Heap Leaching

Heap Leaching, as practiced at Rio Tinto, Spain, while one of the oldest, and probably one of the cheapest, methods of extracting copper from its ores, has not had, until recently, other than experimental application in this country. This has been due partly to a mistaken idea that the Rio Tinto ores possess some obscure and mysterious quality that renders them alone suitable to the process, partly to the fact that tests properly made meant the expenditure of large sums of money and several years time, and partly to the fact that favorable commercial and other conditions are not often found.

This paper gives some views on the chemistry and mechanism of the method, an account of the preliminary experiments, and the final plans adopted for a large-scale installation of the method by the Phelps Dodge Corpn. for treating low-grade ore from Sacramento Hill, Bisbee. If results comparable with those obtained from the test heap are to be had from the large plant, this plant will operate on the lowest-grade sulfide ore now being treated commercially in the country.

The method has several advantages: The installation cost is low. Interest charges must, of course, be made both on the cost of the ore delivered to the plant and for the plant itself, but as these are low, there is considerable advantage in the ability to regulate output by more or less complete shutdowns in accordance with market and other conditions. The amount of labor needed is small, and the cement copper produced may be considered nearer a finished product than the sulfide concentrates from a mill, and certainly means less installation cost per ton to put into final form by present methods.

At Bisbee, conditions were favorable for a trial of the method for, in the preparation of the Sacramento Hill orebody for steam shoveling, a large amount of material below concentrating grade had to be moved to a suitable dumping ground, so the cost of delivering this to leaching beds meant practically no extra expense. The only risk was the cost of the plant which, it was thought, could be amortized by an actual extraction sufficiently less than that obtained experimentally to make the proposal reasonably safe. Up to October, 1921, approximately 380,000 tons, from stripping operations, averaging 0.92 per cent, copper had been delivered to the first heap, but since that time, because of the suspension of stripping and other conditions, no deliveries have been made, and plant construction has been postponed. That part of the paper, therefore, dealing with proposed plant construction represents recommendations and plans made up to that date; whether they will be followed will depend on the future policy of the corporation.

The plant was designed to take care also of the mine water formerly treated at the cementation plant at Bisbee; it embodies certain improvements locally worked out as the result of experience there, which are expected to result in economy.

Factors Involved in Heap-Leaching Process

While the process of heap leaching is a simple one to operate, so little is known accurately of the chemistry involved that it is reasonable to hope for modifications in operating methods that will shorten materially the time required for treatment.

Sponge iron, if it is produced commercially at a sufficiently low cost, will be a much more satisfactory precipitant than the tin cans and scrap iron now used; either this product or pig iron will have to be used eventually, for the present supply of tin cans and scrap is not sufficient to cover much extension of cementation in the southwest.

We can easily define heap leaching in terms of its flow sheet and operations, which are possibly simpler and require less expensive and complicated apparatus than any other method of getting copper from ores. It cannot be said, however, that we know exactly, or even approximately, the reactions in the heaps by which the copper is converted into forms soluble in water.

There is no doubt that the kind of copper mineral, its physical character, and the manner of its original deposition are important factors. We can be quite sure that some ores cannot be leached, if it is known that they are quite free from porosity. This follows from the fact that it seems certain that extraction of copper by this method depends on each piece of rock having pores, or microscopic channels, either open or filled with material that can be acted on by the solutions, and which will permit the solutions both to penetrate and to leave the interior of the piece. It would seem probable, therefore, that ores in which the copper is of secondary origin (that is, enriched by the precipitation of copper from solution in microscopic channels, pores or fractures) will be amenable to heap leaching; and that if an ore, in which the copper is of primary origin, is lacking in pores leading from the outside of a lump to each particle of copper mineral, it cannot be treated by the method. We are, however, dealing in reactions and operations that require at least several years; and while small-scale tests can be made to show definitely that an ore can be treated, negative results in a small way are not conclusive evidence of large-scale results because of the time factor and the difficulty of reproducing large-scale conditions.

From the foregoing, there can be deduced one necessary factor in the method of large-scale operation, namely, that each piece of ore in the piles should alternately be wet completely and dried thoroughly. If any individual piece having the necessary pores or channels and thoroughly dry, is wet on the surface, in a reasonable time each passageway, by reason of capillary action, will be filled with the solution wetting the surface, providing the channels are open at both ends. No reasonable amount of washing will remove the solution from the interior; but if the surface is dried, the reversal of the capillary action will bring the solutions to the surface of the piece, bringing with them any salts dissolved on the way. This action can be observed easily on the outside of a pile or an individual piece, which when dry becomes coated with quite pure bluestone.

If the foregoing is correct, it will give an idea of the amount of solution theoretically necessary for application at each cycle. Assuming that a pile or an individual piece of rock is dry, all that can be accomplished in the way of extraction by a single application of solution or water is to wash off the soluble salts on the surfaces, and no greater quantity than is required for this purpose need be used; a larger amount means simply more dilute effluent liquors. Also, as the volume of the pores is comparatively very small, no large amount of solution is needed for filling them; the amount of solution remaining on the surfaces after washing is probably enough. It is certain that complete immersion of the ore or the use of an excessive amount of solution is unnecessary and undesirable.

In heap leaching, it is probable that carbonates, cupric oxide, cuprous oxide, and metallic copper will be attacked in the order named. Copper silicate is an indefinite compound, with widely varying character and solubility, so no definite statement can be made with regard to it. It seems probable, although not certain, that if an ore carries any considerable amount of its copper as oxide compounds, a certain amount of sulfuric acid will have to be used.

Chalcocite and bornite are readily attacked and dissolved, but chalcopyrite is more refractory. It is probable that an ore containing chalcopyrite as its principal copper mineral will be much more difficult to treat by the method.

The reactions are exothermic, and where, as at Rio Tinto, heaps consist entirely of heavy sulfides, careful attention must be paid to keep the resulting temperature below the ignition point.

A very important matter practically is the character of the gangue and its behavior in the heaps under the influence of the solutions.

As the method depends on alternate wetting and drying to provide for penetration of solutions into a lump of ore and their subsequent removal, the size of lump and proportion of fines in relation to depth of pile are important. In addition, some rocks, under the action of the solutions and alternate wetting and drying, may crumble more or less completely. If the resulting products are coarsely granular, this crumbling is not serious; it may even be advantageous in allowing quicker and more complete penetration of solutions. But when a rock breaks down into very fine or claylike particles, it will be more difficult to leach and the uneven distribution of the effects of such breaking down will result in uneven leaching. Data are not available to show how an area in a pile consisting entirely of such completely broken down rock will behave. The first experimental test of the method at Douglas, however, consisted of 25 tons of sand tailings. This small heap allowed complete penetration of solution which, when the heap was allowed to dry, showed marked tendency to return to the outside surface by capillary action, bringing dissolved salts to the surface. The factor of ease of drying comes in here, however, and it is possible that the depth of pile with a rock that breaks down easily will have to be reduced; but so many factors are involved that at present the depth of pile is purely experimental.

The practice at Rio Tinto is of little help in this or other matters, as the heavy pyrite and the siliceous ores leached there vary widely in composition and behavior from our southwestern porphyries.

These statements concerning some of the physical factors involved show that the method is not so simple as the operation would indicate, and the same is true of the chemistry. It will also prove to be true that both successful operation and any possible improvements can only be had by as complete a knowledge of the physical and chemical factors as possible. This sounds trite enough, but frequently expensive failures have been made in apparently simple proposals through neglect of this obvious consideration.

Chemistry of Process

Turning to the chemistry of the process, we can be quite sure of the occurrence of some reactions. So-called oxidized copper compounds will be present in appreciable amounts in the ores under consideration. Cupric oxide, carbonate, or silicate will be probably acted on as follows:

  1. CuO + H2SO4 = CuSO4 + H2O
  2. 2CuO + Fe2(SO4)3 + H2SO4 = 2FeSO4 + 2CuSO4 + H2O
  3. 3CuO + Fe2(SO4)3 + 3H2O = 3CuSO4 + Fe2(OH)6
  4. CuO + FeSO4 + H2O = CuSO4 + Fe(OH)2

depending on the composition of the solutions and the amounts of ferric and ferrous iron, and free acid present.

Soluble basic constituents of the ore, other than copper, will be acted on similarly, resulting in reduction of acidity and precipitation of iron compounds. It is therefore evident that maintaining the balance of iron and acid in the solutions is important practically, and experiments have shown that some ores require the addition of extraneous acid to maintain this balance, while others do not.

Theoretically, so far as iron is concerned, the reaction

3. 3CuO + Fe2(SO4)3 + 3H2O = 3CuSO4 + Fe2(OH)6 will be exactly balanced by the precipitation reaction
5. CuSO4 + Fe = Cu + FeSO4

This may or may not be the case practically. Both free acid and soluble iron salts are produced in the piles by reaction of the sulfides present, and may be sufficient to make up the losses, caused mainly by the usual methods of precipitation, by which basic salts of iron are formed in considerable amounts.

In large-scale testing of the method at Tyrone, acid was added to the liquors, but at Bisbee this was not done.

Another important difference between Rio Tinto practice and the application of the method to our ores is that there will be a smaller quantity of waste liquors to be disposed of in normal operation. This is obvious, as iron is the leaching reagent and it is not produced in excess as at Rio Tinto.

We have assumed the delivery to the plant of 300,000 gal. of mine water per day, which has to be treated in any case, and the pumping capacity needed for the leaching heap has been based on using 1¼ gal. per ton per day. Treating 400,000 tons of ore for the first year would require 500,000 gal. of water for leaching per day. Soakage and evaporation is estimated at 50,000 gal. per day so that in order to equalize the water problem the following amounts will be needed per day (soakage is the amount of water taken up by the ore):

From these figures it appears that after the second year the amount of waste solution will not exceed about 13 per cent, of the total from the heap, and is less than the amount of mine water added to the system.

There will be an advantage in returning solutions from one part of the heap to other parts, in order to decrease the ferric-iron content of the effluent liquors before sending these to the cementation plant. This method should increase the extraction from the heap and decrease the quantity and increase the grade of the solutions. Furthermore, it should postpone the addition of further units of the cementation plant for the second and third year of operation for the quantity of solution to be depleted of its copper will be approximately the same. If this plan works out successfully, it will also delay the building of reducing heaps.

The reactions by which the sulfides of copper are converted into water- soluble sulfates in the heaps are not known with certainty. The equations given in the latest description of operations at Rio Tinto are as follows:

6. 33FeS2 + 28O + 4H2O = 29FeS2 + 4FeSO4 + 4H2SO4
7. 4FeSO4 + 4H2SO4 + 2O = 2Fe2(SO4)3 + 2H2O + 2H2SO4
8. Cu2S + Fe2(SO4)3 = 2FeSO4 + CuSO4 + CuS
9. CuS + Fe2(SO4)3 + H2O + 3O = 2FeSO4 + CuSO4 + H2SO4 or
10. 33FeS2 + Cu2S + 3H2O + 33O = 29FeS2 + 2CuSO4 + 3H2SO4 + 4FeSO4

showing the oxidation of pyrite to form ferrous sulfate and free acid; the oxidation in solution of ferrous to ferric iron; and the solvent action of ferric sulfate on Cu2S and CuS.

Similar equations have been published for years, but they do not really explain what takes place, and it is difficult to see why they should be written in the forms given.

It seems certain that an important reaction accounting for the solution of copper from a sulfide mineral is

CuS + Fe2(SO4)3 = CuSO4 + 2FeSO4 + S

but unquestionably other reactions take place, and it cannot be said that the chemistry of the process is known with any degree of certainty or completeness.

In the absence of direct evidence, a definite explanation cannot be given, but some of the observed facts point toward what follows.

We have to explain that starting with sulfide of iron and sulfide of copper, in the presence of excess of air and moisture, there are produced in the heaps ferrous and ferric sulfate, free sulfuric acid, and copper sulfate. The theory of capillary action seems quite clearly to explain how copper can be removed from the interior of a piece of ore without showing any macroscopic changes, and if the progress of leaching is followed by a proper microscopic study, this will probably be verified and the actual way in which the solutions get in and out followed. This has been done to a certain extent. It is simple enough to explain the formation of ferric iron from ferrous iron in the presence of excess of air.

The direct oxidation of pyrite by oxygen to produce SO2 and iron oxides is a reaction that starts slowly at quite low temperatures, but the velocity increases rapidly with rising temperature. This reaction may, or may not, have a measurable velocity at ordinary temperatures, but at the lowest temperatures at which it will occur it is probable that moisture will, as it does for many other reactions, act catalytically to increase the reaction velocity. A determination of the lowest temperature at which this reaction occurs appreciably should be made, and the influence of moisture on it; but as there is no evidence to the contrary and this explanation fits the facts it may be the correct one. As the reaction is strongly exothermic and as there is very inefficient radiation in the heaps, the heat produced at the beginning will accumulate and act as an accelerator; we would expect in a pile of closely packed pyrite with small amounts of moisture present, a constantly rising temperature.

This is the “heating up” that has been so often described for Rio Tinto. This accumulation of heat may be sufficient to start the heaps burning, to prevent which, care must be exercised. This is exactly what we would expect if the above were the correct explanation of the reaction taking place.

The velocity of the reaction at low temperatures is probably very small, but we are dealing with reactions that take several years for completion.

The minute amounts of SO2 produced at any one time are under favorable conditions for conversion into SO3, as they are in the presence of relatively enormous surfaces capable of acting as contact surfaces. If oxidized to SO3, any oxide of iron produced, being in intimate contact with the acid, will be readily dissolved in the presence of moisture.

The foregoing, if true, will be as valid for sulfides of copper as for sulfides of iron. It is well known that some copper sulfides, under the influence of air and moisture, oxidize readily to form sulfates and that other sulfides do so much less readily; this probably explains the particular amenability of the Rio Tinto ore to the process, although the experimental results at Bisbee would indicate that the reactions take place easily with the forms of copper sulfide present in these ores.

The relative amounts of copper dissolved by direct oxidation of copper to sulfate and by direct solvent action of ferric salts cannot be determined; the latter reaction is no doubt important, but careful records kept of the ferric and ferrous iron entering and leaving the experimental piles did not disclose any relation between the amount of ferric iron entering and the copper dissolved.

If the foregoing is correct, it is evident that, at Rio Tinto, the most favorable conditions for rapid oxidation cannot be maintained without danger of spontaneous ignition of the heavy sulfides; but this is not the case for a pile of siliceous ore, for which the best conditions can be selected. The investigation has not been carried on sufficiently to determine what these conditions may be. It may therefore be possible that the conditions under which the test heaps at Bisbee and Tyrone were run were not the most favorable, and that further work will result in an appreciable shortening of the time required for extraction.

The recorded history of the method at Rio Tinto goes back many years, as stated by Mr. DeKalb. The precipitation of copper by immersion of iron in solutions of copper salts is one of the oldest known chemical facts, and as, generally speaking, iron has always been cheaper than copper, it is probable that the method was applied at Rio Tinto for many years before the recorded development of efforts at systematic use.

In this country precipitation of copper from mine waters has been carried on for many years, at various localities, but until recently no serious attempts have been made to apply heap leaching. The first work done on the method by the Phelps Dodge Corpn. was about 1900. After a visit to Rio Tinto, when Doctor Douglas was impressed with the possibilities of the process for Bisbee, some experiments were made to determine, first, whether heavy pyritic ore carrying copper from the Copper Queen mine could be treated; second, if at the same time the excess of acid and ferric salts produced could be used to leach the copper from low-grade oxidized ores in “reducing beds” similar to those at Rio Tinto. From the records of these tests, the main criterion seemed to have been as to whether or not the test pile would “heat up;” and as no “heating up” was observed when water was applied to the heap, the experiments were pronounced a failure. The addition of chlorine, as common salt, to the liquors appeared to give favorable results, but was too expensive to be practicable. The role of iron as a reagent and the necessity for its presence in the leach liquors was not recognized in these early tests.

It was shown later, by laboratory work, that the presence of iron salts in the water used for wetting the ore was an essential factor, and it was proved that, by attention to this and other necessary conditions, low-grade ores of various kinds could be leached in the laboratory by the method.

During the large-scale leaching and electrolytic tests at Douglas, authorization was obtained for a trial of heap leaching. The material used for this test was sand tailings from the Tyrone mill, about 25 tons of which were placed on a platform and treated for several months. The results obtained were sufficiently encouraging to recommend a large scale test at Tyrone. A brief description of this experiment, which was carried on by A. W. Hudson, is as follows:

Tyrone Experimental Heap

The ore for the test had mostly been produced from, development work and had been dumped adjacent to the various tunnels or shafts. Some of the ore had been in the dumps for over seven years, while some was from recent operations.

The leaching site was selected on account of its proximity to the ore. It consisted of a hillside with a natural slope of between 12 and 15 per cent. This was too steep to retain the slimes, so surface benches were graded with a gradual slope following the contours of the hill. The area was about 250 by 250 ft.

To assist in waterproofing the ground, mill slimes were spread to a depth of about 6 in.; culverts were made from the largest rocks.

About 20,000 tons of ore were elevated to the site and distributed with wagons, making a heap with an average depth of 6 ft. The following is an approximate analysis of the material: Cu, 2.71 per cent.; SiO2,66.0 per cent.; Fe, 6.0 percent.; CaO, 0.3 per cent.; S, 5.0 per cent.; Al2O3, 14.5.

Scrap-iron precipitation launders were built ahead of the heap so that all solutions leaving the heap were pumped to the plant, where they were depleted of the copper, enriched in iron content, and flowed by gravity to the heap again for washing purposes.

Leaching operations were commenced during February, 1917, and continued intermittently for three years, at which time further work was discontinued because the concentrating mill required all the water from the mine.

During the three years, a total of 38.6 per cent, of the copper was extracted, figured from measurements and assays of solutions on and off the heap. The addition of acid was found necessary.

Following the Douglas leaching tests, considerable experimental work was done at Bisbee by various methods on low-grade ores, mostly carrying acid-soluble copper. After the discovery of the extensive disseminated deposits of Sacramento Hill, and their exploration, it became evident that these contained large amounts of ore too low in grade for concentration, so recommendations were made by the research department of the Phelps Dodge Corpn. that a systematic and complete test of heap leaching be made on this material. This work at Bisbee has been described by Joseph Irving, who was in charge of the operation of the experimental heap.

It is an interesting fact that the proceeds from this small heap paid all expenses of operation, and that the actual cost of the copper produced compared favorably with that of other much larger scale operations during the same period. Experimental work paying for itself is sufficiently rare to make this worth noting.

Bisbee Experimental Heap

A dump containing a quantity of low-grade sulfide ore that had been mined from an air shaft two or three years previous provided the ore for the test.

The leaching site selected for the heap was in a gully, part of the floors being covered with old lumber and part dressed with slimes. The bed of the creek was used as the main drain into which all other culverts drained. The surface area of the heap was about 12,000 sq. ft. About 10,000 tons of ore were moved to the prepared site with mine cars, making a heap with an average depth of approximately 20 ft. The following is an analysis of the ore: Cu, 1.3 per cent.; SiO2, 60.7 per cent.; Fe, 10.5 per cent.; CaO, 1.2 per cent.; Al2O3, 12.1 per cent.; S, 9.9 per cent.

Leaching operations were commenced during April, 1917, with the use of mine water, which was later replaced by waste liquors from the scrap-iron precipitation plant. The solutions were measured daily going to and from the heap, the latter being passed through a series of scrap-iron launders for the removal of the copper. Solutions depleted of copper but enriched with iron were returned to the heap for washing purposes.

This heap was treated for three years with various resting and leaching periods, at the end of which time further work had to be discontinued on account of stripping operations by steam shovels on Sacramento Hill. During the three years, a total extraction of 45.2 per cent, of the copper was effected, figured from measurements and assays of solutions both on and off the heap.

The heap was systematically sampled with a series of drill holes and pits, as well as a steam shovel cut through one end. The general average assay was 0.36 per cent. Cu, showing the actual extraction of copper to have been 72.3 per cent.

This large difference between calculated and actual results was doubtless due to unaccounted for losses of liquor, which drained directly through into the ground, and emphasizes the necessity for proper preparation of the site for commercial work.

Large-Scale Installation for Treating Low-Grade Ores from Sacramento Hill

The material to be mined from Sacramento Hill is divided into four classes, as follows:

  1. Waste………………………………….0.0 to 0.5 per cent, copper
  2. Low grade………………………0.5 to 1.0 per cent, copper
  3. Concentrating………………1.0 to 3.0 per cent, copper
  4. Smelting…………………..3.0 per cent, copper and over

The second grade is classified as leaching ore. It is estimated that the deposit contains 8,500,000 tons of this material, averaging 0.72 per cent, copper.

During March, 1920, the first of the leaching ore from Sacramento Hill was placed on the leaching site, this ore being encountered during stripping operations with steam shovels. From that time until October, 1921, when operations at the Hill were suspended, approximately 380,000 tons have been mined and placed ready for leaching. The proposed general layout of the first unit of the plant and heaps is shown in Fig. 1.

The place selected for the leaching heaps was chosen because of its proximity to the ore, the contour of the ground, and the nature of the floor. The site selected for the first 2,000,000 tons of ore is approximately 1800 by 750 ft. The average slope of the ground is between 3½ to 4 per cent, thus allowing easy drainage of solutions from the heap. The floor consists of caliche and conglomerates.

The ore mined by the steam shovels was loaded into steel dump cars, built by the Western Wheeled Scraper Co., having capacities of 20 and 25 cu. yd., which dump either side of the track. These cars are dumped by compressed air.

Before operations were commenced, a track was built along the north end of the leaching site, and dumping operations were started on the ground sloping from this side to the south. Each trainload, consisting of six cars, was sampled at the heap after dumping. As the dump increased, the track was moved to the edge, the coarser material rolling to the foot of the heap. The large lumps were broken, with powder, to from 8 to 12 in, and some of these were used for building the culverts or drains. Fine ore was kept on the surface of the heap as required for a covering and for building the basins necessary for irrigation purposes.

The culverts were built ahead of the ore being dumped, the hardest and most suitable rock being used for the purpose. These culverts are used both as ventilating flues and as drains for the solutions. The rock must be packed tightly in place to resist the force of the lumps rolling down the face of the heap, otherwise the culvert would collapse and its purpose be defeated. While the cross culverts are continuous across the site, those connecting at right angles are staggered, so that the system is an interlaced network of flues. The culverts are not always built in regular order, for advantage is taken of all depressions or drainage channels on the ground surface. The opening in the culvert is 12 in. in the clear.

Preparation of Site

Before the ore was piled, the site was cleared of cactus, brush, etc. Waterproofing the surface of the ground under the heaps is important, for it was clearly shown, by sampling the Bisbee test heap, that the actual percentage of extraction calculated from an average analysis of the washed ore was much greater than that calculated from the records of quantities and analyses of solutions during operation. Assuming that the latter were correct, the only explanation for any great difference would be in loss of solution by seepage into the ground under the piles.

The question of a suitable and sufficiently cheap waterproofing method for the site of the large plant has not been settled satisfactorily. As long as it is kept wet, a layer of clay or of slime tailings will probably be sufficiently waterproof. As soon as slime tailings from the new mill are available, they will probably be used for additional ore heaps. This will be the cheapest available waterproofing method, as it should cost but little to flume these slime tailings from the mill to the site; and the cost of waterproofing per ton of ore piled under these conditions should be nearly negligible. At present, no slime tailings or clay are available at a possible cost.

Another possible method of waterproofing is by spraying the surface with oil. If this is done, the surface should first be thoroughly dry, and it should be dried between successive coats of oil, and thoroughly dried before the ore is laid down. While an oil-coated surface will be fairly waterproof, the culverts, etc., must be built upon it, which would cause it to be badly broken, unless it was thoroughly dried beforehand.

A good part of the ground under the heaps will consist of caliche, which should have formed upon it a crust of calcium sulfate from the leaching solutions and thus arrest the percolation of solutions into the ground, but it does not seem probable that too much reliance can be placed on the water-repelling character of such a layer, and a positive method of waterproofing should be adopted if possible.


There will be two reservoirs—one above and the other below the leaching heap. The reservoir above will be used for settling out suspended matter from the mine water previous to irrigating the heaps; it will also act as a storage in case of need. The one below will be need as a storage for the liquors coming from the heaps before going to the cementation plant for the recovery of copper. These reservoirs are not yet constructed.

In Fig. 2 is shown the proposed general arrangement of the first unit, of the cementation plant, which has yet to be erected. This plant is

designed so that practically all of the operations will be performed automatically. There will be twelve redwood tanks, each 24 ft. in diameter by 10 ft. high, placed in two rows of six in series. They will contain false bottoms for supporting the scrap iron. Underneath the false bottom will be an acid-proof stirring arm worked from a shaft in the center of the tank and suspended from the top of the tank. This will be used to agitate the solutions when necessary and to move the precipitated copper to the central discharge on the bottom of the tank, which will be partly conical to assist in this operation.

