Freeport Minerals Company

Freeport-mcmoran Corporation

Freeport Research and Engineering Co.

Freeport Research & Development Company

Other

Freeport Mcmoran Resource Partners

Freeport Mcmoran Resource Partners, Ltd. Partnership

Tadeusz K Wiewiorowski

Vivian C Astley

Phillip D Mollere

David J Miller

Freddie J Touro

Willis L Thornsberry, Jr

Michael A Murray

Jody J Taravella

Dwight S Nunley

James T Womack

Jesus W Perez

Thomas A Headington

Michael O Nutt

David M Dyer

A toilet flush control system which provides control over the volume of water used in conventional tank-type toilets. The disclosed toilet flush control apparatus and method can be used in existing toilets by easily retrofitting such toilets with said apparatus. The toilet flush control apparatus may also be incorporated into the design of new toilets. The apparatus comprises a flapper-type flush valve which by itself does not have adequate buoyancy to remain open during the flush cycle of a toilet and which is equipped with an eyelet member; a buoyant member capable of imparting the needed buoyancy to the valve during the flush cycle; a flexible line connected at one end to the buoyant member and passing through said eyelet, and mechanism for adjusting the vertical distance between the flush valve and the buoyant member from outside the toilet tank. The volume of water saved during flushing is equivalent to the volume of water between the level of the flush valve and the level of the buoyant member.

Water saver toilet flush control system

A process for recovering gold from gold-bearing ores includes the steps of forming a slurry of ground ore and treating the ore with an oxidizing and complexing agent such as hydrogen sulfide or a sulfide salt. A chelating agent, SO.sub.2, air, and an anion exchange resin are added to the slurry to subject the ore to a simultaneous sulfurous acid leaching and anion exchange resin adsorption step in the presence of dissolved oxygen. The gold transfers to the resin which is then removed from the slurry. The gold is then separted from the resin by chemical stripping. Alternately, the ground ore is blended with an oxidizing agent such as calcium hypochlorite and then with a gold-complexing agent. The ore blend is slurried in water, with the addition of a chelating agent, SO.sub.2, air and an anion exchange resin to effect the leaching and adsorption of gold values.

Process for recovery of gold from gold ores using a complexing

A process for recovering gold from refractory gold-bearing ores uses sulfurous acid as the leaching agent to form a gold-sulfite complex. The ore is ground, slurried blended with a chelating agent and then subjected to a simultaneous dissolved SO.sub.2 leaching and anion exchange resin adsorption step in the presence of dissolved oxygen. The gold transfers to the resin which is later separated from the resin by chemical stripping.

Process for recovery of gold from refractory gold ores using sulfurous

An improved process for treating spent alkylation acid to recover a sulfuric acid product suitable for use in wet process phosphoric plants particularly those which recover uranium as a byproduct. The process utilizes heat and agitation to polymerize the liquid, soluble organic impurities normally contained in spent alkylation acid to insoluble, carbonaceous solids. The carbonaceous solids formed are relatively inert in acidic environments, easy to handle and have valuable cation exchange and impurity scavenging properties.

Spent alkylation acid treatment process

A process for recovering elemental sulfur from volcanic and non-volcanic ores is disclosed. Ore is processed by crushing the ore to a coarse particle size autoclaving to separate the sulfur from the ore, mild grinding to selectively reduce the size of the liberated sulfur particles, flotation and sulfur separation/filtration steps to produce a sulfur of at least 99% purity. The process recovers at least 90% of the elemental sulfur contained in the ore. The particular crushing of the ore and selective grinding techniques for reducing sulfur particles for reducing sulfur particles size are important factors in the efficiency of the process since the grinding techniques inhibit the formation of fine gangue which is difficult to separate from the liberated sulfur.

Recovery of sulfur from native ores

A process for producing high grade silica and fluorine-containing coproducts from fluosilicic acid is described. To produce silica, the fluosilicic acid is reacted with aqueous ammonia to produce a solid silica product and a solution of ammonium fluoride. The solid silica is washed under controlled conditions and further processed to produce a high grade product. The ammonium fluoride solution is concentrated and sold as a concentrated ammonium fluoride solution product or reacted with metal hydoxides or oxides to produce metal fluorides.

