In recent years the metallurgical field of the copper industry has expanded greatly, the copper ores have become lean and diverse in character, and we are obliged to treat such ores on a very large scale.
On the commercial side, the operating and consulting staffs of the great mining companies have been enlarged and organized, and no great capital expenditure is decided upon until it has been investigated and recommended by a highly organized force of specialists as a committee.
I believe that in the scientific field and in the most valuable proceedings of our various societies the old idea of one-man point of view and one-man treatment of principle and detail together has been splendidly developed and should continue, but for the advancement of our art we should deliberately discuss subjects among ourselves in order to do away with the idiosyncrasies of the individual and to gain a composite view representing the best thought of many men.
In the development of new ideas leading to the introduction of new methods, the societies could enlarge their present great field of usefulness if such discussion be had, and in my opinion further great good could be done if a number of men of different qualifications and experience would join in preparing a joint book, each dealing with his particular problem, but consulting the other men. In this way individualism would be preserved and various processes would be described, and if properly reviewed the paper would be palpably capable of analysis, and the student would have at his command the associated efforts of a group of qualified engineers.
I have neither the time nor the qualifications to make an investigation and study in detail of hydro-metallurgical processes, but for some years I have foreseen the development of such processes, and I have a deep interest in the subject and have given it in some of its phases careful consideration from a long-distance viewpoint. I have assisted in directing many experiments and in reviewing the merits of many processes, and am glad to make some introductory remarks with the understanding that other gentlemen present will freely and candidly enter into the discussion and help to fill out the few bare bones that I can uncover. You will see in my remarks that I am controlled very largely by operating demands. No matter how beautiful the process, simplicity and cheapness of operating costs are the main features that appeal to me, and complexity and increased costs only appeal to me if they definitely promise an increased yield that will more than offset the increased expense; for, given the grade, the maximum profit per ton is demanded.
As you all know, there has been a very large development of deposits of lean copper ores in the Southwest in recent years, and this region has rapidly developed until in the aggregate it produces more copper than any like area in the world, and it still has a large increase of output ahead of it. The vast mass of these lean deposits consists of sulphide ore, and until recently these sulphides alone have attracted serious attention. It is well known, however, that there are large tonnages of oxidized copper ore associated with the lean porphyries.
A few years ago the Calumet & Arizona Mining Co., upon the recommendation of J. C. Greenway, took an option upon a majority of the stock of the New Cornelia Copper Co., whose mining properties are at Ajo, in southwestern Arizona. Mr. Greenway directed systematic drilling upon this property, followed by the sinking of some 70 shallow shafts for test purposes and the sinking of two deeper shafts, with considerable underground development, for the purpose of testing and checking the grade of the sulphide ore. The stock was purchased.
The matrix of the ore is an eruptive granite. There is no overburden. The chalcocite has been oxidized nearly as fast as formed. The zone of transition from oxidized minerals to primary sulphides is narrow. The oxide and sulphide ore is remarkably uniform in grade. Ira B. Joralemon’s paper, in the Transactions, describes this deposit. He shows that there are about 40,000,000 tons of ore containing 1½ per cent, copper, and the tonnage is dividend into about 12,000,000 tons of 1½ per cent, oxidized ore and 24,000,000 tons of sulphide ore, with about 4,000,000 tons of mixed oxide and sulphide between. If the grade be dropped to an average of less than 1½ per cent., these tonnages would rapidly increase.
In parallel with this development, oxidized ore in notable tonnages have been found in the Globe mining district. This ore is more scattered, and, on the whole, is leaner than the Ajo ore, and frequently has a large capping of overburden. In the aggregate, the Inspiration Consolidated Copper Co. alone now has over 15,000,000 tons of oxidized ore running about 1.3 per cent., and a probable tonnage in addition. There are other mines in other districts containing oxidized ore, notably in Utah and Montana, and there are also oxidized ores at other points in Arizona and a vast deposit in Chile. The treatment of such tonnages is of prime importance.
The Ajo ore has a matrix of granite with a large percentage of secondary quartz replacing feldspar. There is little calcium present in soluble form, and the oxidized ore seems to be best adapted to a leaching process. Preliminary laboratory tests showed that the ore crushed with the production of remarkably little slime and that the copper will dissolve in dilute sulphuric acid quite freely when the crushed ore contains fragments no larger than 6-mm. cubes. Tests also showed that a little cuprite is present, which is only partly soluble in sulphuric acid. There are also other soluble salts, notably salts of iron and aluminum. It was decided to investigate a leaching process for the oxidized ore, and since the ore made but little fines and could be leached with comparatively coarse crushing, we decided that we would try to leach in tanks by percolation. The amount of silver in the ore is small. Salt is apparently expensive, and it was decided that, other things being equal, the use of chlorine compounds as solvents was not desirable. I know, of course, that chlorine compounds may have a great future in leaching, but in this particular case, and in the early development of a process, it was considered that sulphuric acid offers fewer difficulties as a solvent.
