PART V

LIME-ROASTING OF GALENA

THE HUNTINGTON-HEBERLEIN PROCESS

(July 6, 1905)

It is a fact, not generally known, that the American Smelting and Refining Company is now preparing to introduce the Huntington-Heberlein process in all its plants, this action being the outcome of extensive experimentation with the process. It is contemplated to employ the process not only for the desulphurization of all classes of lead ore, but also of mattes. This is a tardy recognition of the value of a process which has been before the metallurgical profession for nine years, the British patent having been issued under date of April 16, 1896, and has already attained important use in several foreign countries; but it will be the grandest application in point of magnitude.

The Huntington-Heberlein is the first of a new series of processes which effect the desulphurization of galena on an entirely new principle and at great advantage over the old method of roasting. They act at a comparatively low temperature, so that the loss of lead and silver is reduced to insignificant proportion; they eliminate the sulphur to a greater degree; and they deliver the ore in the form of a cinder, which greatly increases the smelting speed of the blast furnace. They constitute one of the most important advances in the metallurgy of lead. The roasting process has been the one in which least progress has been made, and it has remained a costly and wasteful step in the treatment of sulphide ores. In reducing upward of 2,500,000 tons of ore per annum, the American Smelting and Refining Company is obliged to roast upward of 1,000,000 tons of ore and matte.

The Huntington-Heberlein process was invented and first applied at Pertusola, Italy. It has since been introduced in Germany, Spain, Great Britain, Mexico, British Columbia, Tasmania, and Australia, in the last at the Port Pirie works of the Broken Hill Proprietary Company. Efforts were made to introduce it in the United States at least five years ago, without success and with little encouragement. The only share in this metallurgical improvement that this country can claim is that Thomas Huntington, one of the inventors, is an American citizen, Ferdinand Heberlein, the other, being a German.

LIME-ROASTING OF GALENA

(September 22, 1905)

The article of Professor Borchers (see p. 116) is, we believe, the first critical discussion of the reactions involved in the new methods of desulphurizing galena, as exemplified in the processes of Huntington and Heberlein, Savelsberg, and Carmichael and Bradford, although the subject has been touched upon by Donald Clark, writing in the _Engineering and Mining Journal_. It is perfectly obvious from a study of the metallurgy of these processes that they introduce an entirely new principle in the oxidation of galena, as Professor Borchers points out. Inasmuch as there are already three of these processes and are likely to be more, it will be necessary to have a type-name for this new branch of lead metallurgy. We venture to suggest that it may be referred to as the “lime-roasting of galena,” inasmuch as lime is evidently a requisite in the process; or, at all events, it is the agent which will be commonly employed.

When the Huntington-Heberlein process was first described, it did not even appear a simplification of the ordinary roasting process, but rather a complication of it. The process attracted comparatively little attention, and was indeed regarded somewhat with suspicion. This was largely due to the policy of the company which acquired the patent rights in refusing to publish the technical information concerning it that the metallurgical profession expected and needed. The history of this exploitation is another example of the disadvantage of secrecy in such matters. The Huntington-Heberlein process has only become thoroughly established as a new and valuable departure in metallurgy, a departure which is indeed revolutionary, nine years after the date of the original patent. In proprietary processes time is a particularly valuable element, inasmuch as the life of a patent is limited.

From the outset the explanation of Huntington and Heberlein as to the reactions involved in their process was unsatisfactory. Professor Borchers points out clearly that their conception of the formation of calcium peroxide was erroneous, and indicates strongly the probability that the active agent is calcium plumbate. It is very much to be regretted that he did not go further with his experiments on this subject, and it is to be hoped that they will be taken up by the professors of metallurgy in other metallurgical schools. The formation of calcium plumbate in the process was clearly forecasted, however, by Carmichael and Bradford in their first patent specification; indeed, they considered that the sintered product consisted largely of calcium plumbate.

Even yet, we have only a vague idea of the reactions that occur in these processes. There is undoubtedly a formation of calcium sulphate, as pointed out by Borchers and Savelsberg; but that compound is eventually decomposed, since it is one of the advantages of the lime-roasting that the sintered product is comparatively low in sulphur. Is it true, however, that the calcium eventually becomes silicate? If so, under what conditions is calcium silicate formed? The temperature maintained throughout the process is low, considerably lower than that required for the formation of any calcium silicate by fusion.

Moreover, it is not only galena which is decomposed by the new method, but also blende, pyrite and copper sulphides. The process is employed very successfully in the treatment of Broken Hill ore that is rather high in zinc sulphide, and it is also to be employed for the desulphurization of mattes. What are the reactions that affect the desulphurization of the sulphides other than lead?

There is a wide field for experimental metallurgy in connection with these new processes. The important practical development is that they do actually effect a great economy in the reduction of lead sulphide ores.