Production of silica and fluorine-containing coproducts from fluosilicic

The present invention discloses a new method for obtaining a crystallized phosphoric acid of relatively high purity starting with conventional phosphoric acid. The new method is based on crystallization of orthophosphoric acid hemihydrate from the phosphoric acid. Massive nucleation which would lead to the formation of unprocessable masses is avoided in the new method by providing unusually large amounts of fine relatively pure seed crystals of orthophosphoric acid hemihydrate and operating the crystallization process under conditions which favor crystal growth on the seed crystals and disfavor the occurrence of secondary nucleation. To prevent the crystallizing magma from reaching a viscosity which would render further processing difficult the present invention provides for recycling raffinate in an amount sufficient to maintain the solids content of the crystallizing magma below about 40%.

Phosphoric acid crystallization process

The present invention relates to the processing of phosphoric acid by crystallization. Wet-process phosphoric acid is purified by cooling it to supersaturated conditions and adding relatively large amounts of fine, relatively pure seed crystal. Crystallization thereafter proceeds under conditions which substantially favor crystal growth on the seeds which are added and disfavor secondary nucleation. To prevent the crystallizing magma from reaching a viscosity which would render further processing difficult, the present invention provides for recycling raffinate in an amount sufficient to maintain the solids content of the crystallizing magma below about 40%.

A method for the recovery of precious metals such as gold and silver from various ore types is described which involves subjecting a slurry of the ore to a simultaneous leaching and electrodeposition process by mixing the slurry with a reagent such as an alkaline cyanide solution which provides for the leaching requirement and contacting said slurry with a metallic cathode with a negative electric potential applied thereto providing for the electrodeposition requirement. The cathode is made of a metal selected from the group consisting of cadmium, copper, iron, lead, molybdenum, tin, zinc, cobalt, nickel, silver, titanium, tungsten, vanadium and alloys and mixtures containing at least one of these metals. The simultaneous leaching and electrodeposition occur under conditions controlled to afford at least partial dissolution of the precious metal values from the ore, whereby continuous transfer of the precious metal from the ore onto the surface of the cathode is promoted. The resultant electrodeposition product, i.e., the cathode with precious metal values electrodeposited thereon, is then separated from the ore slurry and subjected to a subsequent precious metal recovery step by conventional methods.

Simultaneous leaching and electrodeposition of precious metals

A method for the recovery of precious metals such as gold and silver from various ore types is disclosed which involves subjecting a slurry of the ore to a simultaneous leaching and cementation process with a reagent such as an alkaline cyanide solution providing for the leaching requirement and a reducing metal affording cementation. The reducing metal is selected from the group consisting of cadmium, copper, iron, lead, molybdenum, tin, zinc and alloys and mixtures comprising at least two of these metals. The simultaneous leaching and cementation occur under conditions controlled to afford at least partial dissolution of the precious metal values from the ore, whereby continuous transfer of these values from the ore onto the surface of the reducing metal is promoted. The resultant cementation product, i.e., the reducing metal with precious metal values cemented thereon, is separated from the ore slurry and subjected to a subsequent precious metal recovery step by conventional methods.

Simultaneous leaching and cementation of precious metals

A process for the selective separation of ferric iron from an aluminum fluoride aqueous solution containing ferric iron is effected by contacting said solution with an organic extractant phase consisting essentially of a mixture of a mono (higher alkyl-substituted phenyl) phosphoric acid and a di(higher alkyl-substituted phenyl) phosphoric acid dissolved in an inert diluent to thereby transfer the ferric iron to said organic extractant phase. The efficiency of the process is further improved by adjusting the aluminum fluoride aqueous feed solution to a temperature in the range of about 140.degree. F. to about 175.degree. F. prior to contacting it with the organic extractant phase, maintaining the temperature within this range throughout the ferric iron extraction stage or stages, then heating the resulting iron-depleted aqueous raffinate to a temperature higher than 195.degree. F., and then crystallizing aluminum fluoride trihydrate out of the raffinate at this higher temperature.