To me the hydro-metallurgical problem divides itself naturally into two nearly separate divisions—the solution of the copper, and the recovery of the copper from solution. We were obliged, however to study the problems of solution on the basis of getting the copper into solution in the form most convenient for its ultimate recovery.
At the start we employed Stuart Croasdale to undertake a study of the solution of the copper, and we also delegated to him experimentation upon the precipitation of copper by iron. Mr. Croasdale ably conducted a long series of most valuable experiments and published a paper upon this subject which you will find in the Transactions of the Institute. At the same time we were fortunate enough to obtain the advice of Utley Wedge upon the manufacture of sulphuric acid direct from calcining-furnace gases.
Mr. Croasdale demonstrated that the ore crushed to an 8-mm. cube, or smaller, would yield over 80 per cent, of its copper when treated with a dilute solution of sulphuric acid. He demonstrated that slime presented no difficulty if the ore was distributed in the tanks uniformly, and that by finer crushing 85 per cent, of the copper could be obtained in solution if treated for a sufficient length of time with an acid solution. He demonstrated not only that the ore was permeable uniformly, but that with comparatively small tanks the height was not a matter of great consequence, and he was easily able to use a column 12 ft. in height, and could probably use one much higher. Later on we found upward percolation preferable.
Mr. Wedge showed us that with care in calcining the sulphide fines for reverberatory furnaces at the Calumet & Arizona smelter at Douglas, sulphuric acid of 55° to 60° Be. could be manufactured with a low operating cost, and a construction cost of not to exceed $3,000 per ton of daily capacity. The point of manufacture is about 300 miles, more or less, according to the route, from the mine. Mr. Wedge also undertook a study of sulphating and chloridizing roasts. His requirements were that the ore be crushed to pass a 1½-mm. screen. He showed that if the sulphide ore could be mined in parallel with the oxidized ore, and in proportion to their respective tonnages, there would be sufficient sulphur for his purposes and that a very high extraction could be obtained. He also showed that sufficient advantage could not be had in using salt to offset its cost and the difficulties that might arise later. Mr. Wedge also showed that a small amount of heavy sulphides from the Bisbee mines crushed and mixed with the oxide ore would give a very high percentage of extraction after the mixture had been subjected to a sulphating roast. He also introduced the idea that a preliminary water leach would give a perfectly pure sulphate solution from which the copper could be recovered electrolytically and that the acid thus regenerated would be sufficient to complete the leach.
Frederick Laist in the meantime was conducting a number of experiments governing processes in a like brilliant way, and among these processes was the development of an oxidizing roast on Anaconda tailings. Like Mr. Wedge, Mr. Laist found that by calcining sulphides or mixed sulphides and oxidized ores a high percentage of the copper could be extracted by leaching. Temperature control was vastly important, as the developments of both Messrs. Wedge and Laist showed. If too high a temperature is used a large amount of the copper becomes insoluble in the strongest acid, but in both cases temperature control was demonstrated to be simple and practicable, and under these circumstances practically no insoluble copper compounds were formed. In Mr. Laist’s case, however, the use of salt is essential on account of the silver contents of the charge, but he finds that it is not necessary to use this in a chloridizing roast and that a solution of salt with sulphuric acid gives him the desired extraction on the product from the oxidizing roast.
The manufacture of sulphuric acid by the chamber process, while simple, requires a very large investment, and in this case the transportation charges would amount to more than the cost of the acid. The chamber and the contact processes, of course, are the two important manufacturing processes, but even a casual study of wet methods of copper recovery shows that there are other methods of acid manufacture involving a principle similar to the chamber method, that might be used in parallel with the recovery of the copper from solution with less investment for plant and with a saving in energy that is important.
The McKay process includes a sulphating or an oxidizing roast, as he chooses, to get sulphide of copper into soluble form, and he adopts the same process for this purpose as followed by Messrs. Wedge and Laist.
The Slater process and the Midland process deserve study. They depend upon the use of complex series of chlorides and hypochlorous compounds that will dissolve many salts of copper that sulphuric acid will not touch. They will attack, among other things, chalcocite, and they deserve serious consideration in many important but special cases, as they avoid the necessity of a preliminary roasting operation.
I regret that I am only slightly familiar with the very important work in the development of a process by E. A. Cappelen Smith for the beneficiating of the Chuquicamata ores.