The liquor entering the tank will be introduced alongside the agitator shaft and delivered underneath the false bottom, flowing up through the iron, where it will discharge over a peripheral launder to the next tank in series. Each tank in the unit will be connected to the next tank in parallel by a launder so that any tank may be cut out for inspection with-out interrupting operations.

The iron will be distributed to the various tanks by a gantry crane, which will take the scrap from the railroad cars at the head of the plant or from storage.

The liquor, after being depleted of its copper, will flow to an equalizing tank, where the required amount will be returned to the heaps for washing purposes. Provision is made for the installation of a scrap-iron launder plant should it be found necessary to remove the last traces of copper from the solution.

The classifier will be of the Dorr type, built of acid-proof material, as will also the thickening tanks of which there will be two, 24 ft. in diameter by 10 ft. high. These will be equipped with acid-proof diaphragm pumps, which will remove the thickened cement copper to the drying floors, the clear overflow being pumped back into the cementation-tank circuit.

The drying floors will be built of concrete and so arranged that all surplus water can be drained to a sump and either returned to the system or go to waste.

Flow Sheet

Subject to variation from data gained from subsequent operation, the flow sheet and method of plant operation that will be followed are as follows; Fig. 3 shows the present plan.

The mine water, after leaving the reservoir, will be measured, sampled, and passed on to the leaching heaps, where it will be enriched and flow to the reservoir at the foot of the heaps. From there it will be measured, sampled, and go to the head of the cementation plant, passing through two rows of six tanks in series containing scrap iron. The solution, now depleted of its copper, will flow to a sump, where the required quantity will be returned to the heap, the remainder being sampled and run to waste. The precipitated copper will be removed, at intervals, from the cementation tanks and will pass through a classifier where the coarse copper will be removed and deposited on to drying floors. The fine or suspended copper will go to thickening tanks, from which the clear solution will be returned to the cementation tanks, while the thickened product will be pumped to and deposited on the drying floors. When sufficiently dry, the product will be shoveled into mine cars and dumped into railroad cars for shipment to the smelter.

As the operations in the heaps at Bisbee and Tyrone were experimental, the procedure there may not be closely followed as experience is gained in the larger heaps.

The first of the ore was laid down before the cementation plant was available. If a part of this was installed, the logical method would be to start wetting the heaps as soon as possible after they were laid down, applying only enough water to moisten the ore thoroughly, and then allow them to wait until ready to start operation. If the first part of the ore was wet in this way, a considerable part of the copper would be made soluble by the time the heap was finished.

Without sufficient precipitating capacity first, however, most of this soluble copper would be lost when the heaps were wet during the rainy season.

The following assumptions have been made in the design and flow sheet:

In normal times, a delivery to the leaching heaps of about 400,000 tons yearly up to the total quantity of ore graded as leaching material. The cementation plant to be increased by units as required. Grade of ore 0.72 per cent. Cu. Period of extraction, 6 years to make a total recovery of 70 per cent, divided as follows:

The period of extraction is an arbitrary assumption and the possible profit will be reduced in proportion to extension of time of treatment.

While all assumptions are believed to be conservative, the operation should be considered as a large scale experiment, the results of which will be of interest in view of the bearing it may have on the future treatment of stripping ore and other low-grade material, which in some cases may mean for some mines the addition of considerable tonnage to what is now estimated as ore.

The main experimental factor is the required time of treatment, and in operation the adoption of a method of treatment of site which will absolutely prevent solution losses.


J. Parke Channing, New York, N. Y.—When the Tennessee Copper Co. started smelting in the Ducktown district the ore was first heap-roasted. These heaps or piles were all protected by sheds which were kept in good condition and I feel quite sure that none of the copper was ever lost by leaching. When pyrite smelting was started and the roasting of ore abandoned, the roast yards were carefully cleaned up and B. B. Gottsberger made a final calculation of the total amount of copper that went into the roast yards and the total amount that was taken out. My recollection is that there was an unaccountable loss of 2 lb. of copper per ton of ore. Inasmuch as we were absolutely sure that this was not leached by rain water, the only explanation that I could give was that this copper was volatilized even at the low temperature of the roast heaps, but we have no scientific evidence to this effect.

R. C Canby, Wallingford, Conn.—The thought that impressed me while reading this paper is the statement that it is not the washing action of the solutions but the capillary action of the solutions to and from the inside of the ore that produces the results. So these periods of oxidation might perhaps more properly be called periods of capillary action.

C. S. Witherell, New York, N. Y.—I think capillary attraction also performs another role; not only does it serve to make the leaching solution penetrate the lumps of ore through numerous small cracks, but when the salts crystallize, an expansion takes place, thus disintegrating the lumps and opening other passages for the leaching solution.

Edward L. Blossom, New York, N. Y. (written discussion).— For successful heap leaching there are two outstanding requirements: (1) Every step of the operation must be carried on at low cost. (2) The copper, if not already in water- or acid-soluble form, must be capable of oxidizing to that form in a reasonable period of time. The authors of this paper seem to have fixed 6 years as the maximum but hope to make the time shorter.

Carbonates and oxides are the most readily leachable copper minerals. Of the sulfides, chalcocite and bornite are the easiest dissolved. Chalcopyrite is more refractory. The material should be broken (cost permitting) sufficiently to expose all copper mineral to contact with the solution, but as few fines as possible should be made.

Water, with or without sulfuric acid, and with or without mine waters carrying iron salts, is the solvent employed. Access of moisture and oxygen to interior of heaps is indispensable for producing the desired results. The authors stress the importance of alternately wetting and drying each piece of ore in the heap, by this means they depend on reversed capillarity to bring the dissolved copper salts from the interior of each piece to the surface whence it can be removed by the next wash. This is the same physical factor as that which brings desert alkalis to the surface of the soil, and its recognition as an important factor in leaching is highly creditable to the experimenters. So far as I know, this factor lias not previously been mentioned in the literature on the subject.

The authors recommend waterproofing the ground on which the heaps are to be spread and present figures showing that losses up to 37 per cent, of the extracted copper have resulted from neglect of this precaution. The authors’ propose to precipitate their copper from solution in tanks resembling Dorr thickeners, each tank having a false bottom above the stirring arm to carry the scrap iron. This also is a novelty. Scrap-iron launders of the old type will be used, if at all, only for removing the last traces of copper from the solution. A portion of the depleted liquor will be returned to the heaps for washing purposes.

The chemistry of the process—how insoluble copper compounds are converted into soluble salts—deserves careful attention. The authors give a number of reactions which probably contribute to the desired end, but of the relative importance of these reactions little is known with certainty. Some of the copper sulfide is probably converted into sulfate by direct oxidation. More of it is certainly dissolved by the action of ferric sulfate, which also attacks any metallics which may be present. Oxides, carbonates, and silicates of copper are more or less completely dissolved by free acid. Ferrous and ferric sulfates, together with free sulfuric acid, are continuously produced in the heaps by oxidation of iron sulfide minerals. These reagents, freshly generated in close proximity to the copper mineral, doubtless play a major role in the extraction. But the same reagents may be introduced with the wash liquors. The efficacy of sulfuric acid is presumably independent of its source, but for some reason the iron salts introduced when mine water or depleted liquor from the precipitation tanks are led onto the heaps do not produce any definitely ascertainable increase in copper extracted. This, I have been told, is also the experience at Rio Tinto. Yet the authors state that in laboratory work the presence of iron salts was shown to be an essential factor and anyone who has observed the powerful solvent effect of ferric sulfate on the minerals in question cannot but be surprised that addition of this reagent in heap leaching should not improve the results.

For this apparent conflict between theory and practice two explanations suggest themselves:

  1. Mine water and tank liquors, having been in contact with reducing agents, contain a, relatively small proportion of ferric, as compared with ferrous, sulfate. The latter does not become an active solvent unless it is oxidized in the heaps.
  2. Mine water and tank liquors (in contrast with freshly made laboratory solutions) usually contain basic iron compounds ready to fall out of solution when the acidity and the concentration of ferric sulfate are reduced. Both are reduced by contact of the solution with reactive copper minerals, and in consequence a film of iron compounds is likely to be precipitated just where it will do most harm, viz., on the surface of partly dissolved particles of copper ore.

Once precipitated, this film is difficult to remove even if the subsequent washes contain free acid, and in consequence the needful contact between mineral and solvent is lessened or destroyed. Oftentimes mere dilution of a neutral iron-bearing solution will bring down the precipitate. Cases are on record where application of mine waters has so cemented a heap as to ruin it for leaching purposes.

The remedy, if there be a remedy, is not easy. A very considerable addition of acid to the wash waters would be required to insure maintenance of acidity in all portions of the heap. In presence of acid-soluble minerals other than copper the cost of this would be altogether prohibitive.

Were it possible to oxidize all the iron in the wash water before applying it to the heaps, interesting results might follow, but too much experimental work has been done on these lines to leave much hope that such oxidation can be economically accomplished.

We are thus brought back to the conclusion that improvements in heap leaching must lie in the increase of oxidation effects within the heap. Mention is made of the catalytic effect of the relatively enormous surface of the ore in converting SO2 to SO3. May not some cheap accelerator of catalytic action be found that can be mixed with the ore?


Coke Oven Design and Operation

The conversion of the beehive coke plants, in this country, to byproduct plants has been slow, because the coal supplies were near the centers of the steel industry. With the growth of this industry, especially with its development around Chicago, it became necessary to transport large tonnages of coal from the eastern districts and then convert it into coke. The losses due to transportation costs were partly offset by the value of the byproducts recovered.

To save the transportation costs, it was desirable that the coal deposits of Indiana and Illinois be utilized. This coal had always been classified as non-coking; it was also considered unsuitable for the metallurgical field because of the high ash and sulfur contents. The conditions, however, were promising enough to start experimental and development work, which crystallized in the design of the Roberts coke oven.

The fundamental features of most coke ovens, with respect to the application of heating gas and the recovery of byproducts, are the same and, in the last few years, the tendency has been toward the better application of heating to the walls, higher thermal efficiency, and, by the control of heating conditions, the increase of byproduct yields. Structural features also have been improved and the use of high-grade refractory material has allowed the use of higher temperatures with the resulting higher rated capacity per oven per day. Great improvements have been made in both recuperators and regenerators, particularly with reference to the individualizing of each oven with respect to its adjacent oven, so that each can be operated as a unit if desired. The highest development of the flue type of oven has been applying individual regenerators for the recovery of waste heat.

After some research work, it was found that the high-volatile Illinois coals could be utilized for coke by obtaining the better application and control of the heating conditions in the oven. Most of this high-volatile coal that had been used gave a good coke structure under certain conditions. The fact that coke was occasionally made from this coal, demonstrated the possibility of making coke from it at all times provided the conditions under which the coke cell or structure formed could be isolated.

One problem in coke-oven design is to heat uniformally a surface as large as a coke-oven wall so that the results of the heat distribution will be uniform throughout the entire coking mass. Yet, the necessity of this is demonstrated by the fact that when a coke-oven wall is uniformly heated and a sufficient quantity of heat made available for the coking mass a good coke structure will be made from a large number of coals that otherwise do not give good results.

A coke-oven wall of standard size is approximately 43 ft. long and 14 ft. high, or about 600 sq. ft. With about 1200 sq. ft. of surface exposed to a cake of coal approximately 14 in. thick and weighing 30,000 lb., it is necessary that each square foot of the surface exposed to the coking mass be as nearly the same temperature as possible. One of the conditions in the coke-oven chamber that influences the quantity of heat required is the taper, which will vary from 1¼ to 2½ so that there is approximately 15 per cent, more thickness in the discharge side of an oven than on the pusher side, which means that at least 15 per cent, more heat must be made available at that zone. As this variation in thickness is gradual, the heat supply along the wall must be graduated. The variable heating conditions in the height of the wall must also be compensated for both, because of structural features in design and because of the different heat requirements in each vertical zone.

The Roberts oven is designed in three types: regenerative, recuperative, and the combination-regenerator oven, which can be heated by coke-oven gas or by blast-furnace and producer gas; or where it is necessary to preheat both the air and fuel gas. The fundamental principles of these designs are the same. The regenerative type necessitates the reversing of the flow of the air and gas; the recuperative type permits the flow of air and gas in one direction continuously. The ovens are mounted on concrete foundations, which are simply large flat pads of reinforced concrete of sufficient strength and area to support a number of ovens in each battery.

Roberts Recuperative Oven

In the recuperative oven, the foundation contains the ducts through which all the air required for operation passes. By this means the tendency to overheat the concrete foundation is overcome in a simple and effective manner; the incoming air absorbs the heat as rapidly as it is transferred to the foundation, so that the foundation is maintained at approximately atmospheric temperature.

A longitudinal section of the ovens and a cross-section of the concrete pad are shown in Fig. 1. Seven ducts passing through the concrete, the full length of the battery, carry air at atmospheric temperature in the bottom of either the recuperator or regenerator oven, depending on which type is used; Figs. 1 and 2 show how these ducts are connected to the recuperator.

The foundation of a battery of coke ovens must be stable for, if because of expansion and contraction the foundation is ruptured, there is great possibility of rupturing the brick structure above, which would allow short circuiting of air and gas, which might result in bad operating conditions.

Directly on top of the concrete pad are two courses of fireclay brick. These act as heat insulation for the concrete and form a surface on which the silica brick slides when expanding or contracting. From the top of this fireclay course to the bottom of the battery nothing but high-grade silica brick is used. The supporting walls between recuperator or regenerator are made up of silica straights and shapes. They are 18 in. thick and are placed directly under the oven heating walls, so that the maximum support is provided for the entire structure above. Expansion joints are provided for the lower part of the oven chamber, the location of which is shown by the heavy black lines in (a), Fig. 3. This construction effectively prevents leakage from the oven chamber in the recuperator or regenerator chambers; the expansion joint also allows for the vertical expansion of the recuperator. The linear expansion is controlled by buckstays on the ends of the oven and by auxiliary buckstays at the end of the recuperators. The small buckstays are fastened to the main buckstays but may move independently. This arrangement permits the free movement, either vertically or horizontally, of the recuperator brick

regardless of the movement of the oven brick or the supporting walls. The recuperator is made of silica brick, which always expands when going from atmospheric temperature to working temperatures thus permitting the control of the joints, so that leaks are reduced to a minimum. The heavy silica blocks are so laid that all vertical joints are broken. In this part of the structure are located the sole flues and the waste-gas equalizing flues for each heating wall.

From the sole of the oven to the top of the combustion-chamber wall, two adjacent ovens are made up of two heating walls and an intermediate wall. This triple wall is about 30 in. thick and forms a structure of exceptional strength between adjacent oven chambers.

The center of the separating wall is composed of large flat blocks and shapes through which pass the ducts for air and the supply of secondary-fuel gas. On either side of this separator wall are the heating walls of one side of two adjacent ovens. In these walls is incorporated the space for the combustion of fuel gas.

The shape of the brick of which this wall is constructed and the way they are assembled form one of the features of the Roberts oven. The brick is shown in Fig. 4; the manner of laying is shown at (b) and (c), Fig. 3. Each wall is backed by the center wall; the three walls together are sufficiently strong to withstand any stress placed upon them by operating conditions. From (b), Fig. 3, it is evident that the only stresses the bricks have to resist are in the direction of the arrows D, and they will resist these the same as if solid. They are so laid in the wall that the ends in registering form an arch and the pressure along the line EE is on this arch. The entire wall, therefore, is practically as strong as if there were no openings in it; at the same time ample open space is allowed for the passage of the burning gases through the wall.

In laying, each course is offset half the width of a brick, breaking the vertical joints, while at the same time the centers of the brick in one course come directly over the openings of the course below, as shown at (c), Fig. 3. From the standpoint of strength, the effect of this is the same as if each brick were reinforced over 50 per cent, of the area exposed to the coal, and the next brick above or below has the same amount of reinforcement offset on a new center, thereby increasing the strength

enormously. The strength of this interlocked wall structure and, the comparatively weaker construction of flues is at once apparent and the effect upon the extraction and disposal of the heat from burning gases in this wall is influenced in a marked degree, as compared to the flow of gases through a flue.

Experience has shown that it is difficult to hold a flue-like structure tight, there always being a tendency to develop leaks between the combustion and coke-oven chambers, causing overheating by the introduction of gases from the oven chamber in the combustion flues. This uncontrolled supply of gas easily causes hot spots and uneven heat effect upon the coking mass.

A flue structure is unsupported throughout its length, amounting to more than 700 sq. in. in the smaller type of ovens, while the greatest unsupported surface in the Roberts wall brick is only a little more than 14 sq. in.

A common comparison is to measure the total length of a joint exposed to the coke in a coke-oven wall, which is all right when comparing flue-type ovens, but this comparison does not hold true with a structure such as the Roberts oven, because the strength of the wall structure is dependent on the fact that the combustion chamber and the two faces of heating wall are incorporated in one brick, the rupture of which depends on the cracking of this brick in the center of the combustion chamber.

The inherent weakness in the flue-type structure is overcome by the use of thicker tiles between the combustion and coke-oven chambers, this tile in some cases being 6 in. thick and the average is well over 4 in. The Roberts wall is only 3 in. thick between the burning gases and the coal charge; this thickness is uniform from the top to the bottom, and from end to end of the wall. All the parts of this wall are uniform in strength, heat absorption, and conductivity. With the thin wall through which the heat is conducted to the coal, there is still the full strength of a solid wall 30 in. thick. This particular construction permits the apparent contradiction of the belief that thin coke-oven walls are fragile.

Near the top of the Roberts wall, and immediately below the nozzle for the introduction of the primary gas, is a short space in which the baffle brick are omitted; this forms the mixing chamber. While this chamber is short and comparatively narrow and the necessity for great strength is reduced at this point in the wall, the same keyed construction is used, only the baffles being omitted. Above the combustion chamber, the wall is solid except for the openings through which pass the gas ducts and the openings for the top air adjustment.

The crown of the oven is silica but, as the temperature above this point is low, clay shapes form the top of the battery. Clay is a better heat insulator than silica and is also more resistant to the weather. It is also possible to introduce high-grade heat insulating material in portions of the battery top, thereby lessening radiation losses.

Above the clay top brick are the gas headers that supply gas to the primary and secondary burners. These headers are rectangular in shape and the spaces between them and the brick are filled with a cement mixture so that the top of the battery is perfectly smooth.

The ends of the ovens are faced with clay shapes, which act as heat insulation and are also a protection from the weather. These clay shapes lie directly behind the buckstays and cover the entire space between the oven doors. A layer of silica, 33, Fig. 1, between the clay and the wall brick, protects the clay from direct contact with the heating gases. The clay shapes at the ends are made as large as possible, in order to reduce the number of joints; also they are so designed that all joints may be readily pointed up in case the cement falls out from weathering.

As the expansion of clay and silica are quite different, the clay used for these ends will not move vertically to as great an extent as the silica in the heating walls. The silica will, therefore, slide on the clay and, because of the great friction caused by the pressure of the buckstays, the clay will have a tendency to follow the silica in a vertical direction, thereby opening the joints between the clay shapes. This will cause leakage of the burning gases from the wall to the atmosphere, which leakage has caused considerable annoyance and damage, particularly when starting up new batteries.

This difficulty is overcome, in the Roberts oven, by placing on the buckstays a steel angle 34, Fig. 1, that extends over the top of the clay insulating brick. The bottom of the buckstay is secured to the concrete pad, so that the clay shapes are securely held in place and their vertical movement is prevented. The silica slides on the clay without breaking open the joints in either the heating or the insulating walls. The entire surface of the insulating brick remains perfectly bonded and unbroken.

At no time during the heating up or when operating can leaks be detected in this part of the oven. The buckstays remain cool and are unwarped; the alignment of the exterior of the oven is most noticeable. The system of rigidly holding the clay insulating brick in place lessens the labor necessary in keeping the jambs pointed, thus making the luting of the doors much easier.

The door consists of a cast-iron frame tapered to fit the oven chamber. The brick forming the lining of the door are placed in the frame from the back and, being wedge shaped, are firmly keyed into place; the space in the frame behind the brick is filled with powdered insulating material and a steel plate is bolted to the frame. This gives a door that fits in its frame in the oven and has the lining brick firmly keyed so that they will not shift. The insulation prevents radiation and the backs of the doors are so cool that the hand may be placed on them.

The jamb into which this door fits is also of cast iron; the cast-iron jamb is preferred as it can be accurately set and so cemented in place that it will not move or leak. It does not spall, as does the brick jamb, and always presents a smooth surface for luting the door. This care in fashioning the door and jamb prevents leaks around the doors and lessens the labor in luting.

The buckstays used on the Roberts oven are heavy, for the strain placed on them is enormous. In most of the early installations, where the buckstays were of insufficient strength, they bent to such an extent that it was difficult to place the doors in the oven.

The usual practice in the United States is to carry the charging car, or larry, on rails laid directly on top of the brickwork of the ovens. As this car carries from 13 to 16 tons of coal, it sets up considerable vibration in the brickwork of the ovens when moving over the top.

In the Roberts system, the larry is carried on rails laid on top of the buckstays, so that the weight is carried direct to the foundation of the battery and the top is relieved of the shock from a heavy moving load. In addition, the battery top is cleared of all obstructing rails, there is less danger to the men on top from the moving larry, also the rails carrying the current for operating the car are placed on the side, well out of the way.

Operation of the Roberts Oven

The operation of the Roberts oven is extremely simple and may be readily followed in Figs. 1 and 2. The air enters through the tunnels 30 and passes up the smaller ducts 1 into the air equalizing duct 2. The openings 1 are regulated by dampers 3, which are controlled from the outside and are easily seen through openings left in the ends of the recuperator walls for this purpose. From the equalizing ducts 2, the air passes up around the outside of the recuperator tile. This passage is alternately between the tile and the supporting pier and then between the tile so that each tile is surrounded on three sides by the ascending air. In this manner, the air is thoroughly heated and the tile maintained at an even temperature differential, and at the same time the supporting piers are kept at a temperature no higher than the recuperators.

As the air flows countercurrent to the waste gas, it meets progressively hotter tile as it ascends and, on reaching the outlet ports 4, has attained the temperature of the waste gases at this point. The air then flows through the ducts 6 to the top of the heating wall, horizontally through the passage 7, then downward through the air ports 9 that surround the primary-gas nozzles 10. The quantity of air admitted to the air ports 9 is controlled by the slide brick 8. This brick is reached from the top of the battery through the openings 35 so that, by means of a short iron rod, the operator may accurately regulate the size of the opening 9.

As the recuperator tile is of silica with a high heat conductivity, the air is raised to 2000° F. at the point where it leaves the tile and enters the duct 6. Passing upward through this duct, the air arrives at the tip of the burner at a constant temperature throughout the entire oven. The advantage of having the air reach the burner at a uniform temperature at all times and throughout the entire battery cannot be over-rated as uniformity of heating can only be obtained, when conditions governing combustion are uniform. The falling off in air temperature because of the cooling of a regenerator is well known; its effect has been measured and is admitted to be a factor in the heat effect of a reversing oven. The uniformity of preheating the air in the Roberts oven is but one of the points attained by the designers in the effort to eliminate “average conditions.”

After the byproducts have been extracted, the fuel gas returns to the battery and then enters a large gas header, which is supported on steelwork attached to the buckstays on the coke side of the battery, occupying on this side approximately the same position as the collecting mains on the pusher side. From this header, the gas is distributed to the individual gas headers 11 which are embedded in the top brickwork so that the top of the battery is smooth and unbroken

These headers B, Fig. 5, are of square cross-section and are divided into two pipes of equal capacity. One side carries the primary gas and the other side carries the secondary gas.

The delivery of the gas to the individual headers is through a manifold, Fig. 6, equipped with shut-off cocks; from the rectangular headers the gas passes to the burners through the burner cock 12. Thus the gas may be shut off from the entire oven or from any individual burner cock. All the cocks mentioned are only for shutting off the supply and not for the regulation, as this regulation is by means of orifices in the manifolds supplying the burner headers and in the body of the burner cocks.

The regulation system of the Roberts oven is based on the fact that the flow of gas through an orifice is proportional to the pressure of the gas. A disk is inserted in each manifold supplying primary and secondary gas just below the shut-off cock. This disk has an accurately drilled hole of such size as to pass the required quantity of the primary and secondary gases. If the amount of secondary gas is to be 20 per cent, greater than that of primary gas, the disk for the secondary gas will have an orifice 20 per cent, larger than the orifice through which the primary gas passes. The proportions of the two gases are accurately known and all the disks are drilled before the ovens are heated and when once set in place need never be changed. To change the coking time, it is only necessary to raise or lower the pressure on the main battery header, and the primary and secondary gas orifices will pass the exact proportions of gas required. The supply of gas to the individual headers is thus correctly proportioned at all times and is independent of setting a valve or cock to what must be only an approximate position.

The gas supply through the individual burner cocks is cared for in a similar manner; a cross-section of the gas cock is shown in Fig. 5. The core of this cock is drilled with two holes set at 180° apart; as the core may be turned through 360°, either hole may be made to register with the gas inlet A. If the core is turned to a position midway between the two openings, the gas supply will be entirely closed. In practice, one of the holes is made the full size of the gas inlet but the other is drilled, with extreme accuracy, to the size necessary to compensate for the taper of the oven.

