How to Refine Copper Cement

Cement copper production in the United States has increased during the past few years, principally because more submill-grade material from open-pit operations is being leached in waste dumps. Since 1960, the cement copper annual growth rate has been about twice the rate for primary copper. If this trend continues, annual production of copper from cement copper precipitates will be 550,000 tons by 1975.

In commercial cementation plants, shredded detined cans are the chief source of iron used as the copper precipitant. The cement copper produced contains from 60 to 90 percent copper. Oxygen is the principal impurity owing to partial oxidation of the moist, fine copper particles during drying. Copper precipitates also are contaminated with silica, unreacted iron, and extraneous debris associated with tin cans. Standard practice is to smelt this impure cement copper in the reverberatory furnace along with copper sulfide flotation concentrates. Because of the increasing production of cement copper and absence of excess capacity at existing smelters, alternative processing routes to present smelting practice should be considered.

The direct smelting and fire refining of cement copper in an electric furnace was investigated by the Bureau of Mines in the early sixties. This approach has not been commercially adopted. Other

Copper Electrolite Solution Filtration

The electrolyte in-between the active electrodes must be purged with higher velocities to maintain a high concentration of the desirable copper ion in the liquid film adjacent to the cathode and the sulfate ion in the liquid film at the anode. This higher velocity also carries away from the cathode the undesirable impurities released from the anode as they are released into the electrolyte as filterable solids.

The Onahama refinery produced 5,300 metric tons of electrolytic copper per month in 1968 with a planned expansion to 12,000 tons per month. There are over 300 electrolytic cells in the tank house. About 540,000 gallons of copper sulfate electrolyte solution are recirculated through the cells at the rate of 5,000 gallons per minute. This means that all the electrolyte in the cells is replaced once every 80 minutes. About 400,000 gallons are in the cells and 150,000 gallons in a surge or service tank.

The copper sulfate concentration in the electrolyte is maintained at 49 gms per liter and the free sulfuric acid at 200 grams per liter. In addition to filtering the electrolyte at the rate of 2600 GFM through diatomaceous earth filter aid the electrolyte is also purified by dialysis. The electrolyte

Anodic Dissolution of Zinc Sulfide Suspensions in Aqueous Solution

A new suspension electrolysis method has been developed, which enables without any gas or dust pollutants a direct anodic extraction of zinc from zinc sulfide concentrates, while simultaneously oxidizing sulfide sulfur to elemental sulfur. The investigations were performed in a laboratory scale electrolytic cell with suspended zinc sulfide concentrate particles with or without graphite powder additions in dilute sulfuric acid electrolyte.


The cathode, a rod of aluminum, is placed in the center of cell. The anode consists of four plates mounted on the cell wall. The anode materials used in this investigations are PbO2(Pb), graphite and Pt. The cathode compartment and the anode compartment are separated by a “RCH 1000 Supralen-M-Type” membran, which prevents the transport of sulfide particles to the cathode area. With the capillary the reference electrode Hg/Hg2SO4 (sat. K2SO4) is connected with the system. To observe the suspension, which is obtained by a magnetic stirrer, the cell consists of transparent acryl glass. The temperature was kept constant by a thermostat in combination with circulating water in the double-walled cell.

With zinc blende particles suspended in 1 M H2SO4 solution a redox potential of about 0,15 V vs NHE could be observed. To avoid during the electrolysis of

Pilot Plant Operation & Manganiferous Zinc Concentrates

Pilot plant operations are described for the treatment of manganiferous zinc concentrates from the Gamsberg ore deposit in South Africa via conventional electrolytic zinc processing. Two methods were employed for the control of excess manganese ions in the pilot plant electrolyte involving (1) chemical oxidation and (2) anodic oxidation.