Utley Wedge, Ardmore, Pa.—It has certainly been very gratifying to hear such a comprehensive discussion of the development of the technique of leaching in connection with the property of the New Cornelia Copper Co. The presentation of this matter in the paper as read I believe to be accurate, but one aspect of the problem has not been presented. The paper relates almost exclusively to the treatment of carbonate ore. The deposit at Ajo, where this method is to be applied, comprises about twice as much sulphide ore as carbonate ore. It also comprises 1,500,000 tons of ore which is a mixture of sulphide and carbonate ore. It would be interesting to have accurate data as to the yield or percentage recovery of copper that can be secured from the sulphide ore by the various methods of concentration. The character of the ore leads me to believe that the percentage of recovery of values from the sulphide ore by any of the wet concentration or flotation processes in use at the present time, will be materially less than the percentages which this paper has shown us it is possible to recover by a leaching process. In considering the ultimate beneficiation of this property, I would suggest that the treatment of the carbonate of copper ore should always be considered in connection with the treatment of the balance of the deposit.
Is it not true that the sulphuric acid leaching plant for carbonate ore as described, if constructed on a large scale, will have largely outlived its usefulness when one-third of the deposit of ore has been worked out? If the same facilities that will handle the carbonate ore would also handle the sulphide ore, so that the same plant and the same processes could be utilized throughout the operation, the eventual total cost of construction might be materially reduced. The leaching of carbonate ore of this character with sulphuric acid is clearly demonstrated to be a success. It is clearly demonstrated that the carbonate ore which is to be leached with sulphuric acid does not need to be finely ground to leach completely and economically. It is apparently demonstrated that with processes recently perfected SO2 can be introduced into an impure electrolyte with advantages in economy as regards consumption of power in electrolysis and quality of product, but these advantages can largely be retained by the following operation:
Let the sulphide ore and the ore which is found between the mass of carbonate and the mass of sulphide ore, which partakes of the character of both, be ground to 20 or 30 mesh and furnaced at a temperature not much above a red heat, and the material can at once be leached with as good percentages of recovery as have been secured in the case of the carbonate ore. The copper solutions secured can be electrolyzed, and sufficient sulphuric acid can be produced from the electrolysis for use in leaching all or a portion of the carbonate ore, which methods can be identical with the methods described in the paper.
The leaching of the finely ground sulphide ore will call for somewhat modified appliances on account of the fineness of the material, but it has been abundantly demonstrated in other plants and in other camps that finely ground material can be economically leached. There are methods of handling finely ground material in leaching which even seem to have advantages in operation as compared with the methods described for leaching material in lumps.
In case the leaching plant projected, as indicated by the paper read, is built no larger in capacity than would in any event be necessary in connection with the treatment of sulphide ore and carbonate ore by the methods indicated, then and to that extent the construction for such acid leaching of the lump material would be serviceable for the life of the property, if the construction is of such character as to last that long, and if the mining and beneficiation of the carbonate ore can be so arranged that it will extend over a like period.
I am of the opinion that sulphatizing roasting of the sulphide ore in the orebody of the New Cornelia Copper Co., coupled with the acid leaching of all or a portion of the carbonate orebody with sulphuric acid by the methods indicated in the paper, will give the lowest possible cost of operation and will give the highest possible recovery of values of any method known to the metallurgical art. This combination of methods would obviate the necessity of construction of a sulphuric acid plant; in the place of this plant there would be constructed a furnace plant for giving the required roast to the sulphide ore. At the start, with only carbonate ore accessible, enough of the carbonate ore could be roasted mixed with small percentages of iron pyrite from neighboring camps to sulphatize the carbonate of copper.
I believe the cost of a roasting plant for conducting the process in this way is amply justified by the large recovery of values which can be made, and I believe that the advantages of this combination of methods are so great that the beneficiation of this property by such methods would yield several million dollars in excess of what would be secured if the sulphide ore were treated by any of the known concentrating processes followed by the smelting of the concentrate.
As a general statement, the class of copper ores known as oxidized ores and carbonate ores, to the extent that they are free from alkaline earths or other gangue materials that are readily attacked by sulphuric acid, will, in general, lend themselves to the treatment that Dr. Ricketts has described: namely, the treatment with sulphuric acid. But the past history of the leaching method as applied to the metallurgy of copper is full of instances where efforts have been made to accomplish results by leaching processes not preceded by the study which was given to this matter in connection with the New Cornelia operation. Such attempts in the past have justified the criticism of leaching processes which was once made in a general way by the father of a gentleman whom I see present, who said that the chief thing he had against leaching processes was that they were so Eine Schweinerei. If I am asked to give the English equivalent of that, I would say that it is apparent that the installation of leaching processes which the German gentleman had seen did not appeal to his esthetic sense.
But another day is very rapidly approaching, and the advance that has been made in the past few years in this line of progress seems to me to be revolutionary. There is so much being done by so many different companies that the efforts of any individual in connection with such practice must seem small.