The oven chamber is wider at the coke end than at the pusher end, so that the coke will readily push from the oven. In the Roberts oven, this taper is generally 2 in.; that is, the oven will be 2 in. wider at the coke end than at the pusher end but this taper is uniform throughout the oven as the walls are built without offsets. It is obvious, therefore, that the thickness of coal at the coke end will be greater than at the pusher end; and this thickness will vary uniformly between the ends of the oven, so that it is necessary to burn more gas at the wider end if the entire mass of coal is to be coked in the same number of hours. Also the quantity of gas burned in any part of the oven should be graduated to this increasing taper of the oven from end to end.

The small holes drilled in the core of the cocks gradually increase in size from the pusher side to the coke side of the oven. This increase in size is accurately proportioned to the increase in the amount of the coal charged. As the cores are iron, they may be drilled to the thousandth of an inch and the areas of the holes graduated with great exactness. In the case of a 2-in. taper and a charge of 15 tons per oven, the Roberts oven will have approximately 15 per cent, more coal at the coke end than at the pusher end. The gas burned at the coke end will, therefore, be 15 per cent, greater than at the pusher end of the oven and each intermediate burner will supply the exact amount necessary to coke the coal in the portion of the oven heated by that particular burner.

As the increase due to taper is known, these cocks are drilled before the oven is put in operation and are not changed; for, as is the case with the supply to the headers, all that is necessary to change the coking time is to change the pressure. As the pressure is changed, each burner cock will carry the correct proportion of gas necessary to do the work at that point in the oven. The result is uniformity of heating and the coking of the entire charge in exactly the same time. Such uniformity is impossible without accurate regulation, such as is attained in the Roberts oven.

The vertical-flue reversing type of oven cannot accurately allow for this taper in the oven for it can only be regulated for average conditions. There is an attempt at regulation by increasing the pressure on the coke-side gas header, but as the individual gas nozzles cannot be regulated, the supply to each nozzle is not under close control.

In the Roberts oven, the fuel gas is introduced at numerous points in small, accurately measured quantities, so that the relative temperature of one part of the wall compared to the other is under control, and no point is subject to overheating from the rapid combustion of a large quantity of gas introduced at one point as in other ovens.

Combustion and Flow of Combustion Products

After passing the accurately graduated gas cock 12, the gas is conducted to the burner nozzle 10 where it meets the air flowing through the air ports 9 and combustion starts in the short mixing chamber 13. The gas introduced at this point is called the primary gas and is approximately one half of the total supplied to the heating wall. The rest of the gas, called secondary gas, is introduced into the secondary-gas ducts 31 through the same type of graduated cock as the primary gas. The secondary-gas ducts pass downward through the center wall to a point 32 about midway to the bottom of the oven and there enter the heating wall.

The initial combustion of primary gas is in the mixing chamber 13, as it is introduced with the total amount of air necessary for combustion of the total gas supplied to that heating wall; that is, this gas will meet twice the amount of air required for its combustion. This large excess of air acts as a depressent to the flame temperature. As the air is not preheated to flame temperature, there will be a reduction in the temperature of the flame corresponding to the amount of heat required to raise the excess air to the average temperature of the mixture of burning gas and air.

The gas and air are also introduced at a neutral pressure so that they mix very slowly, with the result that combustion takes place quietly and evenly. The mixing chamber aids in extracting the small amount of radiant heat in the burning gas. The gas used for heating a coke oven usually contains but a small percentage of the illuminants, therefore, the radiant heat generated is proportionately small.

At the bottom of the mixing chamber 13, the burning gas meets the standard checkered-wall typical of the Roberts construction. The stream of gas in each mixing chamber will here be broken into three parts by the wall brick. As there are 24 primary burners in each heating wall, there will be 72 streams of burning gases in the wall from the bottom of the mixing chambers downward. Such an extremely uniform distribution of the heating gases is not found in any other type of construction.

By impinging on the brick at the bottom of the mixing chamber, the gas and air are more intimately mixed and the extraction of heat generated by this combustion is increased. Sweeping around the brick in the top row of checkers, the burning gas flows downward directly on top of the brick in the next tier below, moving in this way to the bottom of the wall. The result is complete combustion; and as the brick are entirely surrounded by these gases and also present the maximum surface for the absorption of the heat, a high degree of heat extraction is attained.

As a matter of fact, the extraction of heat is so complete that at a point 32 about midway down the wall the primary gases cannot heat the wall to coking temperature. At this point the secondary gas is admitted in the proper proportion to continue the generation of heat in the lower portion of the wall. Sufficient air is introduced with the primary gas to support also the combustion of the secondary gas; the oxygen in this air will, therefore, be available for combustion with the secondary gas at its point of introduction. At this point the temperature of the secondary gas is high, also the temperature of the waste gases from the primary combustion. The combustion of the secondary gas, however, will be subdued by the high proportion of inert gases present. These inert gases are the carbon dioxide and water vapor from the primary combustion and the nitrogen present in the air introduced with the primary gas. The combustion of the secondary gas will, therefore, be quiet and even. As this combustion takes place directly in the same checker brick of the wall, it is complete, and the extraction of the heat is as perfect as in the case of the primary gas.

As nearly all the available heat is extracted from the waste gases by the time they reach the bottom of the walls, the two streams are combined by the ducts 14 and this combination produces sufficient heat to maintain the sole of the oven at coking temperature. The coal in the sole of a Roberts oven is coked in a vertical direction by this method. The heat is thus extracted from the burning gases to such an extent that on reaching the lower sole flue 15 the gases contain only sufficient heat to be used for preheating air.

The upper sole flues 16 are connected to the lower sole flue 15 by six openings, each of which is controlled by dampers 17. By these dampers, the differential in the heating wall is maintained uniformly from end to end. These dampers are readily reached through openings left for this purpose.

From the lower sole flue 15, the waste gases pass down through 18 and through the top pass of the recuperators to the downcomer 19, then through the lower pass into 20 and offtakes 21 to the waste-gas tunnel 22. The waste-gas offtakes 21 are equipped with butterfly dampers 23 by which the draft for the oven is regulated; these dampers are readily set from the exterior of the offtake.

The ideal sought in any coke oven is the completion of the coking of the entire charge at one time; this can only be attained when there is uniform distribution of heat from the top to the bottom of the oven and a gradual distribution of heat from end to end to compensate for the taper of the oven and the increased thickness of the charge toward the coke side because of this taper.

The progress of heat, and consequent heating in an oven, in the vertical planes should be parallel to the walls of the coking chamber and the rate of this progress should be proportional to the thickness of the coal charge between the wall and the center of the chamber. The walls of the chamber are perpendicular, therefore, the thickness of the charge will be practically the same throughout its height, but the thickness will vary from the pusher end to the coke end in proportion to the taper of the oven.

Temperature in the Oven Top

The temperature of the space above the coal charge should be as low as possible, for it is through this space that the gases distilled from the coal pass on the way to the collecting mains. These gases are composed of hydrocarbons, some of which are readily broken down by heat; this breaking down results in free carbon and the destruction of valuable byproducts. The initial application of heat in the Roberts oven is at a point well below the top of the coal charge. The combustion is then downward, so that the tendency to overheat the top is eliminated. The temperature of the upper zone of the chamber may be changed by changing the amount of primary gas burned; this is done by changing the size of the orifice or the gas pressure.

Generation of the Heat

The heat necessary for coking the coal is generated by burning gas in the walls of the ovens. Usually this is a lean coke-oven gas of about 525 B.t.u. per cu. ft. Sometimes producer gas is used, but then it is necessary to preheat the gas as well as the air in order to maintain the necessary flame temperature.

The principal factors governing the burning of a gas are the ignition temperature and the rate at which the particles of combustible matter in the gas combine with the oxygen present. This rate is dependent on the quantity of oxygen present and its dilution by inert gases, and the velocity of the gas and air streams during combustion.

Ignition Temperature

All coke ovens operate at a temperature sufficiently high to ignite the gas. Combustion will, therefore, be sustained as long as air and gas are admitted to the walls. The temperature at the point of admission will, however, have considerable effect on the rapidity of the combustion. If the temperature is high and there is just sufficient oxygen to combine with all the combustibles, a high flame temperature will result and the evolution of heat will be rapid.

Rate of Combination of Gas and Oxygen

The lean coke-oven gas generally used to heat up an oven ordinarily contains a high percentage of hydrogen. There is also a high percentage of methane, which, on burning, decomposes, producing more hydrogen. The calorific power of hydrogen is not great but it is rapidly combustible , and produces a high flame temperature.

The introduction of this gas into a highly heated chamber, such as the heating wall of a coke oven, with the theoretical amount of highly heated air necessary to burn it, produces a high local temperature. If, in addition, the gas and air have a comparatively low velocity, the mixing will proceed at a high rate with the evolution of large volumes of heat.

The rate at which the gas is introduced (the quantity per unit of time), sufficient air for combustion being present, will determine the rate at which heat is generated. In those ovens, both vertical- and horizontal-flue types, having but a few points for the introduction of the gas, the quantity of gas burned at each point will be far greater than in an oven using many points of introduction. In the Roberts oven, there are 96 points of gas introduction while most ovens of the flue types introduce gas at not more than 16 points. It is obvious that, in the same coking time and using the same quantity of gas per ton of coal carbonized, the oven using but 16 points for the introduction of gas will burn six times as much gas at each point as the Roberts oven will burn at each of the 96 points. The danger from local overheating in the latter oven is further lessened by the method of transmitting the heat to the coal.

The temperature produced by the combustion of the gases sets up a heat flow that tends to balance the temperature difference between the heating gases and the oven wall with which they are in contact. This balance is influenced by the area of the wall in contact with the heating gases and the rapidity with which the heat is transmitted through the wall to the coking mass. In the Roberts wall, the area for the absorption of the heat is two and one-half times as great as the area that distributes this heat to the coking mass. It is believed that the Roberts oven will transmit the heat generated by the combustion of the heating gases almost twice as rapidly as is done by other types.

Air Control

The air supply in the Roberts oven is controlled by the damper 3 at the entrance to the recuperators and by the slide brick 8 over the air port 9. By means of the damper 3, the flow of air through the recuperator is equalized so that all parts will receive the same amount of air to preheat; this assures uniform temperatures in the recuperator and so reduces movement of the brick as the result of expansion and contraction. These dampers also regulate the quantity of air supplied to each oven. In other types of ovens, the air port is used first for the passage of air and, on the reverse, for the passage of waste gases. The quantities of air and waste gas are not the same, therefore, the conditions under which they should be governed are different.

In the Roberts oven, all the air is introduced at the top of the heating wall with the primary gas. As the latter is approximately only one half the gas required in the oven, there is no question of there being sufficient air for perfect combustion of the primary gas. The control of the heat generated by the primary gas is, therefore, dependent on the adjustment of the primary burners alone.

After leaving the mixing chambers, the burning gases enter the zone of checked brick typical of the Roberts oven. As this portion of the wall is practically one single chamber with inter-communicating passages, the air and gas are free to move from one part to the other so that not only is combustion of the primary gas complete but the secondary gas will come, into contact with sufficient air to make combustion in the secondary zone complete. The control of the heat generated by the secondary gas is, therefore, dependent only on the amount of secondary gas and this control is consummated with the same accuracy as in the case of the primary gas.

Control of the Heating Gases

The control of the heating gases in the heating walls plays an important part in the distribution of the air, and this movement is controlled by the distribution of the draft. The regulation of the draft or differential (that is, difference in pressure between the point of admission of the gas and air and the pressure of the outgoing waste gases) through the various parts of the heating walls of the ovens is of great importance, because, in conjunction with the air slide regulation, it determines the volume of air introduced also the velocity of flow and distribution of the heating gases through various parts of the combustion chambers.

The Roberts oven is clear of controls and obstructions from the point of admission of the gas and air to the point where the waste gases leave the recuperators. The controls are located at the points of admission of the gas and air and can be accurately adjusted. The gases are unobstructed in their flow after their admission, except for the accurately designed motion through the wall. The products of combustion then pass through the damper-controlled openings that regulate the distribution of draft. These dampers 17 are easily accessible and may be readily set to obtain the proper distribution of draft. The essential feature of this regulation is the maintenance of the proper differential from end to end of the oven necessary to move the graded quantities of heating gases; the maximum differential is maintained at the coke end and the minimum at the pusher end.

Gas Control

The method of controlling the gas supply to the headers and individual burners has been described. The effect of this control and the distribution of the gas through many burners may be summed up as follows:

  1. The number and arrangement of the points of introduction gives accurate control of the heat at all points in the oven walls.
  2. The quantity of fuel gas introduced at any one point is reduced to a minimum, and the amount of heat generated in a restricted space is proportionately small.
  3. The arrangement of these burners is such that a relatively large area of brickwork is exposed to the gases for the absorption of heat as it is generated.
  4. There is 100 per cent, excess air on the basis of the air required for the fuel gas introduced at the top or primary burners, which serves to temper the flame temperature, the entire body of gases in the wall absorbing heat and tending to attain the same temperature.
  5. The combustion of the secondary gas takes place in an atmosphere containing a high percentage of inerts (produced by the combustion of the primary gas) with the production of a low flame temperature at the point of introduction of the secondary gas and a progressive combustion with a sustained heating effect as the gases pass down through the wall.
  6. The introduction of the gas from the top, by means of the evenly distributed ducts, serves to reduce the temperature of the brickwork in the top of the ovens protecting the products of the distillation from decomposition during their travel through this portion of the oven.

Extraction of Heat from the Heating Gases

One result of introducing the gas in small quantities at many points is the production of a very small amount of radiant heat. The gas used is a lean coke-oven gas with a low percentage of illuminants. This radiant heat is extracted in the short duct at the top of the wall in which the initial combustion of the primary gas takes place.

The transmission of the sensible heat of combustion is by conduction, which will take place more rapidly if the particles of the burning gases come directly in contact with the brick. In the flue structure, only part of the gases come directly in contact with the brick, the rest sweep past the flue walls some distance from the surface of the brick so that their heat can reach the brick only by conduction through the outer layers of gas. Practically, all gases are poor conductors of sensible heat so that there is poor extraction of the heat generated by the combustion.

In the Roberts oven, the heating gases are brought into actual contact with the brick in the heating chambers. They impinge directly on top of the wall brick, slide off at a slight angle, and drop to the tier below where they again impinge directly on top of the brick in this tier. By this method, the brick is entirely surrounded by the burning gases and each part of the brick will receive a uniform amount from them. The effect is the same as the baffling of the tubes in a boiler or the checker brick in a regenerative chamber. The passages for these streams of gases are but 3 in. wide, whereas in the flue-type oven the flues are often 15 in. or more in width.

Transmission of Heat to the Coal

There are two principal factors in heat transmission through a solid medium: The temperature differential and the conductivity of the transmitting medium.

The generation and extraction of heat in the Roberts oven has been so well worked out that the heat is supplied to the heating wall progressively and it is transmitted to the coal at the same rate, thereby maintaining a constant temperature differential between the combustion chamber and the coking mass without any tendency toward excessive temperature in any zone.

The production of heat is uniform from top to bottom of the wall and from end to end of the oven. As the wall is of uniform thickness, the transmission of heat will be uniform in all parts. The absorption of heat and its transmission to the coking mass will also be proportional to the surface exposed to the heating gases and the ratio of this surface to the surface in contact with the coal. From the shape of the brick and the method of laying them, a far greater surface is exposed to the heating gases than is possible with a flue construction.

Recovery of Heat from the Waste Gases

The recovery of the heat from the waste gases, after leaving the heating walls of the oven proper, is accomplished in three ways: (1) By waste-heat boilers. (2) By regeneration, (a) common regenerators; (b) individual regenerators. (3) By recuperation, (a) common recuperators; (b) individual recuperators.

The amount of heat reclaimed from the products of combustion is dependent entirely on the temperature to which the gases going to the stack can be reduced, and at the same time have them sufficiently hot to eliminate them from the oven system without the aid of auxiliary power. The choice between regenerative and recuperative settings is a matter of individual opinion.

Heat-Effect Curves

The object of heating a coke oven is the effect on the coking mass and the more uniform the application of heat the more uniformly will the coking be carried on. To produce coke with uniform cell structure, the heat must be applied evenly and continuously throughout the entire oven. Heat-effect curves of the Roberts oven are shown in Fig. 7.

A section of the oven showing the points of admission for air and primary and secondary gas is illustrated at (a). Curve AA is the theoretically perfect curve for the application of heat to the coal charge; curve BB is the curve produced in the Roberts oven. Curve GG represents what would occur if all the gas and air were admitted at the top of the combustion space.

Curve DDEE is what would be expected if the gas and air were introduced at two points with just sufficient air at each point for complete combustion and this combustion took place in a flue. The dotted portion of curve BB represents the rise in temperature as the gas passes through the top of the oven to the mixing chamber. At this point, the gas meets the highly heated air, the volume of which is sufficient for the combustion of both the primary and the secondary gas. There is, then, a great excess of air over that required for the primary combustion, as a result the flame temperature is depressed and the curve flattens out as shown from L to M. At the point M, the burning gases meet the checkered construction, and the curve continues in a practically straight line to the point where the secondary gas is introduced.

At this point, it might be supposed that the introduction of the secondary gas would make the curve take the form EE but, as this gas is introduced into an atmosphere containing a high percentage of inert gases resulting from the primary combustion, the reaction between the combustibles and the oxygen present is subdued and the generation of heat is continued along the line of curve BB. A tendency to follow EE is prevented by the baffled construction.

There is, then, in the case of the primary gas, a reduction of the heat effect of the initial combustion because of the large excess of air present. Beyond the point M, the heat effect is strengthened by the baffled-wall structure. The initial combustion of the secondary gas overlaps the final combustion of the primary gas but is subdued by the inert gases present, and the heat effect of the latter part of the secondary combustion is strengthened by the baffled-wall structure. The curve BB, which has been accurately checked by thermocouples, therefore, follows closely the ideal curve for heat effect. It must be remembered that this heat effect is applied continuously, for there are no reversals with the attendant fluctuations in temperature as shown at RR and SS (b) and (c), which are curves taken on reversing ovens. The effect of the reversals is evident.

Roberts Individual Regenerator

The adaption of individual regenerators to the Roberts oven is possible in most forms of common practice. According to Fig. 8, which is the longitudinal section of the oven illustrated in Fig. 9, a new principle in design is incorporated in this oven regardless of whether the oven is recuperative or regenerative. This plan makes it possible to cool the

ovens to atmospheric temperature without rupturing the brickwork after it has been at working temperature.

The expansion of silica brick, such as is used in American coke-oven practice, is about 1/8 in. per lin. ft., when heated to 2600° or 2700° F.; therefore, a coke-oven wall 40 ft. long, when heated from atmospheric to working temperature expands 5 inches.

If it is necessary, because of the desirability of closing down the plant, to have the oven go back to atmospheric temperature, this wall must contract approximately 5 in. As the brick is not strong enough to withstand this contraction, shrinkage cracks result; and as there is no way of controlling the ruptures thus caused, they are irregular and form passages between the air and gas flues and between the coke-oven chamber and gas flues.

To overcome this fault, one type of Roberts oven is so designed that it can be built in sections; that is, it would be equivalent to a number of short ovens placed end to end for the required length of a complete unit. Each section can contain three or four burners and be complete within itself. As these sections are about 7 ft. long the total expansion of each section will be 7/8 in.; and as the brick will expand equally in all directions, the expansion from the center to either end of a section will be 7/16-in. Experience has shown that a coke-oven wall can expand and contract to this extent without damage, so that with this sectional construction a wall may be repeatedly expanded and contracted without developing cracks that result in operating difficulties.

This method of construction is illustrated in Fig. 8. The expansion joints are indicated at XX, YY and ZZ. These joints are intercommunicating between adjacent ovens, but are bulkheaded off from the sections in which are incorporated the combustion space and also from the center wall.

While the utilization of expansion joints in this manner permits circulation back and forth between adjacent ovens, when the ovens are first charged, this condition exists only a short time because the expansion of the brickwork will tightly close the joints and after the oven has been charged a few times any small leaks will be hermetically sealed by carbon deposits. The coking chambers will then be completely isolated from one another.

The other features of this oven are not different from those of the recuperative type. The primary difference, of course, is the admittance of gas into the heating wall at three points, the top, middle, and bottom. The fuel gas is admitted from the top and carried to intermediate and bottom burners through ducts, the same as the secondary gas is put into the wall in the recuperative oven.

In the reversal of the gas flow, this oven utilizes the dividing wall for eliminating the waste gas from the heating wall when the oven is burning in an upward direction; and, vice versa, the waste-gas ducts become air ducts when the gases are burning downward in the heating wall.

The arrangement of the checker brick in these regenerators is optional. They can be divided into zones or made two pass or single pass in a horizontal direction.

The possibility of keeping the gas ducts, particularly the secondary-gas ducts, free of carbon from the decomposition of the fuel gas passing through them is often questioned. Generally, these ducts are of such diameter that it would take many hours continuous flow of rich hydrocarbon gas to decompose sufficient hydrocarbon to cause a stoppage, and it has been found that by shutting off the gas supply in these secondary ducts at regular intervals and running air through in them for 15 to 20 min. all carbon accumulations on the walls are burned out. The common practice is pass air through these ducts once every 8 hr. When using lean, debenzolized coke-oven gas, once every 24 hr. is sufficient to keep the ducts free of carbon. The ovens are provided with an auxiliary air header (the upper pipe, Fig. 6), which contains air at atmospheric temperature at 1 lb. pressure, supplied by a rotary blower.

Results Obtained with Roberts Recuperative Oven

A plant of eighty ovens of the recuperative type has been in operation since January, 1921. During this period, these ovens have run continuously with more than satisfactory results. The plant was designed to operate on 15-hr. coking time, but when the market would absorb the products it has been operated continuously on 12-hr. gross coking time, using either Illinois or Indiana coal exclusively, or mixtures of the two coals. The resultant coke has been used for the usual purposes, such as the operation of a 500-ton blast furnace, lead smelting furnaces, water-gas practice, and foundry cupola work.

It is believed that this character of coal is not used in any other byproduct plant for the production of metallurgical coke. These coals have been used in other plants, when mixed with coking coals in which the coking coal in the mixture predominated. In the Roberts ovens, the Illinois coal has always formed the larger part of the mixture and when the Illinois coals low enough in ash and sulfur could be secured, they have been used successfully without the mixture of any coking coals.

The average operating results in this plant are as follows:

Average quantity of coal per charge per oven, tons…………….14.5
Average coking time, hours………………………………………………………….12
Average coke yield, per cent………………………………………………………..67.25
Average breeze yield, per cent…………………………………………………….3.7
Average tar yield, gallons……………………………………………………………….8.5
Average ammonium sulfate yield, pounds……………………………….27.6
Average total gas yield, cubic feet………………………………………………10,284
Average surplus gas yield, cubic feet…………………………………………4200
Average B.t.u. rich gas……………………………………………………………………..593
Average B.t.u. lean gas…………………………………………………………………….508
Average quantity of coal per oven per day, tons……………………..29

The benzol plant was put in operation on Oct. 1, 1922, and has produced an average of 2.9 gal. of motor fuel per ton of coal.

The blast furnace operated in conjunction with this plant has produced approximately 260,000 tons of pig iron on an average coke consumption of 1806 lb. The ovens have been operated with the normal crew that would be used on any type of coke oven and have shown no operating difficulties and, after practically two years of continual operation, show no inherent weaknesses from a structural or operating standpoint. There has been no charge for repairs or maintenance on the ovens and, as far as it is possible to judge, the brickwork is in as good condition as when the oven was started.

It is thought by the owners of this plant that the only replacement necessary on the ovens will be the replacement of the false bottom in the sole of the oven; this work is easily done. These false bottoms are 3 in. thick, made of silica slabs, and are held in place in the oven by the curbs on each end.

There has been no difficulty in keeping the gas ducts clean and clear of obstructions, including carbon.

The relative value of recuperators and regenerators is a matter of individual opinion. In the operation of the plant mentioned, it is still a question whether the recuperators are as economical of gas in the operation of the ovens as regenerators would be. It is impossible at the present time to get a comparison of the amount of heat required to carbonize this character of coal as compared to the use of 100 per cent, coking coals. However, if the question of costs is eliminated, recuperators made of silica brick can be constructed that will be as efficient in heat recovery as regenerators. The men operating this plant favor the recuperator because of the elimination of the reversing feature and continuous flow of heat in one direction. As most of these men have had considerable experience in the operation of the reversing type of oven, considerable weight should be given their opinion.

While it is true that more brick is used in the construction of this type of oven than in the single-wall type, the added cost is offset by the greater capacity of the oven. Several demonstrations have been made with these ovens in which it has been proved that a 14-in. oven of this construction can be operated continuously on 10.5-hr. gross coking time. As the tendency in the United States for the last few years has been to build ovens with greater cubical content and operate them at faster coking times, based on the results obtained by the use of the Roberts oven, it is believed that it will be only a short time when ovens that will carbonize 20-ton charges of coal in from 10 to 12 hr. will be in use.