Description of Pilot Plant

The pilot plant was designed to produce a nominal 100 kg of cathode zinc metal per day. This resulted in the treatment of approximately 200 kg of zinc calcine and the production of 1000 liters of impure zinc sulfate leach liquor per day. The individual process steps were carried out batchwise with the exception of zinc electrolysis, which was conducted on a continuous 24 hour basis.

Plant piping was of high density polypropylene. Storage tanks were polypropylene, fiberglass, polyethylene or rubber-lined depending on the temperature of the liquors entering. The cathodes and anodes (7 cathodes and 8 anodes) used were of commercial plant size (1.05 m x 0.6 m) rented from Zincor. The anodes were aged in the Zincor refinery for a period of eight weeks prior to use in the NIM pilot plant. The cathodes were similarly treated at Zincor prior to use at NIM. Total submerged

How to Separate W from AgNO3 Electrolyte

A prevalent technique for the recovery of silver from wastes generated in silver-tungsten electrical contact fabrication is electrowinning. Such scrap is placed in a permeable plastic basket which constitutes the anode compartment and this basket is suspended in a larger cell in which the cathode is mounted.

Process Design

The solution to this problem lies in the selective removal of the W from the AgNO3 electrolyte either by liquid-liquid or liquid-solid extraction. Although W extraction in the liquid-liquid mode with a quaternary amine or other reagents as extractants appears feasible, particularly in view of the low pH (-1.5) of the system, it was eliminated from consideration because it is essentially a batch type process and it is a more elaborate industrial operation than extraction by ion exchange resins.

A useful material is a gel-type, weakly basic anion exchange resin with a polyamine functionality and in the free base form, the commercial Amberlite IRA-68 manufactured by Rohm and Haas Company. Initial slurry tests with W-contaminated AgNO3 electrolyte showed that the resin extracts W quantitatively but it also coextracts substantial amounts of Ag.

Process Description

The flow diagram of the process for the separation of W from AgNO3 electrolyte

  1. Resin conditioning
  2. Column preparation

Acid Dissolution of Metal & Mineral Samples by Microwave


Reduce the time, complexity and expense of dissolving metal and mineral samples in solution in preparation for chemical analysis.


In earlier research, the Bureau of Mines developed a rapid and inexpensive method for the dissolution of mineral and metal samples in plastic pressure bottles heated in a boiling water bath. Using this technique, samples could be prepared for chemical analysis in less than one hour compared with several hours by traditional dissolution methods (see Technology News No. 85, December 1980 for further details). By heating in a microwave oven, sample dissolution time can be reduced to approximately 5 minutes.

microwave acid dissolution

An inexpensive microwave oven provides extremely precise and efficient heating of the sample and acid mixture. Increased temperature and pressure in the sealed bottles accelerate sample dissolution.

The Improvements

Samples are prepared as previously described by grinding them to minus 100- mesh and placing a small weighed portion of each into polycarbonate bottles. A mixture of hydrochloric, nitric, and hydrofluoric acids is added to each bottle. The bottle caps then are screwed on tightly to prevent contamination and to retain volatile gases which could be released during the dissolution

What is the Effects of Impurities in Electrolytes on Electrowinning of Lead from Lead Chloride

Lead is one of the oldest metals known to man and has been used for hundreds of years. The method for producing lead from galena has changed very little. A lead concentrate is mixed with fluxing agents, roasted to remove sulfur, and heated to about 1,000° C with carbon to obtain an impure metallic product, which is refined. The smelting process is a low-cost operation but results in SO2 and lead emissions.

The Bureau of Mines under a cost-sharing program with St. Joe Minerals Corp., ASARCO, COMINCO, Ltd., and AMAX has investigated an alternative method for producing lead. The method consists of leaching galena concentrate with a ferric chloride-sodium chloride solution to make lead chloride. The lead chloride is electrolyzed in a molten-salt bath to produce lead metal and chlorine. The chlorine is used to regenerate ferric chloride in the leaching solution. The reactions involved are:


The low melting point (327° C), of lead, its high electrochemical equivalent, and the availability of stable low- melting electrolytes make the development of a molten-salt electrolytic process possible. Lead chloride is a favorable starting material because it is not hygroscopic and has good electrical conductivity when molten.