The connection that I have had with this matter had led me to emphasize especially the effort to secure pure solutions which would yield to electrolysis without difficulty; and the development which has been brought about by two gentlemen who will be heard from this evening makes it seem probable that solutions quite impure in character will hereafter be economically electrolyzed. I regard the work of these gentlemen as of the utmost importance. The application of leaching processes to copper ores and copper calcines is practiced on a larger scale than is generally known. Over 1,000,000 tons yearly are treated for the extraction of copper by leaching. The largest leaching operation in the world is one operation which treats about 750,000 tons per annum. I refer to the Rio Tinto Co. The process used by the Rio Tinto Co. is in its essential character a sulphatizing roasting process. The ore is piled up in heaps 30 ft. deep, which are called terreras, and the ore is allowed to heat by chemical action until a sulphatizing roasting condition is produced and sulphates of copper are formed, which are leached out. The balance of the leaching of copper ores, done on a large scale throughout the world, is mostly chloridizing leaching, where salt is used. The leaching process with salt is very extensively carried out in Germany and England. There are several hundred thousand tons, I think, chloridized every year in Germany and England. I know individually of quite a number of very large operations there with which I have come in contact in connection with the installation of furnaces for chloridizing roasting. There is a very considerable tonnage chloridized in this country. The chloridizing process is familiar to a great many here. For the benefit of the others, I will say that in general, ore, which has previously been roasted to deprive it of most of its sulphur, down to a point where the sulphur is from one and one-quarter to one and one-half times the copper content, is mixed with a percentage of salt, usually in good practice in the neighborhood of 10 per cent., and this mixture, which is reduced to a fineness of 3 mm. by grinding, is then heated in chloridizing furnaces to a dull red temperature, or even slightly below, and the copper is chloridized and subsequently leached out with water, and with such hydrochloric acid as is caught in the scrubbing towers in connection with the furnaces.
The use of sulphur dioxide in connection with the precipitation of copper is a matter which personally I have been watching with very great interest. The quantity of sulphur gas which is escaping from the different smelter plants in this country is so prodigious that it seems like a waste; it seems as if this material will eventually be utilized and turned to some account. It is very possible that to some extent this will be brought about in connection with either the electrolytic or other precipitation of copper, where sulphur dioxide is evidently going to be very much in demand. There has been a possibility, if not a probability, that liquid SO2 might be required in this connection. It is interesting to note that at the present time, as far as I know, there is no liquid SO2 plant in the United States. There are several in Germany, and one or two in England, and the stoppage of the importation of liquid SO2 from Germany has advanced the price to about 10c. a pound at the present time. This certainly does not look attractive as a metallurgical proposition, but the possibilities of the cheap production of liquid SO2 to permit of its being shipped considerable distances for use, are such that I think it will come to pass in this country in a comparatively short time.
E. A. Cappelen Smith, New York, N. Y.—Our plant is practically finished; the leaching plant itself, as well as the mining plant, is practically completed. We have been slightly delayed with the completion of the power plant on account of the war in Europe—our power plant is being manufactured in Germany; according to our present plans, we expect to have the plant in operation by the first of April. So far in the construction work everything has gone on very nicely, and it now remains only to start the plant.
The general plan of work consists in steam-shovel mining the copper deposit followed by leaching the ore. From the crushing plant the ore will be moved by a belt and unloaded by a bridge that straddles the leaching tanks. The leaching tanks—six large concrete vats set end to end—are each 160 ft. long, 110 ft. wide, and 16 ft. deep. Each tank is designed to hold 10,000 tons of ore. After the ore has been delivered to these by the traveling bridge from the main belt, the first solution is applied, by upward displacement. At first we had planned to conduct the operation entirely by downward percolation, but we found a slight tendency of the finer material to clog the filter bottom; also, we found that channels were formed in the charge. These difficulties were entirely overcome when we instituted the method of applying the liquor by upward displacement. The first treatment solution stands on the ore for 48 hr. This solution is immediately followed by the washing liquors in rotation, run on the top of the ore, and displacing the previous solution by the piston method. In this way, we find that we can wash the ore so that the tailing will contain only from 0.015 to 0.03 per cent, soluble copper. We think this is about the limit, and we can obtain this result without introducing any more water in the process than sufficient to make up for that carried off in the tailing as moisture, and in the solution which is discarded to get rid of the surplus acid produced from the ore.
The solution is removed from the tanks through lead-lined iron pipes, and conducted to a central pumping station, where we will use centrifugal lead-lined pumps, each with a capacity of 5,500 gal. per minute against a lifting height of 62 ft. The pumps have been built by the Worthington company; they have been tested in the manufacturers’ shops, and as far as we can tell they are going to be highly satisfactory.