Zinc Flotation Concentrate Roasting Furnace

This paper describes experiments carried on at the Case School of Applied Science, together with their results. Their success led to the design of the larger furnace herein described, but which has not been built.

A previous article by the authors contained a general description of the new roasting furnace herein described but it did not go into detail as to the metallurgical behavior or the results obtained. Believing that such information would be of great value, they have elaborated on the subject and have given many unpublished details.

The furnace described applies the principle of roasting finely divided zinc-sulfide ores, now produced in large quantity by the flotation process, in gaseous suspension; that is, the ore particles are carried, in suspension, in a current of air and gaseous products of the roast. The relatively great fineness of flotation concentrate presents difficulties and problems of roasting in furnaces of the ordinary type; the fine ore is forwarded through the furnace in the form of a shallow bed and its very fineness leads to dense impervious bedding which prevents oxygen from reaching the interior of the bed, thus unduly lengthening the time of roasting and preventing the elimination of the last of the sulfur. The fineness of the concentrate, normally, should lead to a rapid and complete roast, for the speed of roasting is a function of the surface exposed to oxygen, which surface is greatest, per unit of weight, in very fine material. The difficulty in bed roasting is to get the oxygen to the particle. If the fine ore, during the roasting, could be freely suspended in oxidizing gases, full advantage could be taken of the great surface conferred by its fineness. This fundamental idea, of course, is not new, for the Stedtefelt furnace, familiar to the older metallurgists, is an example of it; but the manner in which this is accomplished may be new.

The numerous efforts to roast in gaseous suspension show that the idea is attractive; in fact, an analysis would indicate that it is the most reasonable way to effect the oxidation of ore, provided certain difficulties can be overcome. Two objections against such a method that formerly had much weight were the cost of fine grinding the ore, also the fact that the finely ground product, even after roasting, was not the best condition of material for further metallurgical operations. The objection of costly grinding, however, has been removed by the production of great quantities of flotation concentrate, which in point of fineness present a material that is ideal for roasting in gaseous suspension.

The inception of the experiments described here is due to David B. Jones and March F. Chase. The idea of the general type of furnace and process was suggested to the authors and the experimental work was carried out in the metallurgical laboratories of the Case School of Applied Science at Cleveland in 1915-16. The original plan was to make a furnace for roasting zinc-blende flotation concentrate that would use the heat value of the sulfide to accomplish the roast and to produce a gas suitable for making sulfuric acid; i. e., of sufficient concentration in SO2 and practically free from the products of carbonaceous combustion. Autoroasting of sphalerite is theoretically possible, and also practically, as was demonstrated in the experiments set forth.

A diagrammatic drawing of the roasting furnace and accessory apparatus as erected is shown in Fig. 1. Here A is the furnace proper; B the two stoves, heated by natural gas, that preheat the air; and C, the Cottrell electric precipitator for the precipitation of flue dust and fume carried from the furnace by the gases, which in a commercial plant would pass to a sulfuric-acid plant. At H is the gas supply for the stoves; X is the dust chamber, and G is the cycloidal blower that furnishes air to carry the ore in suspension in the furnace. The dried, preheated ore (60° to 100° C.) is charged into the hopper 1 whence the endless screw 2, operated by a variable-speed electric motor 3, discharges it into a pipe 4, directly above the nozzle 5. A stream of high-pressure (20 to 60 lb.) moderately pre-heated air is discharged through this nozzle in such quantity as to carry readily the fine ore in suspension into pipe 7, on the injector principle. Pipe 7 is of larger diameter than pipe 4 and is lined with refractory material; it serves as the main injector pipe into the furnace. Air, heated to approximately 800° C., passes from one of the stoves B to the injector pipe 7, through the supply pipe 9; this is the main air supply for roasting the ore.

The amount of air supplied is governed by two principles: (1) The quantity must be correct to roast the ore and to furnish a gas of the correct composition for the manufacture of sulfuric acid. Every pound of

sphalerite requires 35.7 cu. ft. of air (standard conditions) to convert the zinc to oxide, the sulfur to dioxide, with enough more oxygen to convert this into the trioxide. Some excess must be carried; in the experiments from 41 to 55 cu. ft. and sometimes 75 to 100 cu. ft. were used, as measured by a meter. (2) There must be such a relation between the quantity of air per minute and the area of the riser tube 10, that the velocity of the ascending air current will be enough to carry the largest ore particle to the top of this tube and over the edge. Fig. 2 gives definite data on this point for sphalerite. Fortunately the requirements for both conditions can readily be fulfilled.

The mixture of ore and high-pressure air from pipe 4 enters the main injector pipe 7 with a rotary motion and is caught by the ascending hot

air from the pipe 9 and injected into the riser tube 10. The fundamental idea is to have the temperature of the ore-air mixture at the ignition point of zinc sulfide, which varies between 650° and 810° C., depending on the size of the particle as it leaves the injector tube to enter the riser tube, so that the full time of the ore particle in the furnace will be available for oxidation. After leaving the riser tube 10, the ore particle falls in the annular space 11, the area of which is large so that the natural velocity of fall shall not be augmented by the velocity of the descending gases. It is the belief of the authors that the gas envelope surrounding the ore particle is constantly changing; thereby causing fresh oxygen to be supplied to the particle.

The time necessary for the roast is furnished by the passage of the ore up the combustion tube 10 and its fall in the annular space 11. This time depends on the height of the furnace, the length of the path of travel being practically twice’ the height of the furnace, also on the velocity of the ascending air current in the combustion tube. Assuming a definite ratio of cubic feet of air per pound of ore, the velocity will be determined by the area of the combustion tube. Definite figures on this point are given later. In any given furnace, i. e. a fixed area of combustion tube, there is a certain variation allowable in the velocity of the ascending gas current, obtained by varying the amount of air, which will then vary the time of the ore in the combustion tube. Too high a velocity, obtained either by too much air only or too great an ore feed with its corresponding increased amount of air, will shorten the time element so much that the sulfur is not sufficiently eliminated. The greater part of the ore collects in the hopper 14; the gases and fine dust pass, through openings 12, into the flue 13, thence to the settling chamber X, where the coarser dust is settled out, thence to the Cottrell precipitator C for the precipitation of the finest dust and fume.

Thirty-Foot Experimental Furnace

The design of the 30-ft. experimental furnace was, as far as possible, based on the results obtained in the previous work but was limited by lack of space and the capacity of the stoves, blowers, gas supply, etc. The stoves had been erected for the small furnace already described and were known to be inadequate for the larger furnace, but no space was available for enlarging them. It was also desirable to keep the expense as low as possible, so that some things that would have aided in the work were omitted.

The furnace consisted essentially of a brick stack (Fig. 3) approximately 30 ft. in height from the base plate to the top of the cover arch. The base plate and support rested on 4-ft. reinforced concrete posts, thus making the total height about 35 ft. The space between the posts provided room for the calcine hoppers and air and feed inlets and a pit provided the additional space necessary for the cleaning of tools and for making repairs.

The stack was built of two concentric firebrick circles 24 and 36 in. internal diameter; the space between these circles was filled with mineral wool. Somewhat larger circle brick were used near the bottom of the stack to increase the stability. No steel shell was placed outside of the brick, though such a shell would have aided greatly in the construction

of the furnace and in carrying on the tests. The top of the stack was closed by fireclay slabs, a small circular hollow tile being set at the center; a gate at its top permitted the interior of the furnace to be watched while operating. Two hoppers at the bottom of the stack collected the greater part of the roasted ore. Above these hoppers, and spaced at equal intervals around the stack, were four openings to the flue that led to the settler and to a small Cottrell precipitator.

Concentrically with the walls of the stack was placed the riser, or combustion tube, which extended from below the hoppers to within about 5 ft. of the furnace top. This tube, in the earlier trials, was made of

9-in. hexagon stove tile but later 12-in. cylindrical tile was used. The annular space between the combustion tube and the stack walls served as a downtake for the products of the roast.

The injector nozzle entered the bottom of the combustion tube for about 5 ft., thus placing its delivery point about 2½ ft. above the base plate of the furnace. The nozzle was a 2-in. fireclay tube and connected at its lower end with a pipe or mixer, into which the air and ore were injected from the stoves and feeder respectively.


The two stoves were of the central combustion type, the combustion tube being made of 9-in. hexagonal stove tile. The checkers were made of 2 by 2 by 9-in. brick laid to provide the maximum heating surface. A 1/8-in. steel shell with top and base plates insured airtight conditions. The stoves were 4 ft. in diameter, and 10 ft. in height. The usual burner, cold-blast, hot-blast, and flue openings were provided; the stoves were heated with natural gas.

Hot-blast mains connecting the stoves were 6-in. iron pipe lined with 4-in. fireclay flue liners, and covered with heavy asbestos insulation. Connections were provided between this pipe and the cold-air supply so that the temperature of the air entering the furnace was always under control. A bypass was provided at one point so that the air could be sent through a rotary gas meter at intervals; the metering was done cold and corrections made to temperature and pressure.


The feeder used was a modified Dunn pulverized coal feeder. Certain parts, originally made of brass, were later replaced by high-speed, hardened, tool-steel parts to avoid excessive wear from abrasion by the ore. A screw conveyor carried the; pulverized ore from the bottom of a hopper to a pipe in which a suction was produced by a compressed-air jet issuing from a 1/8-in. or 3/16-in. nozzle past its lower end. The ore was carried, by an expanding compressed-air current, toward the mixer and injector pipe. The connection into the mixer pipe was made tangentially and also inclined upwards so that the ore and air mixture met the hot air from the stoves in a rising spiral and thus thoroughly mixed the hot air and ore before they entered the combustion tube.

Approximately 10 cu. ft. of free air compressed to 60 lb. per sq. in. were necessary per pound of ore to carry the ore from in front of the 1/8-in. or 3/16-in. nozzle into the mixer. At times, even this amount was insufficient to prevent stoppage of the pipe, so the amount of air was increased at intervals to remove any accumulations by the arrangement shown in Fig. 5.

Seven thermocouples were placed in the furnace: four being placed in the combustion tube as shown, one in the mixer just below the base of the injecting nozzle, one in the downtake at the exit to the flue, and one in the downtake 20 ft. from the base plate. Lead wires were run to a central galvanometer station where readings were taken throughout each test.

General samples from the products accumulated during the run showed these sulfur contents:

This run began 1 hr. before that indicated (9:55) but conditions did not permit sampling, and adjustments caused unreliable results previous to those shown.

Total ore fed during the entire run, 2080 lb.; of this 160 lb. were fed before beginning the record.


Openings for oil burners were provided at intervals of 8 ft. throughout the height of the furnace. These openings were placed tangent to the inside circumference of the downtake. The burners were for the purpose of bringing the furnace to the operating temperature and were not used after such a temperature was reached and feeding of ore had begun.

The products of combustion of the oil heating were taken off at the top until the burners were going well, when the top was closed and the draft directed through the flues near the bottom of the stack. After a temperature of 800° to 1000° C. in the furnace had been reached, the burners were shut off, the burner openings closed, and the ore and air feed started. Adjustments of ore and air feed were then made as indicated necessary by the SO2 content of the gas, temperature conditions, and quality of roasted product. The furnace responded readily to such adjustments, but the time permitted for the test was not sufficient to permit adjustments to the best results possible.


In order to increase the capacity, the 9-in. tube was replaced by a 12-in. This tube was of specially made tile with 1½-in. walls, which restricted the area of the downtake more than was desirable and probably resulted in slightly poorer results than could have been obtained if the diameter of the furnace had been increased in proportion to the increase in the diameter of the combustion tube. Such a change, of course, was impossible without completely rebuilding the furnace.

Table 1 gives the results of a test made with the 12-in. tube. The temperature at the top of the furnace increased beyond that which was desirable, so in the last run cold air was admitted at 20 ft. from the base plate, thus providing for a complete control of the temperature and preventing the hot top. When cold air was thus admitted, it was necessary to reduce the amount of air fed with the ore, which meant a lowering of the velocity in the combustion tube and resulted in an increase in the bottom temperature. This increased bottom temperature, in turn, permitted the reduction in the temperature of the air coming from the stoves.

It was thought that accretions would form in the combustion tube during the roasting but, with one exception, no trouble was encountered from this source, even though the temperatures at times were allowed to go beyond those that would be permitted in practice. Some little difficulty developed at points near the oil burners, as in preheating the temperature in these regions was necessarily higher than the intermediate zones. The temperature of the various zones, however, were quickly equalized and no trouble was encountered during the first few tests in the furnace, the combustion tube being clean at the end of the test; in later tests, however, after some dust had collected in crevices of the tube, the overheating at the burners caused some slagging between the dust and fireclay, thus developing starting points for accretions. It is not thought that this would cause any trouble in practice for, after the furnace is in operation, it will continue so for a considerable time. In this experimental work, the operations were necessarily short and at infrequent intervals, which meant repeated heating of the furnace with consequent trouble from slagging.

The gases and fine dust were taken from the furnace into the flues through four openings equally spaced around the base of the stack. The flue led first into a settling box, in which the rate of flow of the gases was reduced and screens prevented the channeling of the currents. From this settler, the bases passed through a small Cottrell precipitator for final cleaning. The cleaning was very effective, though this end of the operation was given but little attention. The Cottrell equipment contained four 10-in. tubes 16-ft. long which were thoroughly grounded. In the center of these 10-in. pipes, and connected with the 50,000-volt line, were suspended, on insulators, ½-in. pipes fitted with four knife edges at right angles to one another. The 400-volt a.c. current was stepped up to 50,000 volts and rectified by a Kenetron.

The flotation concentrate used in the furnace was a complex sulfide containing 31.4 per cent, sulfur, 44.3 per cent, zinc, 11.6 per cent. iron. The results of the roasting of this material and conditions existing during the roast are given in Table 1. These show a minimum sulfur content in the roasted product of a 2.2 per cent, and 3.6 per cent, for an afternoon period after the furnace had been adjusted to best running conditions. Other tests gave continuous results of slightly over 2 per cent, sulfur.

Tables 2, 3, and 4 show the forms in which the sulfur existed in the roasted product. The particular point of interest in these results is that the sulfur existed largely as sulfate and the quantity as sulfide was small. Table 2 shows the per cent, of SO2 in the gas corresponding to the roasted ore samples. It is evident that the concentration of gas had no effect in preventing the completeness of the roast; i. e., the sulfur is not higher in roasted products produced in higher SO2 concentration.

Note.—Decreased sulfur content of calcines in the afternoon results from the higher furnace temperatures, as shown by Table 1.

Note.—During this run the SO2 content of the gas in the furnace was fairly constant at 9 to 9.5 per cent, during the greater part of the run.

Hopper sample is calcine collected from the hoppers of the furnace; flue sample is calcine cleaned from the horizontal flue leading to the settler; settler sample is calcine taken from the settler; precipitate or sample is calcine taken from the Cottrell precipitator. The coarsest material was from the hopper and the finest from the precipitator.

Tables 3 and 4 show that the sulfur exists in the roasted product as sulfide and sulfate, and that the sulfate sulfur is predominant in the fine ore. This is shown not only by the general samples, of which the hopper samples are the coarsest and the precipitator the finest, but also by the screen analysis of the roasted products. High sulfide sulfur is present only in coarse material, which in total amount is insignificant in flotation concentrate. The total sulfur is highest, in general, in the finest product, but consists largely of sulfate sulfur. Work that cannot be detailed here reveals that this sulfate sulfur is not the result of an incomplete roast m the furnace, but results from a resulfatization of the roasted ore. This resulfatization is due to relation between partial pressures of SO2 and SO3 and the temperature existing in the lower part of the furnace, in the settler, and in the precipitator. Generally, the partial pressure of the sulfur gases is such that no resulfatization of the calcine can take place if the furnace temperature is between 900° and 1000° C. But when this temperature falls to near 800° and below at the flues, the rapid resulfatization of the ore commences. For example, it will be noted that in Tables 3 and 4, the minus 200-mesh material of the hopper samples, i. e., material separated from the furnace at high temperature and immediately cooled, is much lower in total sulfur, as well as sulfate sulfur, than material of the same size from the flue, settler, or precipitator samples, which were subject to the lowering critical temperatures in the presence of sulfating gas. When the aim of the roasting is a practically complete desulfurization, the furnace must be run hot and the calcines and gas rapidly cooled, or separated from each other as soon as they leave the furnace. If the roasting is carried out as a preliminary for sulfuric-acid leaching, as in the electrolytic-zinc process, the sulfating action of the furnace may be used to any desirable degree up to its maximum.

The Proposed East St. Louis Furnace

An increase in height was provided for in the design of the 12-ton furnace it was planned to build at East St. Louis, but which was not completed because of the conditions brought about by the war. Fig. 6 shows that the combustion tube is set to one side instead of concentrically with the downtake. With the concentric combustion tube, it was necessary to place the furnace on posts to provide space for feed and air connections and for repairs. It was also difficult to make pyrometer and air connections from the outside to the combustion tube, at intervals throughout the height, and to provide support for the combustion tube. To overcome these objections, the side combustion tube was designed.

Air and Feed Arrangement

Provision was made for bringing all the air for the roasting from the stoves through the mixer, as in the smaller furnace, or for splitting the current and having a part of the air take this path and the remainder enter the combustion tube near the top of the injector nozzle. Pipe connections to hot- and cold-air sources were placed at intervals along the combustion tube, thus providing for an absolute control of the

temperature in this tube; Fig. 7 illustrates this arrangement of the ore and air feed connections.


As in the smaller furnace, a screw conveyor takes the ore from the bottom of the feed hopper to a pipe immediately above a compressed-air

injector. The injector was changed considerably from that of the smaller feeder, in order to reduce the amount of compressed air necessary for injection purposes. As shown, the compressed-air nozzle blows directly
through a throat, or Eureka siphon, carrying the ore with it and delivering it into a current of low-pressure air coming from hot or cold sources, or both, by which it is carried forward into the mixer.

Experiments indicate that not to exceed 20 cu. ft. of free air compressed to 30 lb. will be necessary for injecting 1 lb. of ore as against 10 cu. ft. of free air compressed to 60 lb. in the feeder used on the small furnace.

The delivery pipe from the feeder was to be connected to the mixer pipe by the tangential and rising connection that was used on the smaller furnace. The same principle was also to be used in connecting the excess-air pipe into the combustion tube.


The bottom, or base, plate of the combustion tube was to be raised about 10 ft. from the floor line, to provide for feeder and air connections. The space opposite these connections in the base of the downtake was to be used for the settling chambers for the roasted ore. Accordingly provision was made for the removal of the coarse product from above a bulkhead in the downtake at a point opposite the injector nozzle. The fines and gas were to be taken off, somewhat above this, through three radially placed flues, which flues in turn entered tangentially a settling chamber placed beneath the bulkhead just mentioned. This method of entry was to provide a circular or centrifugal motion for settling the dust. The gases and remaining dust were to be taken off through a flue passing upward and outward through the bulkhead, thence to the Cottrell precipitator.

Heating the Air

There is insufficient data as to how much of the air must be heated and to what temperature heating should be carried for the most efficient operation of the furnace. The effect of heated air on the behavior of the furnace and the ignition has been discussed and it is obvious that one of four methods must be used.

  1. Heating all the air and passing all of it through the mixer tube and into the furnace with the ore; this will require large stove construction and will not allow air changes in various parts of the furnace, as will be desirable for proper control.
  2. Heating all the air and passing only a part of it into the furnace through the mixer, the remainder to be introduced into the combustion tube at various points; this method is not desirable as the heat is needed in the bottom of the furnace for. heating the ore and bringing about early ignition.
  3. Heating none of the air; indications are that this will not be possible, as already explained.
  4. Heating a part of the air (say one-fifth), which will be passed into the furnace with the ore through the mixer, the remainder of the air to be admitted cold at points throughout the height of the furnace; this method will make all the stove heat available for ignition and for holding down the combustion zone, and will permit the use of cold air for combustion and temperature control of the top of the furnace. Of course, the stoves should be of ample capacity and of such construction as to permit changing the temperature and amount of air going into any part of the furnace as may be found necessary.

Under the first condition, assuming that 300° C. is necessary for mixer temperature, the air in a 50-ton furnace will have to be heated to 350°
in order to heat the ore, if it is being fed cold. If the ore is taken direct
from the dryer, this temperature may be reduced somewhat.

The authors calculate that three stoves of the following dimensions will readily take care of a 50-ton furnace: 1060 sq. ft. heating surface, 246 flues, 4 by 4 in. in cross-section, 14 ft. long, and 3¼-in. walls; total height allowing for gas and air ports, foundation and dome about 25 ft.; diameter about 12 ft.;.cross-section of combustion chamber 12 sq. ft.

Under the fourth condition, which seems the most rational, such stoves would be more than ample, but not necessarily inefficient, and would provide for emergencies where the heating of all the air was necessary. The ratio of fuel necessary for heating all the air to 250° to that necessary for heating one-fifth to 600° is about as 2.5 is to 1, which is quite a saving in favor of the fractional heating.

Drying the Ore

It is obvious that the feed for this type of furnace must be dry, or it will not pass the feeder and injector. It is likely that most of the concentrate delivered at the roasting plant would have to be dried, so it was planned to do this drying with the waste heat in the gases leaving the furnace, by circulating them beneath the hearth of an ordinary kiln dryer.

That there is sufficient waste heat available for this purpose is readily shown. In the experimental furnace, the gas left the furnace at 800° C. under conditions of best roasting. The volume of this gas (measured under standard conditions) was approximately 50 cu. ft. per lb. of ore roasted and was made up of about 8 per cent, sulfur dioxide; 10 per cent, oxygen, and 82 per cent, nitrogen.

Assuming that these conditions are obtained in a 50-ton furnace:

50 tons per day = 69.5 lb. or 31.5 kg. per min.
69.5 lb. will produce 3475 cu. ft. per min. of gas.
8.0 per cent, of 3475 = 278 cu. ft. SO2 = 22.7 kg. per min.
10.0 per cent, of 3475 = 347.5 cu. ft. O2 = 14.1 kg. per min.
82.0 per cent, of 3475 = 2849.5 cu. ft. N2 = 101.7 kg. per min.

Further, assuming that the gas is lowered in temperature by 400° in passing the drier, the heat available there is:

  1. Sulfur dioxide:
    Heat capacity at 800° C. is 1 X 800 (0.125 + 0.0001 X 800) = 164; 164 X 22.7 = 3722.8 kg.-cal.
    Heat capacity at 400° C.
    66 X 22.7 = 1498.2 kg.-cal.
    Heat available from SO2 = 2224.6 kg.-cal.
  2. Oxygen:
    Heat capacity at 800° C. per kg. = 181 cal.
    181 X 14.1 = 2552.1 kg.-cal.
    Heat capacity at 400° C. per kg. = 88 cal.
    88 X 14.1 = 1240.8 kg.-cal.
    Heat available from oxygen = 1311.3 kg.-cal.
  3. Nitrogen:
    Heat capacity at 800° C. per kg. = 207 cal.
    207 X 101.7 = 21051.9 kg.-cal.
    Heat capacity at 400° C. per kg. = 100 cal.
    100 X 101.7 = 10170.0 kg.-cal.
    Heat available from N2 = 10881.9 kg.-cal.
    Total heat available per minute;
    2224.6 + 1311.3 + 10881.9 = 14417.8 kg.-cal.

Assuming 20 per cent, moisture in the ore, it will require 39.3 kg. of moist ore to provide the 31.5 kg. of dry ore. This means that 7.8 kg. of water must be evaporated each minute, and 39.5 kg. of ore must be raised to 200° C.

It is, therefore, evident that the necessary heat is available for drying and heating the ore. If this heated ore is fed without cooling to the furnace, the temperature of the air from the stoves may be reduced somewhat from that necessary when cold ore is fed. As this amount of waste heat was actually delivered by the small experimental furnace, it is clear that enough heat would be available for completing the roast and maintaining furnace temperatures. As this was possible with a small furnace with a high radiation factor, it would be more readily accomplished in a larger well-insulated furnace.

The dried ore from the dryer will contain some lumps, formed in the drying, and some coarse particles that will not work well in the feeder; it was planned, therefore, to interpose a vibrating screen between the dryer and the feed hopper to remove these, insuring a satisfactory feed at all times.


The capacity of these furnaces will vary as the square of the diameter of the combustion tube. This is substantiated in the results from the 9-in. and 12-in. combustion tubes used in the experimental furnace, which had capacities of 2.0 and 3.6 tons per 24 hr., respectively. Assuming similar conditions as to ratio of air and ore, and the velocity in the combustion tube, the diameter necessary in the tube of a 50-ton furnace will be 45 in. It may be that for a furnace of this capacity it would be better to use two tubes having a combined area equal to 45 in. diameter, i. e., 32 in. diameter each. Such an arrangement would provide for repairs in feeders, nozzles, and injectors on a part of the furnace without complete shutting down the furnace, with attendant cooling.