THE NEW METHODS OF DESULPHURIZING GALENA[18]

BY W. BORCHERS

(September 2, 1905)

An important revolution in the methods of smelting lead ore, which had to a large extent remained for centuries unchanged in their essentials, was wrought by the invention of Huntington and Heberlein in 1896. More especially is this true of the roast-reduction method of treating galena, which consists of oxidizing roasting in a reverberatory furnace and subsequent smelting of the roasted product in a shaft furnace.

The first stage of the roast-reduction process, as carried out according to the old method, viz., the oxidizing roast of the galena, serves to convert the lead sulphide into lead oxide:

PbS + 3O = PbO + SO₂.

Owing to the basic character of the lead oxide, the production of a considerable quantity of lead sulphate was of course unavoidable:

PbO + SO₂ + O = PbSO₄.

As this lead sulphate is converted back into sulphide in the blast-furnace operation, and so adds to the formation of matte, it has always been the aim (in working up ores containing little or no copper to be concentrated in the matte) to eliminate the sulphate as completely as possible, by bringing the charge, especially toward the end of the roasting operation, into a zone of the furnace wherein the temperature is sufficiently high to effect decomposition of the sulphate by silica:

PbSO₄ + SiO₂ = PbSiO₃ + SO₃.

But in the usual mode of carrying out the roast in reverberatory furnaces, the roasting itself on the one hand, and the decomposition of the sulphates on the other, were effected only incompletely and with widely varying results.

Little attention has been paid in connection with the roast-reduction process to the reaction between sulphates and undecomposed sulphides, which plays so important a part in the roast-reaction method of lead smelting. As is well known, lead sulphate reacts with lead sulphide in varying quantities, forming either metallic lead or lead oxide, or a mixture of both. A small quantity of lead sulphate reacting with lead sulphide yields under certain conditions only lead:

PbSO₄ + PbS = Pb₂ + 2SO₂.

Within certain temperature limits this reaction even proceeds with liberation of heat. In order to encourage it, it is necessary to create favorable conditions for the formation of considerable quantities of sulphate right at the beginning of the operation. This was first achieved by Huntington and Heberlein, but not in the simplest nor in the most efficient manner. And, indeed, the inventors were not by any means on the right track as to the character of their process, so far as the chemical reactions involved are concerned.

At first sight the Huntington-Heberlein process does not even appear as a simplification, but rather as a complication, of the roasting operation. For in place of the roast carried out in one apparatus and continuously, there are two roasts which have to be carried out separately and in two different forms of apparatus; nevertheless, the ultimate results were so favorable that the whole process is presumably acknowledged, without reservation, by all smelters as one of the most important advances in lead smelting.

It is useful to examine in the light of the German patent specification (No. 95,601 of Feb. 28, 1897) what were the ideas of its originators regarding the operation of this process and the reactions leading to such remarkable results. They stated:

“We have made the observation that when powdered lead sulphide (PbS), mixed with the powdered oxide of an alkaline earth metal, _e.g._, calcium oxide, is exposed to the action of air at bright red heat (about 700 deg. C.), and is then allowed to cool without interrupting the supply of air, an oxidizing decomposition takes place when dark-red heat (about 500 deg. C.) is reached, sulphurous acid being expelled, and a considerable amount of heat evolved; if sufficient air is then continuously passed through the charge, dense vapors of sulphurous acid escape, and the mixture gradually sinters together to a mass, in which the lead of the ore is present in the form of lead oxide, provided the air blast is continued long enough; there is no need to supply heat in this process—the heat liberated in the reaction is quite sufficient to keep it up.”

The inventors explained the process as follows:

“At a bright-red heat the calcium oxide (CaO) takes up oxygen from the air supplied, forming calcium peroxide (CaO₂), which latter afterward, in consequence of cooling down to dark-red heat, again decomposes into monoxide and oxygen; this nascent oxygen oxidizes a part of the lead sulphide to lead sulphate, which then reacts with a further quantity of lead sulphide, with evolution of sulphur dioxide and formation of lead oxide.”

Assuming the formation of calcium peroxide (CaO₂), the process leading to the desulphurization would therefore be represented as follows:

1. at 700° C. CaO + O = CaO₂ 2. at 500° C. 4CaO₂ + PbS = 4CaO + PbSO₄ 3. at the melting point PbS + PbSO₄ = 2PbO + 2SO₂ (?)

Reactions 1 and 2 combined, assuming the presence of sufficient oxygen, give:

PbS + 4CaO + 4O = PbSO₄ + 4CaO.

Now the invention consists in applying the observation described above to the working up of galena, and other ores containing lead sulphide, for metallic lead; and the essential novelty of the process therefore consists in passing air through the mass cooled to a dark-red heat (500 deg. C.).