The solution is then delivered to the storage tanks, placed at an elevation above the leaching tanks, and from there it flows by gravity to the plant where we are going to eliminate the chlorine. We propose to pass the solution through cylinders 30 ft. long and 4 ft. in diameter, half filled with shot copper. By this means, the chlorine is precipitated in the form of cuprous chloride. The operation is continuous. The mixture of precipitated cuprous chloride, with the copper sulphate solution containing about 3 per cent, sulphuric acid, is run into Dorr thickeners,, where the precipitate is separated from the solution. The cuprous chloride pulp corresponds to between 1 and 2 per cent, of the total amount of the solution. We are installing some Kelly filter presses, which are lead fined, and we are also going to install an Oliver filter, likewise lead lined, in order to compare the two types.
The filter-pressed material will be delivered to the smelter by the usual methods, and at the smelter it will be mixed with limestone and coke finely crushed, (8- to 10-mesh material), put through a pug mill, pressed into briquets, and delivered to the blast furnace. In the blast furnace will be produced copper, and calcium chloride as a slag. The copper will be taken to a small reverberatory furnace; where it will be granulated and again used in the dechlorinators. We find that on account of the oxidizing agents present in the solution, partly in the form of ferric sulphate and partly in the form of nitric acid, the efficiency of the copper shot is only about 50 per cent, of the theoretical; in other words, there is a direct solvent action on the metallic copper in addition to that caused by the reduction of the cuprous chloride. Practically all of the copper produced in the form of cuprous chloride will go back into the process.
The solution, after the chlorine is removed, is fed by gravity into the electrolytic tanks with a strength of approximately 5 per cent, copper. It will run through the depositing tanks, arranged in sections of 16 tanks each, which will have about 13,000 amperes to a tank. The voltage will be about 2.3. We expect a yield of approximately 1 lb. of copper per kilowatt-hour. The solution will leave the last tank containing about 1.5 per cent, copper. We find that the current efficiency remains high down to about 1.5 per cent, copper; if we go lower than that, the current efficiency will drop off rapidly. By holding to these limits we will have an ampere efficiency of approximately 90 per cent. The solution from the last tank in the cascade will be delivered back to the leaching plant to be used over again.
Our problem is somewhat unusual in this respect, that we have available sulphuric acid present in our mineral, and instead of having to look around for means of producing acid, we have to look for a method of getting rid of the surplus sulphuric acid. We will have to dispose of a quantity of the solution on account of its content of sulphuric acid.
There are no soluble impurities in the ore itself, such as antimony and arsenic. Soluble lime and alumina and alkali salts are negligible. Consequently, the usual trouble experienced in connection with leaching processes is not present in our problem. Chlorine, which we have in the upper part of the orebody, was really our most annoying difficulty, but the treatment of the solution with metallic copper has eliminated this difficulty. In addition to this, we also have a small amount of nitric acid in the form of nitrates. So far we have not been able to determine that any especial difficulty arises in the process itself from the presence of nitric acid. It has, however, a rather deleterious effect on lead connections, lead pipes, and lead anodes. We have noticed, however, that this action occurs only at points where we are handling the strong solutions—those containing 8, 9, or 10 per cent, sulphuric acid. Nitric acid is present up to 0.8 per cent., but outside of the difficulty mentioned, it does not give any trouble. Its presence does not seem to have any particular function in the process itself, and we have not been able to determine that its being there causes any lower current efficiency. It does not seem to have any effect on the voltage.
Chlorine seems to have a tendency, which was to be expected, of lumping the deposit to a certain extent, and the use of colloids to prevent this does not seem to be very encouraging. I think it will be a question, more than anything else, of watching the deposit, and perhaps not allowing the cathode to get quite as thick as in ordinary practice.
Underlying the large oxidized orebody, at present developed to the extent of 200,000,000 tons, we have a sulphide orebody. Small-scale tests conducted on this material show that it can be treated either by roasting and leaching, or by water concentration followed by smelting in the usual manner. This orebody is also of enormous extent, but details have not as yet been decided upon for the final working of it, as the 200,000,000 tons of oxidized ore already developed will keep us busy for some time to come.
https://www.911metallurgist.com/method-precipitating-copper/
Raymond F. Bacon, Pittsburgh, Pa.—We have been in the field of copper investigation for some time, as many others have, but we feel it is somewhat premature to say very much as yet. We have been working on several different lines, and we believe no one leaching process will be developed that would be applicable to all ores or all conditions. There are certain leaching processes which are better adapted to the conditions of certain ores, and other processes are better adapted to the conditions of other ores.