Effect of Oxygen-enriched Air in Roasting Zinc Ores

Experiments have shown that the use of enriched air would be of particular benefit in the roasting of zinc ores for the manufacture of sulfuric acid. Enriched air increases capacity of furnace, decreases fuel consumption, and increases SO2 content of roaster gas.

The work here described had for its purpose the procuring of data from which some quantitative estimate might be made of the results obtainable by using oxygen-enriched air in roasting zinc ores on a commercial scale. The principal metallurgical advantages of using enriched air in roasting zinc ores would be:

  1. The rate of roasting would be increased, with consequent gain in the capacity of the roasting furnace.
  2. As less air would be required for roasting, the volume of hot gases leaving the furnace and the heat carried out of the furnace as sensible heat in these gases would be less per ton of ore roasted; partly for this reason and partly because of the increased quantity of heat generated in the furnace by the larger amount of ore that could be roasted, the consumption of fuel by the furnace would be lessened; and by the use of air sufficiently enriched with oxygen the necessity of using fuel might be entirely obviated.
  3. Roaster gas having a higher SO2 content could be produced; this would make possible greater capacity and more economical operation of the sulfuric-acid plant.

Certain phases of the application of enriched air to roasting can be worked out only by experimenting with a furnace of commercial, or at least semicommercial, size. Thus the precautions necessary to secure proper distribution of heat in the furnace; the volume of enriched air, and the proportion of oxygen in this air, necessary to give the desired increase in roasting capacity of the furnace and in SO2 content of the roaster gases; and the most desirable frequency of raking, thickness of ore bed, and rate of advance of the ore through the furnace can be definitely determined only after such large-scale experiments.

On the other hand, by drawing up suitable heat balances, the amount of additional heat, per ton of ore roasted, made available in the furnace by the use of enriched air can be calculated, and from that can be calculated, if it is assumed that proper distribution of the total heat in the furnace can be arranged, the additional amount of ore that must be roasted per unit of time in order to make the use of fuel unnecessary. Furthermore, data concerning the effect of enriched air on the ignition temperature and rate of oxidation of zinc ores can be obtained by means of laboratory experiments in which all other conditions (such as temperature, volume of air supplied, thickness of the ore bed, size of; the ore particles, and frequency of stirring) can be maintained constant. From these data, estimates may be made of the increased capacity that may be expected in a full-size furnace, and of the increase in SO2 content of the roaster gas that may be expected, as a result of the use of enriched air.

Heat Balances of a Hegeler Roaster Using Ordinary Air and Using Enriched Air

Heat Balance of a Hegeler Roaster Using Ordinary Air

When the suggested use of enriched air for roasting zinc ores was first called to the attention of the writers, they drew up a heat balance of a Hegeler roaster using ordinary air; this particular type of furnace was selected because it is the furnace generally used in this country for roasting zinc ores when the gases are to be utilized for making sulfuric acid. The heat balances of different Hegeler roasting furnaces will vary in detail, depending on the design of the furnace, composition of ore roasted, kind of gas producers used, quality of coal used, and manner of preheating the air for roasting, but the variations are in the minor items; the net result, as shown by the consumption of coal per ton of ore, is nearly the same in most furnaces of this type.

The heat balance here given is for a hypothetical case, in that the operating, data were not taken from the actual operation of any one particular furnace. The conditions assumed were, however, representative of actual practice, so that the heat balance is typical. A few simplifying assumptions were made, such as assuming an ore consisting entirely of ZnS, FeS, and SiO2, minor constituents being neglected. The effect of such simplifying assumptions on the accuracy of the calculations is negligible.

This heat balance of a Hegeler roaster, using ordinary air and roasting 45 tons of 60 per cent, zinc ore per day, is summarized in Table 1.

Heat Balance of a Hegeler Roaster Using Enriched Air

In the heat balance of a Hegeler roaster using enriched air, if it be assumed that this use of enriched air is to eliminate the use of fuel, several of the items enumerated in Table 1 will be absent. These are, from the debit side, the sensible heat in the preheated air (as there will be no waste combustion gases for preheating), the sensible heat in the producer gas, and the heat of combustion of the producer gas; and from the credit side, the sensible heat in the combustion gases. There remains then, as a source of heat in the furnace, only the oxidation of the ore, which must balance the heat lost as sensible heat in the roasted ore and in the roaster gases leaving the furnace, and that lost by radiation and conduction.

Assuming the composition of the ore before and after roasting, the temperature of the green ore and of the air supply entering the furnace, and the temperature of the roasted ore and roaster gases leaving the furnace, to be the same as they were assumed for the purpose of calculating the heat balance of the roaster using ordinary air, and assuming that the loss of heat from the furnace by radiation and conduction would remain constant, the tonnage of ore that would have to be roasted per 24 hr. to maintain the furnace at roasting temperature without the use of other fuel was calculated for the following cases:

Case 1.—The enriched-air supply to contain 25 per cent, oxygen; the exit gases to contain two volumes of SO2 to one volume of O2 (in this case 13.3 per cent. SO2 and 6.65 per.cent. O2); 60 per cent, of the roaster gases to be recirculated and returned to the furnace at 200° C. to help in controling the rate of combustion and the distribution of the heat in the furnace and in procuring the desired high content of SO2 in the gases. The flow sheet under these conditions is shown in Fig. 1.

Without going into details, the calculations may be summarized as follows: From the combustion of 1000 lb. of the ore are obtained 1,049,650 lb.-cal. The heat leaving the furnace as sensible heat in the roasted ore (853 lb.; see flow sheet) at 800° C. is 106,350 lb. cal. and in the roaster gases (63,140 cu. ft.) at 600° C., 775,888 lb.-cal.; of the latter, 142,644 lb.-cal. are returned to the furnace in the recirculated roaster gases (37,884 cu. ft.) at 200° C. Thus 1,049,650 – 106,350 – 775,888 + 142,644 = 310,056 lb.-cal. are available per 1000 lb. of ore roasted, to balance radiation and conduction losses.

From the previously calculated heat balance of a Hegeler roaster using ordinary air, it was found that the loss of heat from the furnace by radiation and conduction was 111,696,200 lb.-cal. per 24 hr. From this, it follows that 111,696,200/2 x 310,056 = 180 tons of green ore must be roasted in the furnace per 24 hr. to maintain it at the usual roasting temperature, using enriched air under the conditions assumed in this case.

Case 2.—Conditions the same as in Case 1, except that none of the roaster gases are recirculated; or, what amounts to the same thing thermally, that the recirculated gases are returned to the furnace at the same temperature as that at which they leave. Under these conditions the flow sheet is as shown in Fig. 2.

The heat obtained from the combustion of 1000 lb. of the ore is, as before, 1,049,650 lb.-cal., and the heat leaving the furnace as sensible heat in the roasted ore is 106,350 lb.-cal., the heat leaving the furnace as sensible heat in the roaster gases is, however, only 40 per cent, of what it was in the former case, or 310,355 lb.-cal. The heat available to balance radiation and conduction losses is then: 1,049,650 — 106,350 — 310,355 = 632,945 lb.-cal., and 111,696,200/2 x 632,945 = 88 tons green ore must be roasted per 24 hr. to maintain the furnace at the usual roasting temperature, using enriched air under the conditions assumed in this case.

Case 3.—The enriched air supply to contain 50 per cent, oxygen; the exit gases to contain two volumes of SO2 to one volume of O2, which in this case means that the SO2 content will be 28.6 per cent, and the O2 content 14.3 per cent.; none of the roaster gases to be recirculated. The flow sheet under these conditions is shown in Fig. 3.

In this case, the heat leaving the furnace as sensible heat in the roaster gases at 600° C. is, per 1000 lb. of ore roasted, only 155,178 lb.-cal. and the heat available to balance conduction and radiation losses is
1,049,650 – 106,350 – 155,178 =788,122lb.-cal. Therefore, 2×788 122 = 71 tons of green ore must be roasted per 24 hr. to maintain the furnace at the usual roasting temperature.

The SO2 content of the roaster gases in the examples just discussed was purposely assumed to be very high. If it should be impracticable to obtain such a high SO2 content in the roaster gases, a larger amount of ore would have to be roasted to produce the same amount of available heat in the furnace. Even with the SO2 content assumed as high as it has been, the increase in the amount of ore that must be roasted per 24 hr. in order to dispense with the use of file] is considerable.  It is open to question whether such high SO2 content in the roaster gases, with simultaneous large capacity of the furnace, could be accomplished except by the use of enriched air containing a very high percentage of oxygen.

On the other hand, it might seem possible that the ignition temperature of zinc blende in enriched air would be so much lower than in ordinary air that the roasting furnace would not have to be run at so high a temperature when enriched air is supplied as when only ordinary air is available; this would permit a saving in the heat lost by radiation and conduction, and as sensible heat in roaster gases leaving the furnace. To obtain experimental evidence bearing upon these questions, the series of experiments here described was undertaken.

Ignition Temperatures of Sphalerite in Air Enriched with Various Proportions of Oxygen

In giving data on ignition temperatures, it is necessary to define exactly what is meant by the term ignition temperature. The usually accepted meaning is the temperature at which the oxidation of a substance becomes so rapid that the heat liberated counterbalances the heat radiated or conducted away, thus maintaining a visible spontaneous combustion. The temperature at which this can take place varies with the rate at which heat is radiated or conducted away from the substance; this in turn is affected by the heat conductivity of the walls of the vessel in which the substance is contained and by the volume and temperature of air circulated over it. ”

Oxidation of sphalerite exposed to the air, no doubt, takes place at an extremely slow rate, even at ordinary atmospheric temperatures. It is conceivable, if a pile of finely divided zinc blende could be so insulated that radiation and conduction from the pile would be nil and the air supply so regulated that the heat carried off by it as sensible heat would be as small as possible, that the slow oxidation of the blende would, of itself, cause the pile to become sufficiently hot for active combustion to take place. This would be analogous to the spontaneous ignition of large coal piles. In such a case, it would be difficult to say just what should be called the ignition temperature.

In outlining the series of determinations of the ignition temperatures of sphalerite in air enriched with various proportions of oxygen, it was at first planned to heat slowly a sample of the mineral in an electrically heated tube, passing the enriched air over it at a fixed rate; to read the temperature of the sample at intervals by means of a thermocouple, the junction of which was placed in the sample; and then to plot a curve of the rate of temperature rise. It was thought that at the temperature of ignition there might be a sufficient increase in this rate to cause a noticeable deflection in the curve; it was found, however, that the oxidation of the sphalerite began so gradually that no such deflection could be detected.

It was then decided to determine the temperature at which sufficient sulfur dioxide was formed to cause to turn blue a solution of potassium iodate and starch placed at the exit of the tube containing the sample. The ignition temperature, even as determined by this method, varied according to the rate at which the sample was heated, the rate at which the air was passed over the sample, etc., but by heating very slowly and keeping the rate of heating and other variable factors the same in all the experiments, comparative results were obtained that show clearly the effect on ignition temperature of increasing the oxygen content of the air supply.

Apparatus and Procedure

The oxygen-enriched air for use in the experiments was made by mixing commercial oxygen with ordinary air in a large gas-storage bottle. The pressure in this storage bottle was regulated by raising or lowering a pressure bottle of the same size filled with water, which was placed on a small elevator and connected to the storage bottle by a flexible siphon. Before the gas was passed over the sphalerite; it was passed through two washing bottles containing, respectively, sodium-carbonate solution and distilled water, and through two drying tubes containing anhydrous calcium chloride.

The sample of sphalerite was placed in a pyrex glass combustion tube 20 mm. in diameter, which could be heated by means of a nichrome-wound electric-resistance furnace. There were two sections of this furnace, one of which was used to heat the sphalerite and the other to preheat the air so that it would have about the same temperature as the sphalerite before coming into contact with it. The temperature of the sphalerite was read by means of a thermocouple, the junction of which, protected by a thin quartz tube, was placed in contact with the surface of the sample. The temperature of the preheated air was read by means of a second thermocouple. One end of the combustion tube was connected to the supply of enriched air; the other (exit) end to a small washing bottle containing a few cubic centimeters of a solution of potassium iodate and starch, to serve as an indicator for sulfur dioxide. Beyond this bottle of indicator solution, there was attached a flow meter for measuring the flow of air through the system.

The sphalerite used was a hand-picked specimen of the massive mineral. It had the following analysis: zinc, 65.17 per cent.; sulfur, 32.36 per cent.; iron, 0.48 per cent.; insoluble, 0.79 per cent. As the size of the particles of the sphalerite has a marked effect on the ignition temperature, the crushed sample was separated by screening into four sizes: through 20, on 28 mesh; through 28, on 35 mesh; through 35, on 100 mesh; and through 100 mesh; and a separate series of experiments run on each size.

A 15-gm. sample of sphalerite was used for each experiment. It was placed in the combustion tube, the thermocouple placed in position, the gas train made tight, and enriched air of the desired oxygen content passed until the apparatus was filled with it. The gas flow was then adjusted to a rate of 13.5 liters per hour, which had been selected as a standard for the experiments. The current was turned on in the furnaces for heating the sphalerite and for preheating the air; these were heated rapidly up to 30° to 40° C. below the expected temperature of ignition and then at the rate of 1° C. per min. until the temperature of ignition was reached, as indicated by the potassium iodate-starch solution turning blue. The temperatures of the sphalerite and the preheated air were at all times held approximately the same.

The ignition temperatures as determined by the above method are tabulated in Table 2; Fig. 4 shows curves plotted from the values given in this table.


The results of these experiments show that the ignition temperature of sphalerite is appreciably lowered by increasing the oxygen content of the air supply. This lowering is, however, very, small, the ignition temperature in pure oxygen averaging less than 25° C. below that in ordinary air containing only 21 per cent, oxygen; therefore, the effect of enriched air on ignition temperature would be of very slight practical importance in the roasting of zinc ores.

Rates of Oxidation of Sphalerite in Air Enriched with Various Proportions of Oxygen

Apparatus and Procedure

The apparatus used for this series of experiments was similar to that used for the determination of ignition temperatures, which has been described, except that the bottle of potassium iodate-starch solution was omitted. The sphalerite used was also the same and, as before, separate experiments were run on the following sizes through 20, on 28 mesh; through 28, on 35 mesh; through 35, on 100 mesh; and through 100 mesh.

A 5-gm. sample of sphalerite was used for each experiment. It was spread over the bottom of an alundum boat in a layer about 1/8 in. thick, and was not stirred during the roasting. When starting an experiment, the apparatus was filled with air of the desired oxygen content. The furnace for heating the sample and that for preheating the air were then started and raised rapidly to a temperature of 750° C. The gas flow was adjusted to a rate of 5 liters per hour, and the temperature held constant at 750° C. for 1 hour. The furnace was then allowed to cool rapidly and the partly roasted sphalerite was analyzed for total sulfur and water-soluble sulfur.

The results obtained in the experiments are tabulated in Table 3, and plotted as a series of curves in Fig. 5.


Theoretically, other conditions being equal, the rate of oxidation of zinc blende should vary directly with the partial pressure of oxygen in the air to which it is exposed. The curves in Fig. 5 show that this is borne out fairly well by the experiments in which the oxygen content of the air supplied was less than 50 per cent. With higher concentrations of oxy¬gen, the elimination of sulfur did not increase in the same ratio as the oxygen content of the air. In these experiments with air of high oxygen content, however, the sulfur was reduced to such a low point in 1 hr. that the surface of the blende particles was, no doubt, much reduced and was covered with a coating of zinc oxide sufficient to retard the rate of oxidation decidedly. In the experiments with the -35- + 100-mesh, and the —100-mesh sphalerite, the sulfur elimination was less in pure oxygen than in 50 per cent, oxygen-air; also in 50 per cent, oxygen-air and in pure oxygen the sulfur elimination was less from the —100-mesh than from the -35- + 100-mesh size. This is explained by the fact that the finer sizes, when roasted in air of high oxygen content, tended to sinter and form a cake. No doubt the surface temperature of the sphalerite, because of the rapid oxidation in oxygen, was considerably higher than the temperature of the furnace.

It would probably be safe to state, as a result of these experiments, that under similar conditions, the rate of oxidation of sphalerite varies very nearly directly as the partial pressure of oxygen in the air to which it is exposed, at least for all concentrations of oxygen likely to be used in roasting on a large scale.

A second fact is the large amount of water-soluble sulfur in the calcine from roasting in air of high oxygen content. This indicates that the tendency to form zinc sulfate in the preliminary stages of roasting would be greater with enriched air than with ordinary air. It would be necessary to break these up in the final stage of roasting; this might require a higher temperature or a longer time at a high temperature at the end of the roast than present practice requires.

This series of experiments concerning the effect of enriched air on the rate of oxidation of sphalerite is incomplete. In the experiments just described, the sphalerite was roasted for a definite time in all the experiments, consequently in the experiments with the finer sizes and with enriched air of high oxygen content the sulfur in the sample was reduced to a much lower point than in the experiments with the coarser sizes and with air of lower oxygen content. The rate of oxidation naturally decreased as the sulfur content of the sample decreased, and this effect counterbalanced to a certain extent the effect that the increased oxygen content of the air had of increasing the rate of oxidation. It appears now that a better method of experiment would have been to roast all samples to the same content of sulfur and compare the time required to do this with enriched air of various oxygen contents. It was considered unnecessary, however, to carry this series of experiments any further, as more reliable information is given by the experiments next to be described, which were carried out with a roaster capable of taking a charge of several pounds of ore.

Experiments with a Mechanically Rabbled Laboratory Roasting Furnace

Apparatus and Procedure

The laboratory roasting furnace used in these experiments was an electrically heated, mechanically rabbled furnace, patterned after one used by C. A. Hansen for experiments in the roasting of zinc ores for leaching. It is shown in Fig. 6. A sheet-iron cylinder 30 in. in diameter and 18 in. high was set on timber skids, as a foundation, and a layer of heat-insulating brick laid in the bottom. A heavy sheet of iron was laid level on the layer of brick and on this, concentric with the outer sheet-iron cylinder, was set a thin cast-iron cylinder 16 in. in diameter and 9 in. high. Inside of this inner cylinder was laid a layer of firebrick, covered with about ½ in. of crushed firebrick. On this was set the heating unit, which was a fireclay disk with shallow grooves running transversely across the upper surface in which was wound the heavy chromel wire that served as a resistor. On top of the heating unit a thin fireclay disk was placed. Above this hearth bottom the cast-iron cylinder was lined with a fireclay cylinder ¾ in. thick. On top of the cylinder rested a sheet of asbestos and a heavy iron plate, with a hole in the center for the shaft to which the rabble arms were keyed. All joints in the furnace lining were sealed with alundum cement. An opening 4½ in. wide by 3½ in. high was left in one side of the furnace as a door; it was closed with a firebrick plug.

The rabble arm and rabbles were formed from a single piece of heavy strap iron. The rabbles were so arranged that the ore was thoroughly stirred and, at the same time, maintained at uniform depth over all the hearth. The rabble arm was driven by a small motor and worm gears at a speed of 0.95 r.p.m. The driving mechanism for the rabble arm was supported on a slab of hard asbestos board resting on top of the furnace.

The space between the inner cylinder, forming the roasting furnace proper, and the outer sheet-iron cylinder, was filled with infusorial earth for heat insulation.

The shaft carrying the rabble arm was hollow, and a carefully calibrated platinum-platinum rhodium thermocouple, with silica protecting tube, was inserted through it so that the end rested upon the floor of the roasting hearth. The power input to the furnace was controlled by a voltage regulator; in this way the temperature of the furnace could be regulated to within ±10° C. The temperature of the roasting ore was difficult to determine accurately; occasional readings, taken with a ther-

mocouple thrust into the layer of ore while the rabble arm was stopped, averaged about 20° C. higher than the temperatures read with the thermocouple in the central shaft.

Air for roasting, either atmospheric or enriched, as the case might be, was admitted to the furnace through a ½-in. pipe, curved to direct the incoming air away from the gas outlet and sample tube. The roaster gas left the furnace chiefly through the small cracks around the plug that was inserted in the door of the furnace. Samples of the gas for analysis were drawn off through a silica tube not far from the gas outlet.

Sulfur dioxide in the roaster gases was determined by absorption in potassium-hydroxide solution, and oxygen in the roaster gases and in the air supply by absorption in alkaline potassium-pyrogallate solution, in an Orsat apparatus.

The method of controlling the volume and composition of the air supply is shown in Fig. 7. The atmospheric air was supplied by a small

laboratory blower; a large glass carboy was placed in series with this to act as an accumulator to diminish the effect of minor fluctuations in pressure. Oxygen was supplied from a cylinder of the compressed gas. Flow meters were inserted in both the air and the oxygen supply lines to indicate directly the rates of flow of air and oxygen. By maintaining a constant reading on each of these flow meters, the volume and composition of the air supplied to the furnace could be maintained constant within about 1 per cent. In series with the flow meters were wet gas meters, serving as integrating meters on which could be read the total volume of gas passed in any given interval of time. The oxygen and air supply lines led into a large glass bottle, which served as a mixing chamber; the outlet from this bottle was fitted with a three-way stop cock, of which one outlet led to the roasting furnace and the other to the gas analysis apparatus. The complete equipment is shown in Fig. 8.

The ore used was Joplin concentrate, screened through a 10-mesh screen to give a product of fairly uniform size. Its composition was: zinc, 62.07 per cent.; lead, 1.29 per cent.; iron, 1.41 per cent.; sulfur, 31.58 per cent.; insoluble, 1.84 per cent.; CaCO3, 0.59 per cent. Its

screen analysis is given in Table 4. For each experiment, 7 lb. of this ore was used; this made a layer in the furnace about ¾ in. thick.

When starting an experiment, the furnace was heated to the temperature at which the experiment was to be run the air, of the desired oxygen content, was turned into the furnace; and the charge of ore was placed in the furnace and spread evenly over the hearth. The introduction of the cold ore produced a temporary cooling of the furnace, but within about 15 min. it would again be up to the desired temperature. The temperature of the furnace and the volumes of air and oxygen supplied to the furnace were read every 15 min. The average volume of air supplied (ordinary air + oxygen) in all but one of the experiments was 30.8 cu. ft. per hr. In the one experiment referred to, for which enriched air containing 42 per cent, oxygen was used, one-half the usual volume was supplied, or 15.4 cu. ft. per hr. Samples of the air supply, when enriched air was being used, were taken occasionally and analyzed for oxygen; the variation in the oxygen content was never more than a fraction of a per cent, during the course of an experiment. Samples of the, roaster gas were taken every half hour and analyzed for SO2. Samples of the ore were taken, usually, at intervals of 1¼ or 1½ hr. These were analyzed for total sulfur and water-soluble sulfur; the latter is approximately equivalent to the sulfur present as normal zinc sulfate.

Data Obtained from the Experiments

In Figs. 9, 10, and 11 are plotted the data obtained from a series of roasts made at the constant temperature of 800° C. This series includes one roast with ordinary air, one with enriched air containing 28 per cent, oxygen, one with enriched air containing 42 per cent, oxygen, in all of which the volume of air supplied was 30.8 cu. ft. per hr., and one roast with enriched air containing 42 per cent, oxygen, in which the volume of air supplied was 15.4 cu. ft. per hr. Fig. 9 shows the variation of the SO2 content of the roaster gas as the roasts progressed; Fig. 10 shows the progressive decrease in total sulfur content of the ore; and Fig. 11 the variation in water-soluble sulfur content of the ore. It should be noted that the vertical scale in Fig. 11 is ten times that in Fig. 10.

Theoretically, if the volume of air supplied is the same, the rate of the oxidation reaction, and consequently the SO2 content of the roaster gas, should vary directly as the partial pressure of oxygen in the air supplied for roasting. When air containing 28 per cent, oxygen is supplied, the SO2 content of the roaster gas should be 33 per cent, greater than when ordinary air containing 21 per cent, oxygen is supplied; and with air con¬taining 42 per cent, oxygen, the SO2 content of the roaster gas should be doubled. The time required for roasting should be in inverse ratio to the partial pressure of oxygen in the air supplied. As shown in the first three columns of Table 5, this is borne out approximately by the experimental data.

By halving the volume of air supplied, keeping its composition the same, the SO2 content of the roaster gas can be increased considerably., but the time required for roasting is also increased by about 50 percent., as will be seen by comparing the last two columns of Table 5, and the curves in Figs. 9 and 10.

In the roasting of this ore, made up of fairly evenly sized particles, the SO2 content of the roaster gas was fairly constant until most of the sulfur was eliminated from the ore, especially when air of moderate oxygen content was supplied. This would probably not hold true when roasting an ore made up of a mixture of fine and coarse particles. The total sulfur content of the ore decreased at a uniform rate in all the experiments, until it was reduced to between 1 and 2 per cent., after which it decreased very slowly; this agrees with the usual experience in roasting in practice. In the roast in which half the usual volume of air was supplied, the sulfur content of the ore, when sulfur elimination stopped, was over twice what it was when the larger volume of air was supplied.

