaching yielded HY4g. The HY15 sample derived from a commercial NaY (Union Carbide LZY 52) the transformation involved successive ammonium exchange, self-steaming and acid washing. 

With all t the catalysts from Ni Alg were most active. As to PM catalysts their activity is increased up to 75% A1 (ref.9). As shown above the composition Ni A1 5 is related by a solid solution on the NlAl-basls with a maximum possible deficit of Ni atoms. Alloys with a higher Hi content are close to equilibrium NiAl. whereas with the smaller one they have other phases (HiAl. Ni Al ). Apparently, the highest activity of these catalysts is due to the fact that as a result of MA Nj Al leaching the largest quantity of most dispersive Ni is formed, which is fixed on unleached NiAl particles. The thing is that richer nickel alloys are leached to a less extent and contain less Ni skeleton phase. Whereas Ni2.Al3 and NiAl leaching yields are not so fine particles of an active metal. 

Aluminum. All primary aluminum as of 1995 is produced by molten salt electrolysis, which requires a feed of high purity alumina to the reduction cell. The Bayer process is a chemical purification of the bauxite ore by selective leaching of aluminum according to equation 35. Other oxide constituents of the ore, namely siUca, iron oxide, and titanium oxide remain in the residue, known as red mud. No solution purification is required and pure aluminum hydroxide is obtained by precipitation after reversing reaction 35 through a change in temperature or hydroxide concentration the precipitate is calcined to yield pure alumina. 

Fusion with caustic soda at 500—800°C in an iron cmcible is an effective method for opening pyrochlores and columbites (20). The reaction mixture is flaked and leached with water to yield an insoluble niobate which can be converted to niobic acid in yields >90 wt% by washing with hydrochloric acid. 

Although this procedure yields tellurium as the same compound found in the original feedstock, the copper teUuride is recovered in a comparatively pure state which is readily amenable to processing to commercial elemental tellurium or tellurium dioxide. The upgraded copper teUuride is leached with caustic soda and air to produce a sodium teUurite solution. The sodium teUurite solution can be used as the feed for the production of commercial grade teUurium metal or teUurium dioxide. 

A number of high temperature processes for the production of titanium carbide from ores have been reported (28,29). The aim is to manufacture a titanium carbide that can subsequently be chlorinated to yield titanium tetrachloride. In one process, a titanium-bearing ore is mixed with an alkah-metal chloride and carbonaceous material and heated to 2000°C to yield, ultimately, a highly pure TiC (28). Production of titanium carbide from ores, eg, ilmenite [12168-52-4], EeTiO, and perovskite [12194-71 -7], CaTiO, has been described (30). A mixture of perovskite and carbon was heated in an arc furnace at ca 2100°C, ground, and then leached with water to decompose the calcium carbide to acetjdene. The TiC was then separated from the aqueous slurry by elutriation. Approximately 72% of the titanium was recovered as the purified product. In the case of ilmenite, it was necessary to reduce the ilmenite carbothermaHy in the presence of lime at ca 1260°C. Molten iron was separated and the remaining CaTiO was then processed as perovskite. 

In most ores, sufficient Fe is already present. For some ores, it is necessary to add metallic iron. In practice, the oxidation potential of the solution can be monitored and controlled using the Fe /Fe ratio. Very high leaching efficiencies with H2SO ate common, eg, 95—98% dissolution yield of uranium (39). If acid consumption exceeds 68 kg/1 of ore treated, alkaline leaching is preferred. The comparative costs of acid, sodium hydroxide, and sodium carbonate differ widely in different areas and are the determining factor. 

The principal vanadium-bearing ores are generally cmshed, ground, screened, and mixed with a sodium salt, eg, NaCl or Na2C02- This mixture is roasted at ca 850°C and the oxides are converted to water-soluble sodium metavanadate, NaVO. The vanadium is extracted by leaching with water and precipitates at pH 2—3 as sodium hexavanadate, Na V O, a red cake, by the addition of sulfuric acid. This is then fused at 700°C to yield a dense black product which is sold as technical-grade vanadium pentoxide. This product contains a minimum of 86 wt % V20 and a maximum of 6—10 wt % Na20. 