As regards electrolytic processes, I think a good deal of the work we have been doing has been pretty well covered in different ways by the statements which have been made here to-night. Our work on electrolytic processes has been along two lines. First, to ascertain the optimum conditions for the electrolytic precipitation of copper from sulphate solutions obtained by the leaching of oxidized copper ores. In this phase of the work I may say that we finally obtained conditions under which we ran a solution over the ore through the precipitating tanks and then over the ore again, 20 cycles being completed, without any purification of the solution. There was precipitated on the average in these 20 cycles 1.86 lb. of copper per kilowatt-hour and the current efficiency on the twentieth cycle was fully as good as that on the first cycle. Our second line of work on electrolytic processes has been an attempt to find efficient and cheap depolarizers, so as to increase the amount of copper precipitated by a definite amount of electrical energy. We have found one very cheap substance which for short runs gave us 4 lb. copper per kilowatt-hour.
In regard to the sulphur dioxide process, we have been actuated by this thought—it is necessary in most ores to have quite an excess of acid to take care of mechanical losses and to take care of alkaline substances in the ore. One way of doing this is along the line which has been spoken of, by using the sulphur dioxide in an electrolytic cell. We, also, have used sulphur dioxide in an electrolytic cell, but all of our work has tended, to show that the efficiency of sulphur dioxide as an anode depolarizer in a sulphate solution is rather low. Because of this low efficiency, our calculations showed us that it is cheaper to precipitate directly with the sulphur dioxide under pressure than to use the sulphur dioxide as an aid to electrical precipitation; that is, the heat required for direct precipitation with sulphur dioxide will be cheaper in most localities than the electrical energy required either for straight electrical precipitation or for the combination in which sulphur dioxide assists the current. We all appreciate the advantage which any electrolytic process has of giving directly cathodes of electrolytic copper. On the other hand, it has seemed to us that for most localities and most conditions the disadvantages of electrolytic processes, as at present developed, outweigh this advantage. Some of the disadvantages of electrolytic processes are these: The fouling of solutions immediately and seriously cuts down the current efficiency. Changes in the character of the ore, particularly as regards soluble impurities, are always to be expected in large orebodies. The electrolytic process is very sensitive to changes thus brought about in the electrolyte. The installation and upkeep costs of the electrolytic precipitation process are very high and to obtain efficiency requires constant supervision by high-priced men. One rather fundamental objection to electrolytic processes has not been discussed here to-night. In electrolytic processes, at least of the type which have been spoken of to-night, the copper content of the electrolyte can be reduced in the precipitation cells no lower than 1.5 per cent, copper. That means that it is not easy to so conduct a cycle of leaching, precipitation, etc., that all of the water-soluble copper shall be removed from the ore and the volume of the solution in the cycle shall be kept constant. For example, let us assume in a systematic leaching process that all the copper has been extracted from a ton of ore and that that ore is then in contact with the solution which in the cycle is strongest in acid, having just come from the precipitation tanks, and which, of necessity, still contains 1.5 per cent, of copper. This one ton of ore is drained as thoroughly as possible and, according to our experience, then contains from 40 to 100 gal. of this solution containing 1.5 per cent, copper. In other words, there is from 6 to 13½ lb. of copper still in this ore. By adding the proper amount of wash water to retain the volume of solution in the cycle constant—that is, in this particular case from 40 to 100 gal. of wash water—the copper content of that ore will be reduced to from 3 to about 7 lb. To wash out further this soluble copper, so as to have a negligible quantity of soluble copper remaining in the ore, according to our experience, requires an amount of wash water such as to increase very appreciably the total volume of solution which is in the cycle. It is obvious that in any process in which there is practically complete precipitation of the copper, as in Weidlein’s sulphur dioxide process, this washing difficulty does not arise. In place of losing soluble copper in the ore, one there loses a corresponding amount of sulphuric acid; and in Weidlein’s process sufficient sulphuric acid is constantly being formed, so that this loss can be taken care of. As to the element of danger in processes in which a solution is heated under pressure, I would call your attention to the fact that there are thousands of boilers operating continuously in the United States under pressures considerably in excess of those used in the Weidlein process. We built in Nevada a 40-ton plant and we have had certain mechanical difficulties, as we expected to have. These difficulties we are meeting by the installation of a continuous process of precipitation. The advantages of continuous precipitation will be evident to all mill men. Our expectation that the cost of heat in this process of precipitating pure copper is less than the cost of electricity for electrical precipitation has been borne out. Of course, I recognize that there are localities where electrical power is so cheap that an electrolytic precipitation process should be the one to be considered, but I believe that for most localities in the western United States some of these other processes are more favorable as regards cost. In the Weidlein process, a very large excess of acid is constantly regenerated and all who have had experience in the leaching of ores know how desirable this is. In that connection I will say that we have developed a system of handling sulphur dioxide and a method of concentrating it from dilute flue gases. We believe we have made noteworthy progress in that direction, but I am not prepared at this time to go any further into that. In a short time we will be able to say a little more on the subject.