The curves in Fig. 11, showing variation of the water-soluble sulfur content of the ore, are interesting. This sulfur remained fairly constant at between 0.1 and 0.2 per cent, in all experiments until the total sulfur

content of the ore became very low. It then increased sharply to a maximum and later decreased again, first sharply and then more slowly, with continued heating. The height of this maximum, and the amount of water-soluble sulfur remaining in the ore at the end of the roast, increased with increasing oxygen content of the air used for roasting. This agrees with the observation made, as a result of the preliminary laboratory experiments concerning the effect of oxygen on the rate of oxidation of sphalerite. Decreasing the volume of air supplied per hour greatly increased this tendency to form zinc sulfate.

Before running the above experiments at 800° C., some similar roasts were made at 750° C., but in this series the mistake was made of charging the ore in the cold furnace and heating the latter up to roasting temperature afterward. Thus, a variable amount of sulfur was eliminated before the furnace reached 750° C. and, while the results were similar to those obtained in the roasts at 800° C., the separate experiments are not strictly comparable with one another. For that reason the analyses of SO2 in the roaster gas are not given, but the curves showing the rate of sulfur elimination from the ore in the final stages of the roast are of such interest that they are given in Fig. 12. The water-soluble sulfur is here plotted on the same scale as the total sulfur, as it runs considerably higher than in the roasts at 800° C.

Noticing first the curves showing the variation of the water-soluble sulfur content of the ore as the roasts progressed, it will be noted that, as in the roasts at 800° C., the water soluble-sulfur remained very low until most of the sulfide sulfur was eliminated from the ore, and then increased sharply to a maximum that was considerably higher than in the roasts at 800° C. Instead of again decreasing rapidly, as at 800° C., it remained stationary at the maximum or at least decreased only very slowly with continued heating. Thie tendency for zinc sulfate to be formed is greater at 750° C. than at 800° C., and the sulfate is not so readily broken up again at the lower temperature. At this temperature, as at 800° C., the formation of zinc sulfate was greater in the roasts with enriched air of high oxygen content.

The curves show that the total sulfur content of the ore decreased at a uniform rate until it was reduced to a few per cent. Then the rate of sulfur elimination became slower, at the same time that the water-soluble sulfur began to increase. Finally, the total sulfur in the ore actually increased and followed along parallel with the water-soluble sulfur. The explanation of this would seem to be about as follows:

The rabbles used in these earlier experiments, though they kept the ore spread evenly over the hearth and thoroughly mixed, for the most part, left a small amount of ore caked in the corner formed between the floor of the muffle and the circular wall. This ore roasted more slowly than the rest and continued to give off SO2 after the rest of the ore was almost completely roasted. This SO2, together with the oxygen of the air, especially in the roasts with enriched air, reacted with the zinc Oxide in the main portion of the ore to produce zinc sulfate, to such an extent that the total sulfur content of this main portion of the ore increased.


It may be concluded, from the data obtained from these roasting experiments, that temperature, volume and composition of air supply, rate of rabbling; and other such conditions being equal, the rate of oxidation of a given zinc ore increases approximately in direct proportion with the oxygen content of the air supply; consequently that the SO2 content of the roaster gas varies approximately directly, and the time required for roasting varies inversely, as the oxygen content of the air supply. If air containing a high percentage of oxygen is supplied, but in reduced volume, roaster gas very high in SO2 can be produced, but in this case the time required for roasting is considerably greater than when air of the same composition is supplied in the usual volume. In other words, the use of enriched air in roasting can be expected to give a proportionate increase in both SO2 content of the roaster gas and rate of roasting, but an extremely high SO2 content in the roaster gas can only be obtained by sacrificing the gain in the rate of roasting, and vice versa.

The tendency to form zinc sulfate is greater with enriched air than with ordinary air.

Results that may be Expected from Application of Oxygen Enriched Air to Zinc Roasting in Practice

It is in the roasting of zinc ores for the manufacture of sulfuric acid that the use of enriched air would be of particular benefit and, at least as far as we can foresee at present, the possibility of the practical application of enriched air to zinc roasting is not great except where the sulfur dioxide in the gas is to be made use of in some way. In this country, the Hegeler kiln is almost universally used for roasting zinc ores when the roaster gas is to be used for making acid; hence it is chiefly the application of enriched air to roasting in Hegeler kilns that will here be considered.

The possible advantages to be derived from the use of oxygen-enriched air in zinc roasting are an increase in the capacity of the roasting furnace, a decrease in the fuel consumption of the roasting furnace, and an increase in the SO2 content of the roaster gas. From the increased SO2 content of the roaster gas would follow increased capacity and more economical operation of the acid plant.

Our experiments in roasting with enriched air in a laboratory roaster show that with equal temperature, rate of rabbling, and volume of air supplied, the SO2 content of the roaster gas and the rate of roasting increase in the same ratio as the oxygen content of the air supply. The heat balances given in the first section of this paper (Cases 2 and 3) show that increases in the rate of roasting of 95 per cent, when enriched air containing 25 per cent, oxygen is supplied, and 58 per cent, when enriched air containing 50 per cent, oxygen is supplied, are necessary to obviate the necessity of using fuel. In obtaining these figures, the SO2 contents of the roaster gases were assumed as 13.3 per cent, and 28.56 per cent., respectively. Our experiments indicate that the SO2 content of the roaster gases cannot be raised this high except by greatly reducing the volume of air supplied; and if this is done, the capacity of the roasting furnace is correspondingly reduced. It would seem then that large roasting capacity and roaster gas with high SO2 content cannot be simultaneously obtained except by the use of enriched air of very high oxygen content and that, therefore, the use of fuel cannot be done away with except by the use of such highly oxygenated air.

It should be borne in mind, however, that rabbling in a Hegeler kiln is done only at very infrequent intervals and that the ore is therefore very inefficiently exposed to the current of air passing over it. If it could be arranged to use enriched air and rabble, let us say, twice as frequently, the rate of roasting and SO2 content of the roaster gas would be much increased and roasting without the use of fuel would be more nearly within the realm of possibility. This question can only be decided by experiments with a roaster having a capacity approaching that of a full-size furnace.

The possibility of applying enriched air to Wedge furnaces, such as those in which the autogenous roasting of zinc ore is now being attempted, should also be mentioned. Roasting can be carried on autogenously in these furnaces as long as everything goes just so, but the margin of heat is so small that any disturbance of conditions in the furnace is apt to upset the balance. The use of air only slightly enriched in oxygen would increase the margin of safety so that no provision would be necessary for burning fuel in these furnaces.

In conclusion, the writers wish to state that, while they believe that the experimental data and the heat balances which they have given are reasonably accurate, they realize that their interpretation of them is not the only possible one and that from the same data, other metallurgists may draw different conclusions as to the effect that the use of enriched air may have on roasting zinc ores in practice. It is hoped that the data given may be of help to others who are working on the application of oxygen-enriched air to the same or similar phases of metallurgy, and serve to stimulate further thought on the subject.


Copper Geology and Mining Methods

The Chitina mining district of Alaska is located at the headwaters of the Chitina and Copper Rivers. At present, the only producing mining properties are the mines of the Kennecott Copper Corpn. and the Mother Lode Coalition Co., which are situated 196 miles from Cordova the port of entry.

The first claims, later acquired by the Kennecott Mines Co. and afterwards transferred to the Kennecott Copper Corpn., were discovered in 1900. The Copper River & Northwestern Ry., which connects the mines with tide water at Cordova, was completed in the spring of 1911.

Contemporary with the construction of the railroad, aerial tram equipment was brought to the mines by pack train and a tramway, 3 miles long, connecting Bonanza mine with the proposed railroad terminal, was finished, enabling shipments of high-grade ore to be made immediately on the completion of the railroad. A mill to treat the lower grade ore was begun the same year.

The Kennecott company’s holdings consist of 111 mineral claims. The Mother Lode Coalition Mines Co., which is controlled by the Kennecott Copper Corpn., owns 73 claims adjoining the Kennecott holdings. All data on operations and geology refer equally well to the Mother Lode property.


The general geology of the district has been covered by the U. S. Geological Survey and the geological features of the mines have been carefully studied by A. M. Bateman, in his capacity as consulting geologist to the company.

The formations in the vicinity of Kennecott are shown, by the U. S. Geological Survey, to be as follows:

Quaternary.—Alluvium: flood plain gravels, sands and silts.
Rock glaciers: broken rock and ice.
Moraines: glacial till, partly sorted.

Jurassic or later.—Quartz diorite porphyry: stocks, sills, and dikes.

Upper Jurassic.—Kennecott formation: shales, sandstones, and conglomerates.

Upper Triassic.—McCarthy shale: shale with few thin-bedded limestones.
Chitistone limestone: massive limestone mostly magnesian, ore containing.

Triassic-—Nikolai greenstone: altered basaltic lava flows.

The Nikolai greenstone is a succession of altered basaltic lava flows, its total thickness, exposed in the vicinity of the mines, is at least 3500 ft. and the base cannot be seen. Numerous prospects have been opened on copper showings in this formation, the ore being usually bornite, chalcopyrite, and occasionally chalcocite; however they have not resulted in productive mines. Native copper is known to occur in all placer operations in gulches cutting the greenstone, some of the nuggets weigh several hundred pounds. In the vicinity of the mines, the strike of the greenstone is N 60° W and its dip 23° to 30° to the northeast.

Chitistone Limestone.—All the important orebodies are in this formation. It is a conspicuous heavy-bedded formation intersected by numerous systems of fracturing; weathering along these fracture planes produced a very rugged topography. It conformably overlies the Nikolai greenstone and is estimated, by Moffitt, to be about 3000 ft. thick.

The lower part of the formation consists of a 4-7-ft. bed of shale; above the shale is 12 ft. of thin bedded, smooth, hard, gray argillaceous limestone, then 23 ft. of thin-bedded, rough, pebbly limestone, containing flattened, cylindrical, fossil-like grains which, from its appearance, Bateman has termed “crinkley lime,” and 30 ft. or more of dull gray limestone. The remainder of the formation consists of massive beds of sparkling light-gray dolomitic limestone, with occasional beds of darker rock. The upper part of the Chitistone limestone becomes thinner bedded and shaly, gradually grading into the overlying McCarthy shales.

Porphyries.—Light-colored quartz diorite porphyries intrude the greenstone and all the sedimentary rocks in the form of stocks, sills, and dikes. They occur most abundantly about one mile from the Bonanza mine, where they form a larger stock, which constitutes Porphyry Mountain.

Faults and Fractures

There are numerous faults both parallel to and traversing the bedding of the sedimentaries. The former are known as flat faults; the latter also pass into and displace the greenstone. There are many displacements of from 1 to 25 ft., and several faults caused a displacement of as much as 1300 ft. Most of these were pre-mineral; however, in the Bonanza and Mother Lode mines there are several instances where a portion of the ore- body has been displaced. Bateman considers that the flat faults have had a direct bearing on the deposition of the ore, the selvage or gouge contained in them acting as a dam to the orebearing solutions.

Ore Deposits

The general geological features and the relative position of the mines are shown in Fig. 1. The orebodies are typical replacement deposits in the limestone, the outstanding features being the intensity of the mineralization and the fact that chalcocite is the predominating mineral in the deposits. As usual, deposition took place along a fissure, or series of fissures that seemingly start from the greenstone, contact.

These fissures have a strike varying from N 30° E to N 80° E and have no definite dip, varying from nearly vertical to 40° from the vertical, most of them more closely approach the vertical, however. The ore-bodies have the same strike and dip as the fissures, although often when a fault plane is intersected, they widen out along these planes and form what are termed the “flat orebodies,” and are identical with the “Manta” orebodies of the Mexicans. The mineralization along the fissures is much less as the fissure passes into the dull gray limestone, and in only two or three instances is any ore found in this formation or the “crinkley lime” beds that immediately overlie the greenstone.

In the Jumbo mine, a fault roughly following the contact between the dolomitic and the dull gray limestone is the west limit of an orebody, the largest mass of high-grade ore so far encountered. This deposit had a cross-section of 80 by 100 ft. and extended from the 150-ft. to the 700-ft. levels, of which a portion 50 ft. wide and 50 ft. high, extending from the 300-ft. to the 600-ft. level, was practically pure chalcocite.

The lower 1000 ft. of the dolomitic limestone appears to be the most favorable zone for ore deposition. All the productive orebodies lie in it and have their greatest width in the lowest beds, gradually becoming , smaller and of lower grade as they extend east into the upper beds. The eastern extension of the fissure is usually filled with calcite. Thus, the orebodies have a rake or pitch practically paralleling the greenstone contact.

There is every degree of intensity of replacement, from large bodies of practically pure chalcocite and its oxidation products, covellite, azurite, and malachite, to the lime containing small bunches or veinlets of these minerals too low grade to mine. There are no defined walls; the grade of the ore is the limiting factor in mining.

In width, the orebodies vary from a few feet to over 100 ft., not including the local widening of the flat orebodies, which sometimes extend another 100 ft.; in length they vary from 150 to over 1000 ft. In some places, practically the entire width is high-grade ore with only a few feet

of lower grade; in others, the high-grade is in veins from 1 to 10 ft. in width, which are separated from one another by lower grade ore. As the eastern ends of the orebodies are reached, with but one exception, no high-grade deposits are found. There are several places where it would appear that pre-existent fissures or veins were filled, but this occurrence is rare.

The Glacier mine exploits a unique and interesting orebody. It is made up of ice, limestone, some greenstone, and chalcocite. The outcrop of the Bonanza mine was a massive deposit of chalcocite located on the edge of a small amphitheater; the debris, resulting from disintegration of this orebody and country rock, fell into this basin and was occluded in a glacier, which now partly fills it. The orebody is 800 ft. long and 85 ft. wide, and the broken ore in payable quantities extends to a depth of 40 ft.; 45 per cent, of the volume is ice, the remainder is broken country rock and chalcocite with a small amount of carbonate ore.

The principal mineral is chalcocite and its oxidation products covellite, malachite, and azurite. Enargite, bornite, and chalcopyrite are occasionally found together with cuprite, luzonite, and other rarer copper-bearing minerals. During the past five years, the ore produced has averaged 70 per cent, sulfides and 30 per cent, carbonates. The ore is divided in two grades: that which is shipped direct to the smelter and the lower grade ores, which are treated in the mill and leaching plant. The high-grade shipments average between 50 and 55 per cent, copper.

Silver exists in the ore in the ratio of about 1 oz. silver to each 130 lb. copper.

General Description

The Jumbo and Bonanza mines are located on the greenstone-limestone contact at an elevation of 6000 ft.; the Erie mine, on the same contact, is at an elevation of 4500 ft.; and the Mother Lode mine is at an elevation of 5200 ft. This last mine was opened in the higher beds of limestone, the vertical shaft intersecting the contact at an elevation of 4400 ft. Contrary to all expectations, the temperature at the elevation of the mine is not extremely cold, rarely falling below —20° F. and during the winter is often 40° warmer than at the mill camp 4000 ft. lower. Freezing or near freezing temperatures prevail even at the lowest levels of the mines, so the mines are dry and dusty; veins of ice are commonly encountered. The only pumping required is during the summer months, when the snow melts and a small part of the water finds its way through open fissures to the upper levels.

The topography is extremely rough and rugged; snow lies on the ground nine months of the year and snow falls throughout the year. Because of the topography, space for bunk houses and other buildings is limited. All hoists, compressors, and other machinery are located underground. Aerial tramways transport the ore to the mill or railroad terminal, all supplies to mines, and, during the winter months, carry all the passengers to and from the mines.

All the mines, except the Erie, are connected underground; a tunnel is now being driven to connect, this mine with Jumbo. Jumbo and Bonanza mines are opened by inclined shafts paralleling the dip of the greenstone and are located about 50 ft. above the contact. These shafts are 14 ft, wide, have two skipways arid a manway, and are 7 ft. high above the rail. The shaft of the Jumbo mine has, a slope distance of 3051 ft. and the shaft of the Bonanza 2416 ft. On account of the flat dip, the manways have stairways in place of ladders.

The skips used at Jumbo have a capacity of 80 cu. ft. and those at Bonanza, 60 cu. ft., with a track gage of 40 in. in both shafts. The Mother Lode mine was opened by, a two-compartment vertical shaft 800 ft. deep. A new incline shaft has been sunk a slope distance of 1405 ft., after the same manner as at the other mines. All are located underground, being connected with the surface by a tunnel. On account of the flat pitch of the orebodies, the vertical shafts would require an excessive amount of development work to open the various levels.

Formerly, levels were driven each hundred feet, this distance was increased to 200 ft., which was found to be too great, and 150 ft. has been accepted as the best distance, all things considered. Two or three pockets are commonly cut at each level and the skips loaded by chutes without a measuring hopper. One pocket for the mill ore is usually capable of holding about 300 tons; the others, for the high-grade and waste, have a capacity of 50 to 100 tons.

Exploration, Sampling, and Estimating

In common with most deposits in the limestone, it is impossible to foretell or estimate accurately the amount or grade of the ore that a block of ground will produce without an unreasonable amount of development work. Diamond drilling has been used to good advantage for exploring unknown ground; in all over 70,000 ft. of drilling has been done. The usual and more reliable method of exploring has been to drive a drift or crosscut in the dolomitic limestone paralleling the strike of the greenstone, and about 100 to 150 ft. from it; thus any mineral-bearing fissure that is encountered can be followed.

Only occasionally is any sampling done underground. After becoming acquainted with the ore, it is possible to estimate closely the grade of the ore by the amount of glance or carbonates it contains. When the limits of the ore are reached, samples are sometimes taken. It has been found that the sample values are usually considerably higher than the actual recovery obtained in the mill; this is probably due to the friability of the glance and the soft chalky nature of some of the carbonates.

Mining Methods

The shrinkage method of stoping has been used, except for the open-pit mining on the Bonanza mine outcrop. A departure from the usual method, however, is practiced. Where the high-grade portion of the orebody is of sufficient size, as much as possible is mined by the shrinkage method and completely drawn out. The mill-grade ore is then stoped, filling the void left by the extraction of the high-grade and the excess is drawn off as usual.

After as much of the high-grade ore is mined as is practical, other veins, lenses, and masses are met and broken with the mill ore. No attempt is made to sort the ore in the stopes after the mining of the mill ore is commenced; but at all the mines, the ore from the skip pocket on the top level passes over a picking belt, where pieces of high-grade ore are hand picked from the mill ore and any mill ore that may be mixed with the high-grade produce is picked out.

The character of the ground makes almost an ideal condition for the method employed. The work must be given close attention to guard against leaving ore that makes along bedding planes, faults or cross fissures, away from the main orebody; although in most instances as the broken ore is drawn from the stope, it is safe to follow it down and, by using a Jackhamer, recover the ore that may have been overlooked. A great many of the floor pillars left are recovered after a level is finished; but it has been found that it is well not to be too hasty about the recovery of pillars and destroying the level, as oreshoots from a lower level have been found in ground that was considered barren. Until recently, no attempt was made to fill these old stopes, as they would stand empty with practically no caving; the waste from development work is now being used for this purpose.

The Glacier mine is worked but three months per year, when surface mining is carried on. During the months of July, August, and September, the ice of the glacier melts sufficiently to release about 30,000 tons of ore; this is recovered by scraping the thawed ground with a Bagley scraper. To date, while some experimental work has been done, thawing by artificial means has not been attempted; possibly operations might be successfully carried on during the cold months, but it would be at a much greater cost. The scraper used has a capacity of 50 cu. ft. and is operated by an electric double-drum engine of 75 horsepower.

Development Plans

As the inclined shafts are located on the western limits of the ore, crosscuts are driven until the orebodies are reached. The drifts on the ore are kept, as far as possible, in the high-grade ore, chute raises are driven 25 to 35 ft. apart, and widened in the usual manner so that they connect, leaving a pillar 25 to 30 ft. thick between the level and the bottom of the stope. Often, if the ore becomes leaner in the drift, work in the stope is carried ahead from the last chute raise, thus determining the direction in which the drift should be driven. In the wide portions of the orebody, a second, and sometimes a third, drift is necessary to draw the ore evenly from the stopes. In other words, the main idea, after the ore is located on a level, is to follow it, as local swells and pinches in the orebody and the method of mining followed preclude any definite layout of the haulageways as in lower grade and more regular orebodies.

In order to mine the ore on the extreme west end of the orebody, it is necessary to drive raises through the underlying dull gray and crinkley

limestone and the greenstone; when the levels are driven 200 ft. apart, a sublevel is driven to eliminate the long raises that would be necessary.

Fig. 2 shows, in plan and projection, a typical orebody and the development work required to stope it. The main haulageways are driven 7 ft. wide by 7 ft. high on a grade of 0.5 per cent, in favor of the loads; the prospecting drifts and crosscuts are 5 by 7 ft.; 16 and 30-lb. rails are used, the gage of track is 18 in. A compressor plant at Bonanza mine furnishes air for all the connected mines, a 6-in. line being used.

Very little timber is used, only an occasional set being necessary in passing through faults or on the greenstone contact; usually native round timber is used with round poles for lagging.

Loading machines are used in driving the larger headings; while they expedite the removal of the broken material, thus avoiding any delay when the miners are ready to set up for the lifters, a crossbar being used, they have not reduced the cost per ton removed. Scrapers are used at the Glacier mine, as noted; they are also employed advantageously when the main inclines are raised out, instead of being sunk.

Tramming is done by hand, horse, and storage-battery locomotives. Hand tramming is used where the distance is short and a small tonnage is moved; horse tramming, when the distance is greater; for the long hauls and on the levels producing the greatest tonnage, 4-ton Baldwin-Westing-house locomotives with Edison cells are used. This type of locomotive has given very satisfactory service.

For horse and hand tramming, 20-cu. ft. end-dump cars are used; with locomotives, cradle-type and side-dump cars of 36 cu. ft. capacity are used, usually in trains of six or eight cars. .While, on several levels, the locomotives run on 16-lb. rails, the practice is to use 30-lb. rails; curves have a minimum radius of 40 ft.

Hoisting is done in balance, the hoists at the Jumbo and the Bonanza are duplicates; they are of single-reduction, herringbone-gear type with a rope speed of 600 ft. per min., driven by two 85-hp., a.c., 2200-volt, three-phase, sixty-cycle motors; they were manufactured by the Allis Chalmers Co. The Mother Lode incline will be equipped with a double-drum hoist, with double reduction gears, driven by two 75-hp. motors; the rope speed will be 450 ft. per min. The cables are six-strand, nine- teen-wire, Lang lay, 7/8 in., in diameter. When hoisting men, the skips are removed and a man car used. Neither skip nor man car is fitted with a safety device, as a satisfactory one has not yet come to the company’s attention.

The air-compressor plant furnishes air for all mines, except the Erie, where an Ingersoll-Rand Imperial type 10, 600-cu. ft. capacity, electrically driven compressor is installed. The plant contains: One Ingersoll Rand type P. E.-2 compressor, 1500 cu. ft. capacity, driven by a 250-hp. synchronous motor; two Ingersoll Rand Imperial type 10 compressor, 500 cu. ft. capacity, each driven by a 85-hp. motor; one Ingersoll Rand Imperial type 10 compressor, 650 cu. ft. capacity, driven by a 105-hp. motor.

Because of the numerous openings to the surface, natural ventilation, with the exception of small fans belt-driven by a 10-hp. motor in development, aided by doors to course the air, is satisfactory.

Electric lights are used on the levels and incline shafts. The miners use carbide lamps, furnishing their own caps and lamps, the company keeping them in repair.

Each level has a telephone connecting with the foreman’s office, compressor room, and hoist room. The mine telephone system is independent of the general system.

Electric pull bells, modeled after those commonly used in other mines, are used.

Operating Data

Types of Drills

For drifting Ingersoll Rand, 248 Leyner machines are used; for stoping and raising, Ingersoll Rand C. C. 11, except when drilling in chalcocite, when it is necessary to use a water-type drill. Ingersoll Rand B. C. R. 430 and Sullivan D. P. 33 are used for blockholing and where occasional flat or down holes are to be drilled. Four-point, cross, high-center drill bits are used on all machines, made up of the following sizes of steel: 1-in. quarter octagon for stoper; 7/8-in. hollow hexagon for Jackhamer; 1¼-in. hollow round for Leyner. The bits are:

Stoper, 1 7/8-in. for starters; 1 ¾-in. for seconds; 1 5/8-in. for thirds; and 1½-in. for fourths.

Leyner, 2-in. for starters; 1 7/8-in. for seconds; 1¾-in. for thirds; 1 5/8- in. for fourths.

Record of Unit Production

(a) Ore broken……………………………………………………297,502 short tons
Ore produced…………………………………………………….294,202 short tons

Labor Data

(c) Stoping labor includes: Miners in stopes, muckers in stopes, bulldozers in stopes, rockbreakers in stopes:
Tons broken per man per hour…………………………….1.3964
Man-hours per ton……………………………………………………0.7161
(d) and (e) Exploration and development labor, miners only:
Tons broken per man per hour……………………………1.2941
Man-hours per ton…………………………………………………0.7726
(g) All underground labor including above labor:
Tons produced per man per hour……………………..0.4586
Man-hours per ton…………………………………………………2.1807
(h) Surface labor, exclusive of office force:
Tons produced per man per hour……………………11.5536
Man-hours per ton……………………………………………….0.0866
(i) All labor including office force:
Tons produced per man per hour…………………….0.4243
Man-hours per ton……………………………………………….2.3570

Supplies Data

Safety Measures

Hoisting ropes are thoroughly inspected once each week; every six weeks 2 ft. are cut off from both ends. The ropes are changed end for end after six months’ use. The sheaves are inspected once every week, and the hoists each day. In the vertical shaft, the safety catches are tested every Sunday.