How to Make Lead Metal by Molten Salt Electrolysis

Molten-salt electrolysis of lead chloride is an integral unit operation in a ferric chloride leaching process that was developed by the U.S. Bureau of Mines for treating galena concentrates as an alternative to smelting. Prior to the Bureau’s work, several other investigators had studied molten-salt electrolysis of lead chloride. As part of the Bureau’s research effort, several monopolar and bipolar electrode designs were investigated in bench-scale electrolytic cells, which ranged in capacity from 5 to 400 A. The Bureau also built and operated a 3,000-A lead electrowinning cell that was tested at 450° C with a LiCl-KCl-PbCl2 electrolyte as part of an integrated, semi-continuous operation of the process. The 3,000-A (fig. 1) cell used horizontal graphite plate electrodes and was capable of producing 225 kg of lead metal per day. The anode consisted of two graphite plates 37 by 61 by 7.6 cm thick. The bottom surfaces of the anode plate were grooved with six 0.95-cm-wide channels which were sloped at 4° from the horizontal to direct chlorine gas away from the anode surface. The cathode was a single graphite plate measuring 74 by 61 by 5.1 cm thick. The cathode had six 0.64-cm wide grooves slanted to direct the

Precipitation of Mercury in Cyanide Gold Leaching

Numerous low-grade gold-silver ore deposits are being mined and milled throughout the Western United States. In addition to gold and silver, many deposits contain as much as 15 ppm of mercury. During cyanidation, 10 to 30 pct of the mercury and 85 to 90 pct of the gold are typically solubilized. The reactions for gold and mercury extraction with cyanide are —

2 Au + 4 CN- + O2 + 2 H2O → 2 Au(CN)2- + 2 OH- + H2O2………………………………………….(1)

2 Au + 4 CN- + H2O2 → 2 Au(CN)2- + 2OH-………………………………………………………………..(2)

Hg²+ + 4 CN- → Hg(CN)²4-…………………………………………………………………………………………(3)

2 Hg + 8CN- + O2 + 2H2O → 2Hg(CN)²4- + 4 OH-……………………………………………………….(4)

Mercury builds up in the recycled leach solutions because only part of the mercury is adsorbed on carbon in the loading circuit. Mercury and gold are stripped from the carbon with caustic cyanide solution and electrowon on steel wool cathodes. Mercury must either be recovered or precipitated because of the health hazard during smelting of cathodes and regenerating of activated carbon. One gold mill recovers mercury by retorting the cathodes prior to smelting. Another mill autoclaves the ore to extract minimal mercury. Soluble mercury can be precipitated with sulfides, as

Electro-depositing PGM Platinum Group Metal Coatings

The Bureau of Mines has evaluated the substitution of platinum-group metal coatings for bulk platinum-group metal objects as a means of reducing the consumption of the platinum-group metals. The Bureau has conducted several studies of the electro-deposition of the platinum-group metals from molten alkali metal cyanide baths during the last two decades. The major incentive for this work has been the need to protect materials from increasingly hostile environments imposed by modern technology. To withstand high-temperature environments, structural materials must possess high-temperature strength as well as resistance to oxidation and corrosion. Refractory metals such as molybdenum, tungsten, and columbium, and alloys of these metals that have the required high-temperature strength, are readily oxidized. Protection of the refractory metals could be accomplished by coating them with a suitable platinum-group metal, and a composite material with highly desirable properties would result. Bureau research has focused on producing platinum-group metal coatings on refractory metals as well as on more common materials of construction. This research has shown that high-quality deposits which are thick, adherent, and coherent can be prepared of each of the platinum-group metals, as well as select binary alloys of these metals.

Historical Perspective

In 1937, Atkinson obtained a patent describing

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