Ferric iron removal from aluminum fluoride solutions

Process for the recovery of uranium from wet process phosphoric acid solution in which an organic extractant, containing uranium values and dissolved iron impurities and comprising a dialkylphosphoric acid and a trialkylphosphine oxide dissolved in a water immiscible organic solvent, is contacted with a substantially iron-free dilute aqueous phosphoric acid to remove said iron impurities. The removed impurities are bled from the system by feeding the resulting iron-loaded phosphoric acid to a secondary countercurrent uranium extraction operation from which they leave as part of the uranium-depleted acid raffinate. Also, process for recovering uranium in which the extractant, after it has been stripped of uranium values by aqueous ammonium carbonate, is contacted with a dilute aqueous acid selected from the group consisting of H.sub.2 SO.sub.4, HCl, HNO.sub.3 and iron-free H.sub.3 PO.sub.4 to improve the extraction efficiency of the organic extractant.

Process for recovery of uranium from wet process phosphoric acid

An ammoniacal solution containing nickel and cobalt is contacted with an extractant comprising an oxime or oxine compound dissolved in an organic diluent to selectively extract the nickel into the organic phase while leaving most of the cobalt in the aqueous ammoniacal phase from which it is subsequently recovered. The nickel-loaded organic extract is washed with water and aqueous carbon dioxide to remove and recover ammonia which undesirably transfers to the organic phase during the extraction. The ammonia-free, nickel-loaded extract is then stripped with aqueous sulfuric acid at low temperature to remove the nickel. The ammonia-free nickel sulfate solution produced in the stripping operation is particularly suitable for recovering the nickel by electrolytic deposition. The stripped organic extract is further stripped with aqueous sulfuric acid at high temperature to remove and recover the small amount of cobalt which transfers to the organic phase during extraction and hampers reuse of the organic extractant in the process. Removal of the cobalt permits recycle of the extractant to the process.

Solvent extraction of nickel from ammoniacal solutions

An ammoniacal solution containing nickel and cobalt dissolved as nickel-ammonia complexes and cobalt-ammonia complexes is treated with a material capable of providing free ammonia in the solution, such as gaseous ammonia or aqueous ammonia, in order to increase the proportion of higher nickel-ammonia complexes to lower nickel-ammonia complexes in solution until at least about 85% of the dissolved nickel is in the form of higher nickel-ammonia complexes, i.e., complexes in which the number of NH.sub.3 molecules is greater than 3. The attainment of this high concentration of higher nickel-ammonia complexes is readily determined by various analytical procedures such as, for example, free ammonia electrode measurements and spectrophotometer measurements. The solution is then treated with a sulfiding agent in an amount sufficient to selectively precipitate out the dissolved cobalt as cobalt sulfide. The resulting slurry is separated into a nickel-enriched liquid fraction and a cobalt-enriched solids fraction. Surprisingly, when the dissolved nickel is present as a higher ammonia complex, less tends to undesirably coprecipitate with the cobalt during sulfiding. The highly desirable result is a precipitate containing up to 50% more cobalt and 20% less nickel, and a mother liquor more enriched in nickel, than normally obtained in a conventional selective sulfiding.

Enhanced selectivity in the separation of nickel and cobalt from

Aqueous AlF.sub.3 enriched crystallization mother liquors contaminated with P.sub.2 O.sub.5 values and having less than 5% free acidity are treated with a material which provides bismuth ions in the mother liquor in amounts small enough to insure that the Bi/PO.sub.4 mole ratio remains below about 1.05. Under these conditions, the bismuth and P.sub.2 O.sub.5 react essentially quantitatively to form a BiPO.sub.4 precipitate by which the P.sub.2 O.sub.5 can be separated from the mother liquor. Bismuth can be regenerated from the BiPO.sub.4 and recycled for further use in the process.

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New Orleans, LA, United States of America

Tadeusz K Wiewiorowski