There is a fire extinguisher on every level station; fire doors are provided. When located near timber or snow sheds they are of concrete and steel; otherwise they are built of wood, care being taken to make them as air-tight as possible.

No safety engineer is employed, the engineering department reporting to the general mine foreman and superintendent any unsafe practices that come to its notice.

Each bunk house is equipped with a pool and reading room, a number of magazines and other periodicals being provided. Moving picture shows are given twice a week at Jumbo and Bonanza camps.

A well-equipped hospital is located at the mill camp with a competent surgeon and corps of nurses in attendance.

At the plant, last year, there was one fatal accident; no serious accidents causing total permanent disability; three partial permanent disability; 28 causing loss of more than 14 days time; and 120 minor, loss from 0 to 14 days.

Compensation paid, under Territorial Act, amounted to 1.058 per cent, of the payroll.

Mining Methods and Costs

The Globe Mining District is in the southeast central part of Arizona, in Gila County. Globe, with a population of about 7000, is the terminus of the Arizona Eastern R.R., a branch line 130 miles long that connects with the Southern Pacific R.R. at Bowie.

In 1874, prospectors crossing the Pinal Mountains from the west located what is generally known as the Old Dominion mine. For several years, it attracted little attention, because of the greater interest aroused by the discovery of high-grade silver ores in some of the foothills northeast of Globe. About six years later, the prospector turned his attention to the abundant copper ore revealed by surface workings along the Old Dominion vein, and, in 1884, the Old Dominion company erected two 30-ton furnaces. From 1888 to 1893, the Old Dominion company is said to have maintained an average annual production of about 8,000,000 lb. of copper. Until Dec. 1, 1898, all supplies had to be freighted into Globe by wagons and the mines of the district operated intermittently because of high expenses, but with the advent of the railroad the Old Dominion company continued to be a large and steady producer.

In the Globe district, the production of copper far exceeds in importance that of any other metal. There are four operating companies along the Old Dominion vein and the total annual product of these properties for 1923 was about 45,000,000 lb. of copper.

The unit size of the mineral tracts in the district is the regulation mining claim, 600 ft. wide by 1500 ft. long, and all ownerships are held in fee.

Globe is 3600 ft. above sea level, and lies between the Apache Mountains to the east and the Pinal Mountains to the west. The principal drainage of the district is northward through Pinal creek into the Salt River. The general slope from the high point along the vein, where the Superior & Boston mine is, to Pinal creek, where the Old Dominion mines are, is about 250 ft. to the mile.


The oldest rocks in the district are pre-Cambrian crystalline schists known as the Pinal schist, which are the basement upon which all the later rocks were deposited. These latter rocks comprise shale, conglomerates and quartzites with a total thickness varying from 500 to 800 ft. and are thought to be Cambrian in age. Overlying these rocks is a series of limestones, known as Globe limestones, that vary in thickness from 300 to 500 ft. and range in age from Devonian to Pennsylvanian.

These rocks have been cut by numerous faults, and following or accompanying the faulting large sills and masses of diabase were intruded between the sedimentary beds. A long period of erosion followed, during which the region was deformed by further faulting to its present topography and during which the original ores were deposited.

The main fault in this district and the one along which most of the mining is carried on is known as the Old Dominion fault; it varies from 3 to 50 ft. in width and is developed for a length of approximately 3 miles. The fissure has a variety of strike and dip but is roughly north-east and southwest, with a dip of about 80° to the south.

The fault is fairly conspicuous and is easily followed, except where it is wholly in diabase, when its course is marked by a zone of brecciation stained with hematite and salts of copper.

The vein is commonly made up of brecciated shale or quartzite and mineralized with oxide of iron and the ores of copper, the overburden varying from 200 to 600 ft. The mineralogical character of the ores along the vein is simple. The oxidation of the sulfides has resulted in simple products. The pyrite and chalcopyrite have their sulfur replaced by oxygen, carbon dioxide, or silica and become hematite, limonite, cuprite, malachite, or chrysocolla. The secondary sulfides recognized in the district are chalcocite and bornite. Native gold, silver, and copper have been observed in small amounts within the zone of oxidation.


Most of the exploration along the Old Dominion vein has been done by test pitting, tunneling, trenching, shaft-sinking, drifting, and cross-cutting. All development is carefully sampled, the outline and extent of ore body carefully determined, and ore estimated on a basis of 11 cu. ft. per ton. The production, as indicated by exploitation, has proved the method sufficiently accurate.

Change in Mning Method

The principal mining method formerly in use was the square-set method but the decreasing copper content and the increasing cost of timber, together with the increasing cost of labor and supplies, made it imperative that a cheaper method be substituted. In some places along the vein where the ground is heavy and where it is imperative to keep timber close to the working face, square-setting is used, but in general that method has been superseded by newer and more economical methods; the selection of method depends on the size and shape of the orebody and the character of vein filling and walls.

Sampling and Estimating Reserves

The method of sampling is far from elaborate; grab samples only are taken from each round blasted daily. A record is kept of all samples taken in the block of ore lying between any two raises and the numerical average for the month is assumed to represent the value of the ore mined that month.

When estimating the reserve, which is done the first of every year, the area of each block of ground remaining between any two raises is measured on the profile tracings with a planimeter. This area multiplied by its average width gives the contents in cubic feet. This figure divided by 11 (11 cu. ft. ore in place is equivalent to 1 ton) is reported as the tonnage for that block. The numerical average of the year’s samples of ore broken, together with the assay values of the drift over this section, is reported as the assay value of the block remaining.

The total ore reserve is computed by multiplying the number of tons by the per cent, for each block. This product divided by the total reserve tonnage gives the average per cent., which practically checks with the heads reported by the mill. Blocks of ground that have been worked out have been found to check within 0.5 per cent, on both tonnage and value reported.


The section of the Old Dominion vein along which the Iron Cap Copper Co. is mining is about 3500 ft. long, and the width of the vein varies from 3 to 40 ft. with an average dip of about 80°. The vein material is a hard brecciated quartzite or shale between fairly good walls. The distribution of values through the deposit is irregular and some sorting is resorted to. Very few waste bodies occur in the ore zone, however, and when found are usually left in place until stopes are ready for waste filling; they are then blasted down and become part of the gob.

The inclined cut-and-fill system is used throughout the mine. This method requires very little timber; stope floors are carried on an incline of about 34°, which eliminates most of the labor of shoveling but which is not so steep as to constitute a hazard from rolling boulders.

The average stope temperature is about 78°, average relative humidity 88 per cent. Production is approximately 150 tons per 8-hr. shift of 50 men. This includes all underground labor, but only about 50 per cent, of the shift are on actual stoping operations.

Fig. 1 shows the successive steps from the starting of a stope to its finish. This plan calls for a main shaft for the handling of all men and materials and the opening up of the mine by a series of levels placed approximately 100 ft. apart.

The Iron Cap Copper Co. has two three-compartment shafts, each compartment 4½ by 5 ft. in the clear, timbered with 10 X 10-in. timber sets on 5-ft. centers and lagged with 2 by 12-in. lagging. The Iron Cap shaft, the only one at present operating is 1540 ft. deep, and is in the hanging wall. Commencing at the 800-ft. level, stations 15 ft. high by 40 to 60 ft. long are cut every 100. ft. Station sets are 10 by 10-in.

timber on 5-ft. centers with a drop of 6-in. on each set, leaving the back end of the station 11 ft. high.

Loading pockets were cut above the 1100-ft. level and under the 1300-ft. level; raises driven from each pocket accommodate the ore mined on three levels.

The shaft is so conveniently located with respect to the vein that all ground between shaft and vein constitutes the length of the station. Drifts 5 by 7 ft. are then driven along the footwall, no timber being used until stopes are started. Fig. 2 shows the arrangement and dimensions of shaft, station crosscut, and drift.

Mining Methods

As soon as the drifts have advanced far enough, raises with a minimum cross-section of 5 by 10 ft. are driven on 125-ft. centers. The sill floor sets of these raises are timbered with 10 by 10-in. timber. Posts are 8½ ft. high, sets placed on 5-ft. centers. Above the sill floor set, however, the timber is 8 by 8 in. Posts are 5 ft. 4 in. long.

The manway only is timbered with what is termed a “clap-me-down” set. The posts are set as nearly over each other as possible,

the cap is cut to fit the ground and is well blocked; 3-in. by 12-in. by 6-ft. lining boards keep the manway clean. A 6 by 6-in. sprag in the chute end flush with the last cap serves as staging for the machine men.

All headings are given definite numbers, which indicate to a certain extent their location with respect to the shaft. Headings to the east of the shaft have an even number and headings to the west, an odd number. Raises are numbered consecutively from the shaft in each direction. For instance, 902 raise No. 4 would be the fourth raise east of the shaft on the 900-ft. level. Stopes are designated as 902 stope east or west of raise No. 4.

Stopes may be started as soon as the raise has been holed through to the level above. The back of the original drift is first broken down to a height of 15 ft. and the ore mined from footwall to hanging wall. This sill-floor stope is then timbered with 10 by 10-in. sets with square framing. Sets are placed on 5-ft. centers; posts are 8½ ft. long. If the vein is not over 8 ft. wide, the cap is cut to fit the ground. In some cases where the vein is 6 ft. wide or less, and walls are exceptionally hard, hitches are cut to receive the cap and no posts are used.

A temporary chute is placed 25 ft. on each side of the original raise; permanent chutes are placed 50 ft. on each side of the original raise with a manway between.

Filling Stopes

As soon as all sill-floor timber has been placed, the sets are lagged over with a double floor of 3 by 12-in. planks and stoping starts on the first floor at the original raise. The ground is blasted out around the raise as high as safety will permit, the ore is then removed and filling poured in from the level above, the waste taking its own angle of repose. When the waste filling is about 3 ft. from the back of the stope it is roughly leveled off and floored with 3-in. by 12-in. by 5-ft. planks. Cleats of scrap timber are nailed to the floor to enable machinemen to move around easily and a cut about 6 ft. high is taken each side of the raise. When this cut is completed, the floor is taken up and piled out of the way and the opening is filled as before; the flooring is again laid and another cut taken. The flooring is used until it is worn out.

As the stope passes the temporary chute, the timber from this chute is salvaged for use elsewhere. As the toe of the incline reaches the permanent chute, the original raise timbers are salvaged as the stope progresses and are used to build up the permanent chutes and manway.

In all main drifts 2-in. air lines and ¾-in. water lines are carried, 1-in. air lines and ½-in. water lines are run down each original raise and up each center manway, so that drilling connections may be made at either the top or the bottom of the incline.

As stopes hole through to the level above, drift timbers are caught up and held in place until they can be supported by 8 by 8-in. sets placed upon the filling; these are eventually filled in and the original drift left intact.

Waste fill for stopes is obtained from development work, from shrinkage stopes above the ore zone, and from old filled stopes where no damage can be done by allowing them to cave.

Drilling and Blasting

All stoping is done with wet hand-rotated stopers using 7/8-in. quarter-octagon hollow steel. The starter bit is 1¾ in. and decreases 1/8 in. on each length of steel. Thirty-five per cent, gelatin powder is used in blasting all holes in stopes.

A grab sample is taken from all rounds blasted and a copy of the assays is furnished to bosses daily. All broken material is hand trammed in 16-cu. ft. end-dump, roller-bearing cars run on 12-Ib. rails.

The method adopted for drilling drifts, raises, and shaft sinking do not call for any special mention. This work is usually done on contract, the company providing all tools, equipment, and supplies; and the contractor providing all labor. The average price paid for drifting is $4.40 per linear foot with a minimum cross-section of 5 by 7 ft. The price for raises with a minimum cross-section of 5 by 10 ft. is $4.50 for the first 50 ft. and $5 per foot for the remainder. Shaft sinking averages about $50 per foot, depending largely on the character of rock being drilled and the amount of water likely to be encountered. All drifting and shaft sinking are done with water Leyners using 1¼-in. hollow round steel with double-taper cross bits. The starter bit is 2 in. and decreases 1/8-in. on each length of steel.

Blasting in all development work is done with 40 per cent, 1 1/8-in. gelatin powder. All development is carefully sampled and accurate assay maps brought up to date every 30 days. About 1 ft. of development is done for every 12 tons of ore mined.


Tables 1 and 2 show average detailed costs of stoping and development work per ton of ore for 85,211 tons mined in 1923. These costs constitute approximately 50 per cent, of the total cost of mining. At

present 30 miners break all ground in development work and stopes, including waste to fill stopes, and supply 300 tons of ore daily. There is an average of 100 men employed underground and a total of 135 at the mine. All labor in stopes is on a “day’s pay” basis.

Machinery and Surface Plant

The surface equipment consists of an Allis Chalmers double-drum hoist driven by a 250-hp., 440-volt, 25-cycle electric motor, each drum holding 1800 ft, of 1 1/8-in. 6 X 19 Lang lay cable. All hoisting is done in counterbalance at a speed of 700 ft. per min. Ore is hoisted from the pocket through two compartments in 4-ton skips and dumped direct into a choke feed Austin No. 7½ gyratory crusher, which breaks to about 2 in. This material then passes through a trommel and the oversize is fed to a Symons 48-in. vertical disk crusher, the final product being ½ in. A belt conveyor carries it to the ore bins, and from there it is taken to the mill, ½ mile away, by a Westinghouse 6-ton locomotive operating over a 24-in. gage electric railroad on 550-volt d.c. current.

A cage for hoisting men is hung under the skip, and provision is made for connecting a second cage underneath if necessary. An unbalanced dinkey cage is operated in the third compartment by a 10 by 18-in. duplex, direct-acting sing!e-reel Ottumwa hoist; this cage is used only to lower supplies. Skips and cages are both equipped with safety catches and are inspected daily.


At present, the water is handled by electrically driven pumps in two lifts, from the 1500 to the 1300-ft. level, and from this level to the mill on the surface. Both pumps are on the 1300-ft. level. A Lane & Bowler 500-gal. deep-well pump, six-stage, driven by a 40-hp. motor lifts the water from the 1500-ft. level through an 8-in. pipe and discharges into two 60,000-gal. concrete sumps on the 1300-ft. level. The water gravitates into a 300-gal. Aldrich quintuplex plunger pump driven by a 150-hp. motor and pumps through a 6-in. pipe direct to the mill on the surface, about ½ mile from the collar of the shaft.

Air Compression

Air for drilling is furnished by a steam-driven 3000-cu. ft. O. R. C. Ingersoll-Rand compressor. A Sullivan 1500-cu. ft. tandem, compound, direct-connected, steam-driven compressor is idle at present but can be used in an emergency.


Ventilation is provided by one Sturtevant multivane fan pulling 65,000 cu. ft. at 3½-in. water-gage pressure, driven by a 75-hp. motor, belt-connected. This exhaust fan is at the collar of the Williams shaft and is operated 14 hr. per day. Air is drawn down through the Iron Cap shaft to the lowest mine level and allowed to work upward through the stopes, finally finding its way out through the Williams workings and up the Williams shaft.

Lighting and Signaling

Stations are lighted by 32-c.p., 100-watt, 110-volt electric bulbs; main tramming drifts are lighted by 16-c.p. 40-watt, 110-volt electric bulbs. Lights in stopes are provided by carbide lamps carried by each miner.

Western Electric, local-battery system telephones are located on each shaft station and at the collar of the shaft. All signals between cagers and hoisting engineer are over an electric signal system, supplemented by rope bell.

All electric power used is bought from the Inspiration Copper Co. at Miami, Ariz., and brought over a private line a distance of 7 mi. All steam power is furnished by three 250-hp. Babcock & Wilcox water- tube boilers.


A Member.—What do they put on the filling?

A. L. Walker.—They put lagging and use that lagging until it is worn out, gradually moving it from one place to another.

A. Neustaedter, Roselle Park, N. J.—Do they put up square-set raises?

A. L. Walker.—They put up square-set raises, but use square sets only when it is imperative. If the ground is at all soft or dangerous they use stulling.

A. Neustaedter.—Cribbing would not do.

A. L. Walker.—It might in certain cases. Of course, in the old days, the square-set system was used altogether. In the Old Dominion mine, where the orebodies were 40 or 60 ft. wide and sometimes 60 ft. high, square sets were used altogether.

A. Neustaedter.—Do they work the square-set stopes on the rail too?

A. L. Walker.—Yes; all the orebodies above the eighth level in the Old Dominion were worked with the square-set system. About 1893, however, that system became so expensive that we developed a system of heavy stulling to support the roof whenever possible.

Mining Methods of Michigan

The Marquette range, on which are situated the iron mines of Marquette County, together with a few in Baraga County, Mich., extends from a point 10 miles southwest of Marquette westward for 30 miles. The tracts are usually a multiple of the standard 40-acre parcel, which is the smallest government subdivision of a square mile or section.

About half of the mines are held in fee, these being owned by the older mining companies. Some of them date back to about 1880, and a few are as early as 1850. During the past 40 years most of the mines opened have been on leased lands, the royalty being either for a stated amount per ton or a percentage of the selling price of the ore at the lower lake ports, or the selling price on board of cars at the mine.

The first merchantable body of ore discovered in the Lake Superior district was found, in 1845, at what is known as the Jackson mine, near Negaunee. This ore was hard hematite. After unprofitable attempts had been made, in 1848, to smelt it in forges, shipments were begun to lower lake ports, which increased rapidly upon the completion of the locks at Sault Ste. Marie, in 1855, and the railroad from the mines to Marquette, in 1857.

Beginning with the Pioneer, in 1858, a number of small charcoal furnaces were built to smelt a part of the product. At various points in the upper and lower peninsulas, large charcoal furnaces are still making iron from the ores of the Marquette range. In connection with these furnaces are byproduct plants.

As the deposits were opened up, soft ore was encountered, but for a few years this was disregarded, as only the hard ore was used. Underground mining was begun about 1880 and during the next few years the open-pit mines producing high-grade ore were exhausted.

The first mines in the district were owned by the Jackson, Cleveland, Lake Superior, Lake Angeline, Champion, Iron Cliffs, Humboldt, Republic, and Michigamme companies. Many of these companies have been merged in the holdings of The Cleveland-Cliffs Iron Co., which controls most of the tonnage of the district.

Ore shipments are made from Marquette through the docks of the Duluth, South Shore & Atlantic railway, constructed in 1857, and from the Lake Superior & Ishpeming railway, constructed in 1896. A portion of the tonnage is also shipped over Chicago & North Western to Escanaba. Since the above dates, larger and more modern docks have been built.

The average number of men employed in the district for 25 years is 4215. The average annual shipments from 1902 to 1920, inclusive, have been 3,827,659 tons; the total shipment to the end of 1921 is 137,237,513 tons.

Geology of District

The geology of the Marquette range is described, in Monograph 52 of the U. S. Geological Survey, by Van Hise and Leith. The iron formations occur in the Huronian series of the Algonquin group of pre- Cambrian rocks. The sedimentaries, in which occur the principal mines, stretch from Marquette through Negaunee, Ishpeming, and Champion to Michigamme and Republic, with a separate area at Gwinn. The series consists mostly of quartzites and slates, interbedded with them being the jaspers of the iron formations. All of these rocks are faulted and folded and are crossed by dikes of greenstone or diorite. The Negaunee iron formation, or jasper, in which most of the mines occur, is a combination of iron oxide and silica containing, according to the U. S. Geological Survey, about 29 per cent. iron. No commercial use has been made of this material. Its greatest thickness, where proved by drilling at Negaunee, is over 2000 ft. It rests upon slates belonging to the formation, which is locally known as the Siamo, and is overlaid by quartzites or slates of the Goodrich formation. The soft ores occur as secondary concentrates either near the base or near the top of the jasper, the latter resting upon interbedded diorite intrusions. These orebodies are often limited in depth by dikes or faulted portions of diorite or slate, which serve as impervious bases on which concentration has taken place.

The hard ores occur only at the top of the Negaunee formation, being underlaid by a few hundred feet of hard-ore jasper, which again lies on the jasper of the soft-ore formation. The hanging wall of the hard ores is quartzite or slate.

The Michigamme slate formation, which overlies the upper quartzite and the Negaunee formation, contains interbedded iron formations, which in places produce limonite ores.

The Gwinn, or Swanzy, subdistrict of Marquette range is placed, by the U. S. Geological Survey, in the Michigamme formation. Here the ore is a soft hematite, found at the base of a jasper 100 ft. or more in thickness, and resting on comparatively thin beds of black slate and quartzite or arkos overlying the granite. The iron formation is overlaid by the slates of the Michigamme formation.

The physical structure of the ores on the Marquette range is excellent, none of them having a fine enough structure to be objectionable to furnace men. The hard ores are used in lump form in the open-hearth processes; they are crushed for use in the ordinary blast furnace.

Description and Topography

The district is about 800 ft. above Lake Superior, or 1400 ft. above the sea level. The surface is hilly and rocky. Lakes and swamps are bordered with terraces of glacial origin, above them rising rocky hills of iron formation, quartzite or diorite, the tops of which are from 50 to 200 ft. above the glacial terraces. Because of the proximity of Lake Superior the summer climate is cool. The prevailing northwest winds bring heavy snow falls from the beginning of November until the middle of April. The winter temperature is modified by the lake, which never freezes over entirely. The average yearly rain fall, which includes the equivalent in snow, is about 32 inches.

Mining timber is brought in by rail. Practically all the white pine of the Upper Peninsula has been removed, and in the neighborhood of the mines, most of the hardwood also has been cut. There still remain districts where there are large stands of hardwood, together with tamarack, hemlock, and cedar.

Labor conditions have been excellent. The mining population, derived chiefly from northwestern Europe, has been industrious and thrifty. Most of the men own their homes, the lands upon which they stand being either purchased or leased from the companies on easy terms, which induce building. However, since the war, many workmen have been attracted by high wages to the large cities, this movement being accelerated when the mines were shut down during the recent depression.

The ore is transported by rail from the mines to the docks in hopper- bottomed cars of 50-ton capacity. It is there dumped into pockets of 200-ton capacity. In loading vessels, the ore is delivered through spouts, which are lowered to the hatches. These spouts are 12 ft. from center to center, and the hatches on the boats are 12 ft., 24 ft., or a multiple of 12 ft. apart.

Pumping at the mines varies from a few hundred to about 3000 gal. per min. The overlying sand and gravel are often saturated with water, but owing to embedded clays and other courses, this is seldom drained until broken by the extraction of the ore, under the caving system.


The earliest explorations were for the purpose of finding the ledge or deposit from which had come the many broken masses of hard ore found lying upon the surface or in the glacial material. Succeeding explorations were conducted by test pitting through the overburden, drilling the same way from outcrops, or tunneling into the rocks themselves. In many cases, shallow shafts were sunk, from which drifts were driven.

The diamond drill was used, as early as 1869, for deep holes in hard orebodies, but its use was not customary until about 1878; since then it has become the usual method of exploring for both surface and underground work. The churn drill has not been used much for exploring, owing to the hardness of the rock capping to be penetrated.


As all explorations of orebodies at present are by diamond drilling, the only sampling is that of the core and sludge, from the drill holes. These are collected after each 5-ft, run and later analyzed; the weighted average of the two, figured on the proportion of each covered by the length of the run, is the analysis for that run. These analyses, combined for the entire hole, give the depth and grade of known ore encountered. Duplicate samples of core and sludge from each run are preserved for reference. If the presence of soluble sulfur is suspected, the amount of water pumped down the hole and the amount coming out are measured, and samples of this water taken at definite intervals, these samples being analyzed for sulfur. This is then combined with the analysis of insoluble sulfur in the core and sludge. When drilling through hard ore, almost complete recovery of core can be made; while in soft ore practically no core is obtained, and the only analysis is that of the sludge. The sludge is collected by causing the water from the drill holes to flow into boxes 4½ by 1½ ft., with two baffle boards, in which the particles of ore held in suspension settle out.


The methods of estimating are those customarily used in the Lake Superior district, both for ore found by diamond drilling and for developed ore underground. This is a comprehensive subject and should be treated in a separate paper. Plans and cross-sections are made both of explorations and mine workings to show the area and depth of the deposits as soon as ascertained. The limits of formation are shown on these drawings as a result of careful geologic examination. Considerable latitude for judgment must be permitted in the case of partly developed orebodies, which for the purpose of estimating the cost of development must be divided into ore in sight and prospective ore. The number of cubic feet per ton varies from 8 to 9 in hard ores and is about 12 in soft hematites, while for limonites it is as high as 13 or 14. The U. S. Geological Survey states that the soft ores of the Marquette range average 12 cu. ft. to the ton. This is borne out by a series of careful tests, which have recently been made in several mines.