https://www.911metallurgist.com/leaching-copper-iron-electrolyte/
J. Parke Channing, New York, N. Y.—The problem presented in the Miami district is a little different from that in the districts referred to in the discussion of this evening, inasmuch as the ore to be treated there is one which is a mixed chalcocite and carbonate. The gangue, as you know, is particularly free from any impurities that go into solution, but the fact that both chalcocite and carbonate are present makes it a rather difficult problem. The particular ore deposit upon which we worked was that of the New Keystone, but the same class of ore existed in the Miami, and also on the Inspiration. The problem therefore presented to us was either a straight leaching one, which would have to be proceeded with by complete dead roasting, or else it would be a mixed process, water concentration followed by leaching, or possibly leaching followed by water concentration; and then finally, if flotation came in, we might have to make three bites of the cherry—that is, use water concentration, flotation, and leaching. You can readily see there would be a great many combinations; and whether it would be desirable to leach first, then water concentrate, and then float, or whether it might be desirable to make the sequence different, we do not know. This Keystone ore averages about 2¼ per cent, copper, of which anywhere from 30 to 40 per cent, is oxidized. After making a number of concentration tests we decided upon attempting a dead roast; in other words, converting all to the copper into a soluble form and leaching it.
It is not possible for me to give you the details of the numerous different things we tried. I may say, however, that for the purpose of precipitation, we tried the electrolytic method. The problem with respect to these mixed ores is by no means solved and we are still conducting experiments along these lines, and I think Mr. Canby, the gentleman in charge of the experiments, can give you in a general way some idea of what we have accomplished and what is still left undone.
R. C. Canby, Wallingford, Conn.—The only thing I can think of which has not been mentioned is the circulation of the electrolyte. In our electrolytic work I designed the cells so that the solution, instead of flowing lengthwise against the flat surfaces of the anodes and cathodes, flowed crosswise of the cell. In that way, in the use of sulphurous acid gas, I had hoped, by having a rapid circulation, to use a much smaller quantity of both copper and the depolarizing element in the electrolyte, with a correspondingly stronger current, and also to be able to precipitate the copper without heating the electrolyte. I was able to use a current density as high as 10 or 12 amperes per square foot with the temperature of the electrolyte about 18° to 20° C., and get a very satisfactory coherent deposit. I think that was the only novelty in the electrolytic work.
The other things were done very much along the lines which have been discussed here this evening. We think, however, it would be very much better if all of the oxides could be treated in the present plant, and we are now working to that end, so as not to have a separate installation for the semi-oxidized ores. It is along that line that experiments are now being carried on, so as to treat the oxidized material, which is mostly in the flotation tailings. I think there were perhaps five or six different methods of manipulation tried, such as leaching resulting slimes and leaching before crushing for concentration, etc., but all of these combinations contained really nothing of any special interest, as they are thoroughly familar to everybody.
J. Parke Channing.—How about roasting with the oil furnace— would that be of interest?
R. C. Canby.—That might be of interest. The problem was to use our 14° Baume fuel oil, to be burned at a temperature of 2,000°, and yet keep our ore bed at not to exceed 800° to 1,000° C. I accomplished this by using a furnace similar to that used for the Huntington and Heberlein process at El Paso, and at three other American smelting plants, the hearth revolving past the flame, so that I was able to roast the ore without at any time bringing the temperature of the roasting ore above the desired 1,000° C., whereas the fuel was being burned at about 2,000°, or a little over. The furnace was entirely of reinforced concrete construction and stood well.
Richard Lamb, New York, N. Y.—During 1907 I designed and erected a leaching and electrolytic copper extraction plant at High Hill, Va., on the Virginia Copper Co.’s property, to treat the bornite and chalcocite ores found there.
Owing to the large percentage of silica this ore is difficult to concentrate economically. The large bulk of silica crowds off the values on any form of concentrating apparatus, and at least 50 per cent, of the values is lost if machines are worked at a speed necessary to make them efficient. There is a large body of copper ore in the Virginia copper belt, but many failures have been made in the attempt to concentrate it. However, the ore is ideal for leaching. The magnesia and alumina are mostly in silicate form, and do not go into the acid solution, and the large percentage of silica is in its favor when leaching.
