Leaching Copper: Iron form Electrolyte

Leaching Copper: Iron form Electrolyte

The solvent we would use; how we would precipitate the copper from the solution, and what we would do with the iron in the electrolyte. Without going into the details of the reasons, we decided that sulphuric acid would be the most satisfactory solvent and that we would recover the copper from the solutions by electro-deposition. As to how we would take care of the iron in the electrolyte, we had to choose between controlling the iron (i.e., keeping it in the ferrous state) in the solution, and eliminating it from the solution. The removal of the iron appeared to offer the more attractive field.

Briefly, the method followed involves leaching with sulphuric acid, removing a portion of the iron from the solution, and precipitating the copper electrolytically,

Regarding the leaching per se, I do not think I could say anything which would be of any particular interest. There are certain factors which govern the extraction of copper from an ore, and when these factors are given due consideration, I think we can get about the same extraction with one process as with another.

In considering how to remove iron from the electrolyte, we were reminded of the work which had been done by Hoffman at Argentine, where he precipitated iron from bluestone solution. Hoffman accomplished this by heating the solution in a suitable tank, agitating with air, and adding copper in an oxidized form. I think he used a roasted lead-copper matte. This work of Hoffman’s gave us our idea for the removal of iron from the solution.

For the copper oxide necessary in this work we have used high-grade copper carbonate ore, oxide formed by the roasting of cement copper, roasted high-grade sulphide ore, also roasted copper sulphide concentrates. In our demonstration plant (10 tons daily capacity of carbonate ores) we used roasted concentrates. The concentrates were roasted in a small Wedge furnace, the results obtained being far better than those attained in our preliminary work when we roasted the same class of concentrates in a laboratory muffle furnace.

In order to obtain a satisfactory reaction between the copper oxide and the iron, etc., in the solution, it is imperative to have the oxide material very finely divided. The roasted concentrates we used were ground in a small Abbe mill. We found that with 5 or 6 hr. grinding we obtained a product 93 per cent, of which would go through a 200-mesh screen. It is probable if we had had time to investigate this grinding operation, we would have found that it was not necessary to grind for 5 hr., but, owing to the shortness of the period at our disposal for securing results, we did not have the opportunity to bring this operation down to the shortest time limit.

From the Ajo carbonate ore we removed a little more than 2 lb. of iron per ton of ore treated. In order to keep the iron in the solution below the percentage deleterious to electrical efficiency, it was necessary to treat approximately one-quarter of the solution. We found that by manipulating the solution going on and off the ore in a certain way we could concentrate a large proportion of the iron in a certain part of the solution. This was accomplished as follows: The first solution going on the ore carried about 2.5 per cent, of free acid and was kept on the ore by circulating until it was approximately neutral. This neutralizing precipitated from the solution oil the ore a considerable proportion of the iron, alumina, etc. This neutral solution was drawn from the ore, and after acidification by electrolyte coming off the electrolytic tanks was sent to electrolytic tanks for precipitation of the copper. As the first solution was withdrawn from the leach it was replaced by fresh solution. This fresh acid solution not only removed some more of the iron from the ore but also picked up the iron which had been precipitated from the first solution used. This second solution now high in iron was circulated until it was nearly neutral, when it was replaced by fresh solution carrying 4 to 4.5 per cent, free acid, which completed the leach. The total leaching period averaged about 80 hr. The solution high in iron, etc., was sent to that part of the plant where removal of iron and other impurities was accomplished.

In purifying the solution is was first heated to 195° to 200° F. When this temperature was attained agitation with air was begun and the finely ground oxide material was gradually added. Maintenance of a temperature of about 195° F. and agitation with air was continued for 3.5 hr. During the first hour all the ferric iron was precipitated. The removal of the ferrous iron was slower and we found that it was not advisable to endeavor to precipitate it all. In 3.5 hr. we easily precipitated 90 per cent, of the total iron and at the same time removed 65 to 75 per cent, of the alumina, 50 per cent, of the manganese, and all of the arsenic, antimony, and bismuth which might be present.

After the above operation the pulp was passed through a Schreiber wooden filter press. The solution came out beautifully clear and was sent to the high-acid electrolytic circulation system. We anticipated that with the precipitate produced we might have trouble with the filter cake; that it would be slimy and difficult to handle. We did not find it so. The cakes were 2 in. thick, firm, washed well, and the amount of copper in the form of unused oxide they contained was well within commercial limits.

As you are well aware, when the iron sulphate, aluminum sulphate, etc., react in the agitation tank with the copper oxide, they give up their acid radicals to the copper, forming copper sulphate, which in the electrolytic tanks yields sulphuric acid, so that the only loss of acid is that due to combination with the alkalies and that due to loss by entrainment in the tailing. This acid loss we found we could easily replace by a proper control of the roasting of the concentrates in the Wedge furnace; i.e., we could so control the roast as to maintain in the calcine sufficient copper sulphate either as water-soluble sulphate or as basic sulphate to compensate for the acid losses to which I have referred.

Regarding the electrolytic part of the work, we simply followed standard tank-house liberator-tank practice. Our anodes were antimonial lead, 4.0 per cent, antimony. We tried both grids and sheets. With the grids our voltage was 0.3 volt higher than with the sheet anode and we discarded the grids in favor of the sheets.

With reference to the use of a lead anode, I might say we commenced our work with some apprehension, fearing that there might be a serious loss of lead. We made a number of protracted tests and as a result we calculate that the cost for lead anodes at Ajo with lead at 8c. per pound would be less than 0.2c. per ton of ore treated.

In the process I have outlined there is really nothing new or novel. It simply consists of an idea picked up here and another there, and the putting of these various ideas together.

If we were to criticise the process, the pros and cons might be summed up about as follows:


  1. Simplicity.
  2. Every step in the process has been in use in commercial plants for years.
  3. Saturated solutions avoided.
  4. Wash water not excessive, and if it should become so, it can be taken care of.
  5. Impurities which would injure the quality of the electrolytic copper are eliminated from the electrolyte.
  6. Acid control very simple.
  7. No step in the process is dependent upon fine adjustments.


  1. High and low acids used. This necessitates more solution-storage tanks, more solution lines, and more pumps than if a single solution were used.
  2. There is an extra operation: namely, purification of electrolyte, which involves the installation of a roasting furnace, boilers, compressor, and filter press.
  3. The lead anodes will probably have to be changed at the end of four or five months, since the lead peroxide scale formed will increase the voltage.