Accuracy of Method

Due to the irregular shape of the various orebodies, large discrepancies have often been found between careful estimates based on exploration and the tonnage eventually recovered. Drill holes in some cases follow chimneys, the formation proving barren except for a small cross-section in the vicinity of the hole. On the other hand, some deposits have proved to be much greater than estimated by drill holes. Reasonably correct estimates can be made in shallow deposits from a number of short holes at close intervals, but the difficulty increases for depths over a few hundred feet, due to the deviation of the diamond-drill holes and the lack of knowledge of the geology of the formation.

In the case of shallow regular orebodies, as on the Mesaba range, the tonnage can be accurately estimated and the percentage of extraction determined. This is not the case on the Marquette range, where the ore deposits, as a rule, are deep and extremely irregular.

History of Principal Mining Methods

The early shipments of iron ore previous to 1860 were made largely from loose masses found scattered on the surface. After this source was exhausted, open pits were developed, some of which continued to operate until about 1880. Drilling by hand, blasting with black powder, loading into carts drawn by horses or mules, and again loading into 7-ton cars on the railroad constituted the usual method. Shafts were started for collecting the water in the pits, so that it could be pumped. As the pits grew deeper and available ore seams were followed, winzes or stopes were sunk, the ore being raised by horse whims. Most of the early appliances were introduced from Cornwall, whence the first miners came. Details of the mining methods, including cost, are given in Volume 1, Geological Survey of Michigan, published in 1873.

As the open pits were continued, they increased in depth until a point was reached, on account of the ore dipping under the rocks, where it was not profitable to remove the overburden. It was then necessary to sink inclines and provide mechanical means of hoisting. The hard ore was then mined in open stopes, pillars being left to support the capping.

From 1875 to 1880, much of the mechanical equipment was introduced, including rock drills, electric lights, and electric signals. At this time dynamite replaced black powder as the chief explosive. Hard-ore mining continued in about the same manner at depth. The breast stoping method of room and pillar was used, in which only enough ore was left in pillars to support the hanging.

The first soft ore was found near hard ore deposits. The demand for this class of ore was limited, previous to 1880. It was mined in open pits, but when it became unprofitable (on account of the increased depth and the dipping of the ore under the rock, to remove the overburden) underground mining was started. This soft ore could not be mined by the open stope and pillar method; timber was necessary to keep the places open. The principal method of mining soft ore was, in the middle eighties, almost entirely by the square-set system of rooms and pillars. The rooms were usually three sets, or 21 ft., wide and as long as the orebody. As a rule, the pillars were of the same width as the rooms. In many instances the rooms were carried to a height of ten or twelve sets, or 70 or 84 ft. After the ore had been mined in the rooms, an attempt was made to get what was left in the pillars by raising and running them. This system was extremely wasteful, as the ore soon became mixed with rock and the grade lowered so that work had to be stopped.

The caving system was introduced by miners from the north of England, where it originated. The method there was to mine from a sub-level immediately below a mat of timber, which was kept propped up until the retreat of mining began. Every effort was made to maintain this timber mat, for when it was destroyed a new mat had to be made at great cost. Modifications of this caving system were introduced during the early eighties, until it became the accepted method about 1887 to 1890 (see J. P. Channing in the Lake Superior Mining Institute, Volume 19); the introduction of the caving system was hastened by the decreasing local supply of large timber.

Permanent shafts in the foot wall were rare before 1895. Timber was used exclusively for shafts, shaft houses, and trestles. About 1900, steel replaced wood in shaft houses and, about 1910, concrete and steel were used for shaft lining. The most recent practice, inaugurated in 1919, is to build enclosed shaft houses of reinforced concrete. Electric haulage was introduced in 1892, since which time its use has become general.

Certain changes in mining conditions were brought about by legal restrictions, though in justice to the mining companies it should be said that much of the beneficial legislation enacted was prompted by the managers. The election of the mine inspector in each county, provided for in 1887, assured a greater degree of care for the safety of the workmen, which was increased after the passage of Michigan’s workmen’s compensation law in 1912. All large companies now have a department for safety inspection and first-aid training, although these are not required by law.

Early operations were wasteful, because of the lack of system and the necessity of marketing only the higher grades of ore. As the number of leased properties increased, it became necessary that all ores should be taken out in a workmanlike manner and unnecessary waste prevented. The principal causes of loss in the early years of the district were: (1) the lack of preliminary exploration and the beginning of caving before the limits of the orebodies had been determined; (2) the fact that softer material could be mined more cheaply than hard, which consequently was left in place; (3) in attempting to reduce the cost too great a vertical distance was taken between sublevels; (4) the lack of proper maps and systematic method of laying out the work.

The larger and more progressive mining companies, realizing these mistakes, inaugurated geological investigation, systematic development and close supervision, which resulted in the present methods by which losses have been reduced to a minimum.


The demand for crushed soft ores for charcoal furnaces necessitated the introduction of crushers in shaft houses.

Attempts have been made to concentrate some of the lean ores but, principally because of the intimate association of silica with the iron oxide, these have failed. The first concentrating plant in the district was built at the Jackson mine in about 1880. It failed because jaw crushers and rolls could not be made hard enough to withstand the wear and tear (manganese and other alloys of steel were not then in general use) and because, regardless of how fine the crushing might be, the particles of iron oxide and silica were too closely associated to be separated. Magnetic concentration was attempted, about 1890, by Thomas A. Edison at Humboldt and Michigamme by means of machinery introduced from Sweden for the magnetic treatment of magnetites. These attempts failed, owing to the small amount of ore available for concentration. At the American-Boston, a concentrator for soft ores was used until the mine was shut down. This method depended on a special structure of low-grade ore. Crowell & Murray’s “Iron Ores of Lake Superior” gives valuable information on the various ores of this range.

Mining Methods in Use

There are no rich ores close enough to the surface to permit open pit mining. Doubtless many deposits, now exhausted, that were-opened underground could have been stripped with modern equipment and the ore mined at a profit. Only the lean ores are now mined in open pits. The factor deciding the question of open pit or underground mining is the cost of stripping the overburden as compared with underground cost. Climatic conditions do not interfere, as shipments are made only in the summer. In open-pit mining, the systems used are: steam shoveling directly into standard railway or narrow-gage cars; milling into raises to underground drifts, then tramming the ore to a shaft, where it is dumped, hoisted, and run into ore cars.

Varying conditions necessitate a number of minor differences in mining methods, because hardly any two orebodies in the district are of the same size, shape, or physical structure. Underground mining may be separated into hard- and soft-ore mining. The hard ores are comparatively unimportant, as only a few mines on this range contain this grade. The systems used are either breast stoping into rooms and pillars, or shrinkage stoping if the vein is narrow, steeply dipping and has firm hanging and foot walls.

Underground Mines

Most of the soft ore mined on the Marquette range is won by the top-slicing method. The typical orebody is a large mass with width exceeding thickness, lying in a basin of slate or diorite, or both, with a flat pitch and with the overlying jasper for a hanging wall or capping. The width may be as great as 1200 ft., the thickness 200 ft., and the length indefinite. In orebodies of such shape and dimension, top slicing is the only system by which great loss of ore can be avoided and the ore obtained without the grade being seriously affected by its being mixed with rock. The top-slicing system is flexible, in that any horses of jasper or large dikes can be left. By daily sampling from the breast of the working places, the ore can be hoisted and shipped, or stocked, according to grade. In large orebodies, slicing is carried on at different elevations, as extremely large sublevels are practically impossible to keep open. A large product can soon be obtained by starting work in a number of different places, each of which must be immediately below the hanging jasper. The disadvantage of this system is the amount of timber required and the heat generated by the decay of the timber. However, as practically all of the mines have two openings to the surface, good ventilation can be obtained by natural or mechanical means and the temperature kept within reasonable limits.

Mine Opening

Mine openings are entirely by shafts, all modern ones being vertical; the deepest in the district is about 2500 ft. The size regarded as standard is 10 ft. 10 in. by 14 ft. 10 in. inside. This is divided into four compartments, the arrangement of which is shown on Fig. 1. Modern shafts are constructed with concrete walls and steel sets; some are circular, others rectangular. Great care is taken to locate them in the solid foot wall at a point where they will not be disturbed by caving.

The arrangement of the loading and discharging pockets is shown in Fig. 2. There are usually three 60-ton storage pockets, from which the ore is drawn into two measuring pockets, each holding enough for one skip. In some cases, additional storage is gained by raises from one level

to another, thus avoiding the expense of installing pockets on each level. In some mines storage raises at the shafts, 200 ft. high, are satisfactory.

Underground Development Plans

In Fig. 3 are shown the main levels, sublevels, raises, and chutes, with their relative dimensions and intervals.


Drilling and Blasting.—In drifting through hard rock, water-feed hammer drills on cradles are used, while for drifting and raising in ore hand machines of the jackhamer type, fitted with auger bits, are employed. For rock raising, the stoper is in common use. For shaft-sinking, air or water-feed hand sinking machines are used. The size and shape of the steel and bits are shown in Fig. 4; Fig. 5 shows the arrangement and depth of holes for cuts in ore and rock drifts. For tamping,

paper bags filled with fine material are supplied. In most soft ores, the explosive is 40 and 50 per cent, low-freezing ammonia; while in some of the harder ores 50 and 60 per cent, gelatine is used, occasionally 80 per cent.

gelatine is necessary. The air pressure at the drill is usually between 70 and 80 pounds.

Drifting and Sloping.—In laying out the main haulage levels, parallel crosscuts at intervals of 150 ft. are driven. From these crosscuts, raises are put up at intervals of 40 to 60 ft.; these raises are carried through to the capping. At the top of the raises, immediately below the jasper, sublevels are started. Crosscuts are driven to the proper limit and slicing is commenced. As the ore is removed, the floors are well covered with either lagging or 5/8-in. covering down boards. Succeeding sub-levels are driven at intervals of from 10 to 12 ft. On each sublevel, during the process of slicing, the floors are well covered. Fig. 3 gives details for position of drifts and raises.

Timbering.—In the main and sublevel drifts, round hardwood, hemlock, and tamarack timber are used. On this range, timber has not been chemically treated, although preparations are being made to do this. It is thought that before long all of the timber for main levels will be chemically treated.

Timber that has been framed on the surface is delivered at the bottom of raises on timber cars. From these points it is hoisted to the working places by means of small air hoists. It is not the practice, on this range, to remove any timber during the process of working the caving system. The details of the timber framing are given in Fig. 6.

The legs of the standard level set are usually 8 ft. long, though 9-ft. legs are used at the bottom of raises and in some haulage drifts. Logs are delivered in 8-ft. and 16-ft. lengths, the 7-ft. stub from a 9-ft. leg being used as a short leg or cap in sublevels. For main-level sets, legs are 12 to 14 in. in diameter; for sublevels, 8 in. to 10 in. and 10 in. to 12 in. Sets are usually 5 ft. apart and are braced as shown. A longer brace, behind the lower one shown, is spiked to both legs. Lagging and, where necessary, blocking is placed above the caps.

Specifications for Timber

Stull Timber (Legs and Caps)

Hemlock, hard maple, soft maple, yellow birch, tamarack, and Norway pine. Must be sound, straight, and green. Ash, white birch, poplar and balm of gilead not accepted.

Tamarack in top diameters of 8 and 9 in. containing not more than ¼ in. sap rot in depth, accepted; in diameters of 10, 11, 12, and 13 in., containing not more than ½ in. sap rot in depth. Lengths, 8 and 16 ft.


Tamarack, spruce, pine, hemlock, maple, and yellow birch. Must be sound, straight and green. Top diameter, 6 to 8 in. Lengths, 5 ft. 4 in., 10 ft. 8 in., and 16 ft.


Straight sound cedar, tamarack, and spruce (10 per cent, jack pine permitted).
Round, 3- to 4½-in. top, larger than 4½ in. to be split.
Split, not less than 2½ by 4 nor greater than 3 by 6.
In 5-ft. lengths, 160 cu. ft. per cord.
In 7- and 8-ft. lengths, 128 cu. ft. per cord; not less than 125 pieces.

Covering-down Boards

No. 3 maple and birch. To be resawed from 2-in. maple and birch hearts to 5/8 in. To be sound, green lumber.
Widths, 6 in. and wider.
Lengths, not under 6 ft. and not over 9 ft.

Underground Track Ties

4 ft. 6 in. long, in- thick, 4½-in. face.


10 ft. long, cut from 20- and 30-ft. lengths, 3 to 4½ in.

Underground Sampling.—All working places are sampled at intervals of about 5 ft.; these samples make it possible to grade the ore. In addition, the ore from each chute is sampled as the motor cars are filled; this gives a check sample on the grade. All cars, before they are dumped into pockets at the shaft, are sampled, the samples being kept separate, according to mine chutes. On the surface, during the shipping season, a sample is taken from each skip as the ore runs into the railroad car; during the stocking season, it is taken from each car before it is dumped on the stock pile. The samples taken on the surface are the ones reported for the grade, those taken underground are simply used as a check.

Loading Machines and Scrapers.—Two heavy types of loading machines (the Shuveloder and the Hoar) have been successfully used in main level drifts. These machines are not practical, except on main levels. For sublevels, the John Mayne sublevel loader, Fig. 7, has been successfully used for over two years; this machine will be on the market in a short time. It is simple in construction and has few moving parts. It stands up under continuous work and can be operated by any miner. Records for a year for a gang using this loader show an increase over hand shoveling of 93.6 per cent, in tons per man per day; a decrease in

price to contractors of 32.6 per cent., and an increase in monthly earnings of miners of 19.5 per cent. Scrapers operated by double-drum air hoists have been used successfully in a limited way in both hard and soft ores. They load rapidly in a straight drift up to 75 ft. from a raise.

Tramming and Haulage.—Electric haulage is used almost exclusively in the district. Direct current is generated from alternating by rotary converters, situated on the surface or underground. Locomotives are 6 tons in weight and the current is received from an overhead trolley. The ore cars are of steel, side dumping with saddle backs, of 64 cu. ft. capacity, or approximately 4 tons of ore. The gage of tracks is 30 in. and the weight of rail 30 to 40 lb. The grade is 0.5 per cent, with the load. The cars are usually equipped with roller-bearing wheels and are dumped by hand at the shaft though the most recent installations have been rotary dumps, using round bottom cars.

Hoisting.—At most of the mines electricity is used for hoisting, although some still use steam. These hoists vary in horsepower from 400 to 900, depending on the depth of the shaft and the size of the skips. As alternating-current hoists of this size throw a heavy variable load on the power line, the newer hoists are equipped with flywheel motor-generator sets following the Ilgner system. With the skip hoist operated by direct current, an approximately constant power load is

maintained. The drums are from 8 to 10 ft. in diameter and from 8- to 14-ft. face. Hoisting is usually done in balance, at a speed from 1000 to 2000 ft. a min. For a 4-ton skip, which is the average size, a 1¼-in. plow-steel rope is used, except in the deeper shafts, which use 1 3/8-in. ropes. At almost all mines, a separate hoist is provided for the men. The cages have a capacity of from 24 to 30 men, and the speed, when men are being handled, is about 800 ft. per min. The hoists are provided with Lillie overwind device and the cages are equipped with safety catches and are balanced by counterweights. The counterweights are cast-iron cylinders, operating in 12-in. iron pipe. Most of the head frames are of steel, though one mine has an enclosed concrete structure.

Pumping.—Both plunger and centrifugal pumps are used, the motive power usually being electricity. These have capacities from 500 to 1600 gal. per min. against heads from 500 to 2400 ft. The amount of water pumped varies, in the different mines, from 150,000 to 3,500,000 gal. per day.

Air Compression.—Most mines are equipped with two-stage inter-cooled compressors of a capacity of 2000 cu. ft. per min.

Ventilation.—Nearly all mines have two openings and sufficient ventilation is therefore provided by natural means. In a few cases, where there is only one shaft and the connecting drift with another mine does not furnish sufficient ventilation, fans are installed underground. These are multivane blowers, with a capacity of 40,000 cu. ft. per min. against 3 in. water-gage pressure, operated by 50-hp. 2200-voIt a.c. motors. Where such installation has been made, the fan is placed on the bottom level, the cage compartment being used as the intake and the skip compartments as the outlet on the upper level, the lathing between these compartments having been made tight. By means of doors on the various levels, the air is forced through all the working places and discharged into the skip compartment on the top level. In order not to interfere with haulage, these doors are opened and closed by pneumatic cylinders controlled, from a distance, by ropes, red and green lights indicating the position of the doors.

Lighting.—In the shaft houses and on the underground plats, 55-volt, a.c. lamps are used; while in the main haulage levels 250-volt d.c. lamps are installed, which obtain electricity from the trolley wire. These are placed at all switches and at intervals from 100 to 200 ft. along the drifts. All men in the mines use carbide lamps.

Telephones.—Telephones are installed at all main level plats, in pump rooms, at underground hoists, in the lander’s station in the head frame, in hoisting houses, and in the various surface buildings, such as office and shops. Signaling in the shaft is by means of a.c. electric bells, all mine plats being connected to the engine house on two lines. The repeating bell system is used for operating cages. The cage rider is not permitted to open the door of the cage until the stop bell has been received from the hoisting engineer. A wire-pull bell is always installed in the cage com¬partment for emergency signals. No method of signaling in the haulage ways is used except that, on large levels operating more than one electric motor, there are colored lights at the entrance to each crosscut, which are automatically lighted when the motor is in that crosscut.

Disposition of Ore after Reaching Surface

During the shipping season, which is from May 1 to Dec. 1, ore from the skips goes directly into standard railway cars and is hauled to the ore docks at Marquette or Escanaba, for shipment by boat to lower lake ports. During the winter months, ore is stockpiled from trestles about 40 ft. high, usually built with wooden bents. As wooden trestles must be torn down each shipping season and re-erected in the fall, permanent steel stocking trestles are used at mines of a long life where there is sufficient ore to warrant the larger initial expenditure.

Safety and Welfare Work

Most of the large mining companies employ safety inspectors, who make regular or periodical trips through the mines. A book of safety rules is given to each employee, who signs a receipt for it. Examinations are held on these rules by a special committee. Failure to pass is sufficient cause for dismissal. Once a year, a special committee of workmen visits other mines to compare the safety appliances with rules as enforced in the different properties. This committe makes a number of safety suggestions, which are considered by the officials and almost invariably are accepted. At regular intervals, a certain number of men in each mine are trained in first-aid and apparatus work. Once a year, teams are selec¬ted from the various mines and field meets are held to compete for prizes in first-aid work.

The amount of welfare work varies with the different conditions. Some of the companies have hospitals, attached to which is a staff of doctors. In these hospitals, not only the cases resulting from mine accidents are treated, but also cases of sickness in the community. Nurses are employed, who visit the houses of the workmen, give help, and advise in case of sickness. A small uniform charge is paid monthly by all employees to cover doctors’ fees for ordinary consultation and visits, while the amount paid by the company for compensation takes care of cases of accidents. Some companies also have a system of pensions. Most of the companies take an active interest in the community by providing in the mine locations, where there are no theaters and other means of recreation, club houses with reading rooms, gymnasiums, bowling alleys, moving-picture machines, etc.

Precipitation Efficiency of Zinc Dust

It is generally realized that in cyaniding the precipitation efficiency of zinc dust is due to the fine division or extended surface of its metallic particles; but frequently it is thought that the presence of other metals, say2 to 3 percent, lead, is advantageous, causing more complete precipitation. The results of testing about fifty brands of commercial zinc dust have led to the conclusion that there is a distinct relation of precipitation efficiency to fineness and that the effect generally can be estimated by examining the size of metallic particles. The presence of lead was not found to be of any importance.


Generally the term “97 per cent, to pass a 350-mesh screen, 95 to 97 per cent, uncombined metallic zinc” is used by the leading European exporters. Among the many methods of determination of metallic zinc, I have found the iodine test (iodine in potassium iodide) very satisfactory and rapid. It has been controlled by the other methods, samples of the same product having been sent to three different analysts:

Iodine method………………………………………98.11 per cent, metallic zinc,
Volumetric method………….07.63 per cent, metallic zinc, Ledoux Co., New York
Bichromate method…………..98.16 per cent, metallic zinc, Wataon Gray, Liverpool, England
Bichromate method…………..98.20 per cent, metallic zinc, Norway Inst, of Tech., Trondhjem

Determination of Precipitation Efficiency

The method devised by W. J. Sharwood was used. The method proved to be satisfactory, the tests being merely for the comparison of different samples, hence the personal factor in manipulation was eliminated. As nearly all tests showed more zinc in solution than was accounted for by the silver precipitated, the term “dissolved zinc” was introduced—it means zinc dissolved by the action of cyanide and oxygen:

Zn + 4KCN + H20 + O = K2Zn(CN)4 + 2KOH;

or more probably, resolution of its equivalent precipitated silver, as tests stirred two hours showed more “dissolved zinc” than those stirred

one hour. To determine “dissolved zinc” the silver precipitate was dissolved in nitric acid, silver titrated with thiocyanate, and solution titrated with ferrocyanide (after removing silver precipitate) giving the amount of intact metallic zinc left in the zinc dust silver precipitate. The difference between active plus intact zinc and total metallic zinc is “dissolved zinc.”

Sources of Zinc Dust Tested

Samples 1 to 4 are American zinc dust; 6 to 8 are Norwegian, electrothermic fumed dust; 9 to 14 are from an electrothermic experimental plant; 15, origin is unknown, sample was furnished by Ste. Generale de Commerce & Exterieur, Paris; 16 is Belgian dust; 17, German; and 18, electrothermic blue powder (byproduct from electrothermic zinc smelting). Samples 4 and 5 were atomized, the other distilled.

Precipitation Efficiency as a Factor of Fineness

The microscopic examination showed that the distilled zinc dust consists of almost perfect spherules. The appearance is almost clean and metallic except in samples 16 and 17, where numerous particles of oxide are shown. The atomized dust (samples 4 and 5) has a coke-like surface and is very coarse; especially sample 5, which was made by a 100 lb. air pressure.

The number of spheres in a pound zinc dust, assuming the specific gravity as 7. is 0.1235/d³, hence

if the diameter is 1 mm., 1 lb. will contain 123,500 particles
if the diameter is 0.1 mm., 1 lb. will contain 1235 million particles
if the diameter is 0.01 mm., 1 lb. will contain 123,500 million particles
if the diameter is 0.003 mm., 1 lb. will contain 4600 billion particles

In sample 15, spherules of 0.003 mm. diameter were found to be pre-dominant, hence the number of particles in 1 lb. (89 per cent, metallic zinc) is 4600 billion X 0.89 = 4100 billion. In the precipitate was left 28 per cent, of the dust’s metallic zinc content, the diameter of remaining intact zinc spherules is then: 4100 billion = 0.1235/d³X 0.28 and d = 0.002 mm. The original spherule was 0.003 mm., hence the thickness of active surface is 0.0005 mm.

In sample 8, the major particles were of 0.004 mm. diameter. The number of particles in 1 lb., is 2000 billion. In the precipitate was left 43 per cent, of the metallic zinc, then the diameter of the intact particle is 2000 billion = 0.1235/d³X 0.43 and d = 0.003 mm. The thickness of active surface is 0.0005 mm. and so forth.

In the same manner, the efficiency of a zinc dust may be estimated on the basis of fineness as:

The diameter examined being 0.002 mm., the efficiency is 88 per cent.
The diameter examined being 0.003 mm., the efficiency is 70.3 per cent.
The diameter examined being 0.004 mm., efficiency is 58.0 per cent.
The diameter examined being 0.008 mm., efficiency is 33.2 per cent.
The diameter examined being 0.01 mm., efficiency is 27.0 per cent.
The diameter examined being 0.03 mm., efficiency is 9.7 per cent.
The diameter examined being 0.07 mm., efficiency is 3.5 per cent.
To get a fair comparison between the found efficiencies and those from fineness estimated values, it is necessary to eliminate what is called “dissolved” zinc. This is possible by figuring the precipitation efficiency from the difference of metallic zinc left in precipitate; which is here called “true efficiency.”

From the foregoing data there is little doubt as to what role fineness is playing. The consumer frequently calls for high content of metallic zinc, but mostly he buys in accord with the efficiency obtained in practical running. The producer should, therefore, direct his attention to improving the fineness—under maintenance of the highest content of metallic zinc—until it becomes really fume.


G. M. Brown, New York, N. Y.—Andre Dorfmann of the Mclntyre Porcupine Mine made a similar test, some years ago, relative to consumption of zinc and the results he obtained confirm the statements in this paper. In addition to the amount of zinc left in the precipitation presses, he determined the amount of zinc in the barren solution. As this solution was recirculated through the system, he also determined the amount of zinc precipitated from solution, in the ball-mill, tube-mill and agitators, before the solution was again returned to the precipitation presses.

Charles E. Locke, Cambridge, Mass.—The thing which struck me in looking through the table is that the maximum figure is about 60 per cent, efficiency, when based on the metallic zinc content, and the ordinary efficiency, if I interpret it correctly, ranges from a maximum of 57 per cent, with the finest dust down to 6 per cent, with some rather coarse samples of zinc dust.