I built the first experimental plant for treating this ore in Hoboken, and first tested leaching the ore in open tanks. In crushing this ore to say 10 mesh at least 25 per cent, goes into slime. This adds to the difficulty of leaching if the usual leaching tanks are used. I therefore substituted a chlorination barrel for the leaching tank in order that the crushed ore might be stirred while leaching. Stirring arms in open leaching tanks when using sulphuric acid disintegrate quickly. The lead-lined barrel used stirs the ore without injury from the acid. As the ore is a sulphide it has to be roasted. In the Hoboken test plant I used a small Bruckner roaster. In the large test plant at the Virginia Copper Co.’s mine I used an Edwards roaster with a capacity of 30 tons of ore per 24 hr. I was able with this roaster to control the heat. While we had only 1.5 per cent, of iron in the ore, to avoid its polluting the electrolyte I sought to change it from an oxide to a sesquioxide of iron, which is insoluble in dilute sulphuric acid, by keeping the heat while roasting to about 900° F. Dr. James Douglas told me that he had succeeded in changing the iron to a sesquioxide of iron in a copper ore he was leaching, having over 25 per cent, of iron in same, by roasting with the proper degree of heat. I found that if the ore is crushed only to a size that exposes the sulphide on the surface, the roast will be as good as it would be if the ore was ground exceedingly fine. There is some oxide and car-bonate of copper ore in the form of cuprite, malachite, melaconite, and green carbonate produced, but these are not separated when coming out of the mine, so they were sent to the roaster, although the roasting was superfluous, as sulphurous acid will take the copper of those oxide and carbonate ores into solution without their being roasted. The ideal roast would be to change the sulphide ores into sulphate, but since this condition would require careful manipulating I sought to make the sulphide ore oxide and sulphate, but not to make a sweet roast, only seeking to leave as little of the ore in a sulphide condition as possible. If the ore could all be roasted to a sulphate condition the sulphur of the ore would go into the solution while leaching and thus add sulphur to the acid supply.
The ore, ground to 10 mesh and roasted, was put into a lead-lined barrel. A solution of sulphurous acid was made by burning sulphur in an inclosed oven with a supply of air sufficient in volume to supply the oxygen to produce sulphurous gas, and under a pressure just sufficient to overcome the hydraulic head due to the height of the water from the submerged outlet of the sulphurous gas delivery pipe to the surface of the water in the tank. The water is acidulated to about 10° Baume by absorbing the sulphurous gas.
This acid is pumped into the leaching barrel with the roasted ore, and the barrel is revolved until the copper has gone into solution. This takes about 3 hr. Compressed air is then passed into the barrel and the solvent is forced from the barrel through the filter in same. This filter is the same as is used in the barrels used in chlorination work. This solution entered a 20,000-gal. cypress stock tank, painted with Mogul acid-proof paint. The elevation of the tank was such that gravity would produce a flow of the solution or electrolyte from the stock tank through 16 lead-lined cypress vats, placed at elevations with reference to each other so as to continue the flow of the electrolyte through the 16 vats. An acid-proof pipe line was also provided, situated below the vats, so that any vat could be discharged or shunted. Each tank was provided with 1/8-in. thick lead anodes and cathodes, with exposed surface 25 by 36 in., in all 1,030, the cathodes providing over 6,000 sq. ft. of depositing surface. An overhead crane was provided to carry the anodes and cathodes to their respective places.
The electric current was supplied by two Holtzer-Cabot 30-volt, 1,200-ampere generators, with a 120-volt, 1.5-kw. exciter.
By the use of sulphurous acid the anodes were prevented from per-oxidizing. The sulphurous acid of this electrolyte is changed to sulphuric acid by the electrolysis and becomes a stable acid. After passing through the vats any loss of acid from being neutralized by the alkali of the ore is replaced by sulphurous acid, as before described, and the electrolyte is used over and over again to leach, more ore. When the electrolyte becomes polluted it is run into a larger vat having an extra number of electrodes so that the current density is very slight and the copper is practically depleted. The electrolyte from this hospital tank can be run to waste or further treated for its sulphur contents.
The voltage supplied to each tank was 1.8, and a current density of 3 amperes per square foot was used. The electrolyte was permitted to run from ¼ to ½ per cent, in copper, when used over for further leaching.
The output was 30.6 lb. of pure electrolytic copper per horsepower day of 24 hr. With water power at $12 per horsepower per year this would represent a cost of 0.12c. per pound for electrolytic treatment.
With steam power costing $48 per horsepower per year the cost for electrolytic treatment would be about ½c. per pound of copper produced. Producing copper in the Virginia copper belt, from underground to finished product, would cost about 6.9c. per pound working upon a scale of say 5 tons of copper per day. In my belief, using the large available water power nearby, and making 5,000 tons per year, the cost of copper will not exceed 6c. per pound. The capacity of the trial plant described was 2,800 lb. per 24 hr. In building this plant I was employed by one man, who alone, on the advice of his brother, an experienced metallurgist, undertook to lease the Virginia Copper Co.’s mine and put up the experimental plant. The plant was run only a few days, the owner having lost his money in the fall of 1907, and the plant was thrown into the courts, with disastrous results to the owner. Enough was done to demonstrate that with the above described combination of well-known and tried-out apparatus, using the cheapest of solvents, sulphurous acid, there is no reason why a suitable copper ore should not be treated on a large scale so as to produce electrolytic copper at a lower cost than by the old dry-process methods.