The ore is ordinarily ground to pass through a ca 1.2-mm (14-mesh) screen, mixed with 8—10 wt % NaCl and other reactants that may be needed, and roasted under oxidising conditions in a multiple-hearth furnace or rotary kiln at 800—850°C for 1—2 h. Temperature control is critical because conversion of vanadium to vanadates slows markedly at ca 800°C, and the formation of Hquid phases at ca 850°C interferes with access of air to the mineral particles. During roasting, a reaction of sodium chloride with hydrous siUcates, which often are present in the ore feed, yields HCl gas. This is scmbbed from the roaster off-gas and neutralized for pollution control, or used in acid-leaching processes at the mill site. 

For solvent extraction of pentavalent vanadium as a decavanadate anion, the leach solution is acidified to ca pH 3 by addition of sulfuric acid. Vanadium is extracted in about four countercurrent mixer—settler stages by a 3—5 wt % solution of a tertiary alkyl amine in kerosene. The organic solvent is stripped by a soda-ash or ammonium hydroxide solution, and addition of ammoniacal salts to the rich vanadium strip Hquor yields ammonium metavanadate. A small part of the metavanadate is marketed in that form and some is decomposed at a carefully controlled low temperature to make air-dried or fine granular pentoxide, but most is converted to fused pentoxide by thermal decomposition at ca 450°C, melting at 900°C, then chilling and flaking. 

For direct precipitation of vanadium from the salt-roast leach Hquor, acidulation to ca pH 1 without the addition of ammonia salts yields an impure vanadic acid when ammonium salts are added, ammonium polyvanadate precipitates. The impure vanadic acid ordinarily is redissolved in sodium carbonate solution, and ammonium metavanadate precipitates upon addition of ammonium salts. Fusion of the directly precipitated ammonium salts can yield high purity V20 for the chemical industry. Amine solvent extraction is sometimes used to recover 1—3 g/L of residual V20 from the directly precipitated tail Hquors. 

Other Reductions. Ductile, pure zirconium has been made by a two-stage sodium reduction of zirconium tetrachloride (68) in which the tetrachloride and sodium are continuously fed into a stirred reactor to form zirconium dichloride [13762-26-0], heating with additional sodium yields zirconium metal. Leaching with water removes the sodium chloride from the zirconium. Bomb reduction of pure zirconium tetrafluoride with calcium also produces pure metal (69). 

The filtered sodium thioantimonate solution obtained from the leaching of stibnite with sodium sulfide may also be reduced directly to metal by elemental sodium (17). Yields ia excess of 95% of 99.5 pure antimony are claimed (18). 

BaS is leached from the black ash by hot water. The resultkig solution is filtered and then treated with soda ash or carbon dioxide or a combination of the two to precipitate fine BaCO crystals, which ki turn ate filtered and dried. Sulfide values can be recovered as H2S, NaHS, Na2S, or elemental sulfur. Carbon dioxide, detrimental to high BaS yields, is repressed according to the Boudouard equiUbtium... 

In the hot leaching step, barium metasiUcate is insoluble, and the orthosiUcate yields only one-half of its barium value to give barium hydroxide... 

Electrowinning. Vat leaching often yields copper solutions having concentrations sufficiently high for direct electrowinning. However, high concentrations of cations other than copper and low copper concentrations make it more difficult to obtain high purity electrolytic copper by direct electrolysis of leach solutions than by electrolysis of purified solutions obtained from solvent extraction. 

Acylglycerols can be hydrolyzed by heating with acid or base or by treatment with lipases. Hydrolysis with alkali is called saponification and yields salts of free fatty acids and glycerol. This is how soap (a metal salt of an acid derived from fat) was made by our ancestors. One method used potassium hydroxide potash) leached from wood ashes to hydrolyze animal fat (mostly triacylglycerols). (The tendency of such soaps to be precipitated by Mg and Ca ions in hard water makes them less useful than modern detergents.) When the fatty acids esterified at the first and third carbons of glycerol are different, the sec-... 

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