This feature sharply distinguishes it from other known processes. It is true that in previous processes (compare the Tarnowitz reverberatory-furnace process, the roasting process used at Munsterbusch near Stolberg, and others) the lead ore was mixed with limestone or dolomite (which are converted into oxides in the early stage of the roast) and the heat was alternately raised and lowered; but in all cases only a surface action of the air was produced, the air supply being provided simply by the furnace draft. Passing air through the mass cooled down, as indicated above, leads to the important economic advantages of reducing the fuel consumption, the losses of lead, the manual labor (raking) and the dimensions of the roasting apparatus.

In order to carry out the process of this invention, the powdered ore is intimately mixed with a quantity of alkaline earth oxide, _e.g._, calcium oxide, corresponding to its sulphur content; if the ore already contains alkaline earth, the quantity to be added is reduced in accordance. The mixture is heated to bright-red heat (700 deg. C.) in the reverberatory furnace, in a strongly oxidizing atmosphere, is then allowed to cool down to dark-red heat (500 deg. C.), also in strongly oxidizing atmosphere, is transferred to a vessel called the “converter,” and atmospheric air is passed through at a slight pressure (the inventors have found a blast corresponding to 35 to 40 cm. head of water suitable).[19] The heat liberated is quite sufficient to keep the charge at the reaction temperature, but, if desired, hot blast may also be used. The mixture sinters together, and (while sulphurous acid gas escapes) it is gradually converted into a mass consisting of lead oxide, gangue and calcium sulphate, from which the lead is extracted in the metallic form, by any of the known methods, in the shaft furnace. The operation is concluded as soon as the mass, by continued sintering, has become impermeable to the blast. If the operation is properly conducted, the gas escaping contains only small quantities of volatile lead compounds, but on the other hand up to 8 per cent. by volume of sulphur dioxide. This latter can be collected and further worked up.

“In place of the oxide of an alkaline earth, ferrous oxide (FeO) or manganous oxide (MnO) may also be used.”

According to the reports on the practice of this process which have been published,[20] conical converters of about 1700 mm. (5 ft. 6 in.) upper diameter and 1500 mm. (5 ft.) depth are used in Australian works. At a new plant at Port Pirie (Broken Hill Proprietary Company) converters 2400 mm. (7 ft. 10 in.) in diameter and 1800 mm. (5 ft. 11 in.) deep have been installed. These latter will hold a charge of about eight tons. In the lower part of these converters, at a distance of about 600 mm. (2 ft.) from the bottom, there is placed an annular perforated plate, and upon this a short perforated tube, closed above by a plate having only a limited number of holes.

No details have been published with regard to the European installations. The general information which the Metallurgische Gesellschaft[21] placed at my disposal upon request some years ago, for use in my lecture courses, was restricted to data regarding the consumption of fuel and labor in roasting and smelting the ores, which was figured at about one-third or one-half of the consumption in the former processes, to the demonstration of the large output of the comparatively small converters, and to the reduced size of the roasting plant as the result. But the European establishments which introduced this process were bound by the owners of the patents, notwithstanding the protection afforded by the patents, to give no information whatever regarding the process to outsiders, and not to allow any inspection of the works.

On the other hand, a great deal appeared in technical literature which was calculated to excite curiosity. Moreover, as professor of metallurgy, it was my duty to instruct my pupils concerning this process among others, and it was therefore very gratifying to me that one of the students in my laboratory took a special interest in the treatment of lead ore. I gave him opportunity to install a small converter, in order to carry out the process on a small scale, and in spite of the slender dimensions of the apparatus the very first experiments gave a complete success.

However, I could not harmonize the explanation of the process given by the inventors with the knowledge which I had acquired in my many years’ practical experience in the manufacture of peroxides. It is clear from the patent specification that in the roasting operation at 700 deg. C. a compound must be formed which functions as an excellent oxygen carrier, for on cooling to 500 deg. C. the further oxidation then proceeds to the end not only without any external application of heat, but even with vigorous evolution of heat. No more striking instance than this could be desired by the theorists who have of recent years again become so enthusiastic over the idea of catalysis. Huntington and Heberlein regarded calcium peroxide as the oxygen carrier, but that is a compound which cannot exist at all under the conditions which obtain in their process. The peroxides of the alkaline earths are so very sensitive that in preparing them the small quantities of carbon dioxide and water must be extracted carefully from the air, and yet in the process, in an atmosphere pregnant with carbon dioxide, water, sulphurous acid, etc., calcium peroxide, the most sensitive of the whole group, is supposed to form! This could not be.

The only compounds known as oxygen carriers, and capable of existing under the conditions of the process, are calcium plumbate and plumbite. I have emphasized this point from the first in my lectures on metallurgy, when dealing with the Huntington-Heberlein process, and, in point of fact, this assumption has since been proved to be correct by the work of L. Huppertz, one of my students.

During my practical activity (1879-1891) I had prepared barium peroxide and lead peroxide in large quantities on a manufacturing scale, the last-mentioned through the intermediate formation of plumbites and plumbates:

2NaOH + PbO + O = Na₂PbO₃ + H₂O

or:

4NaOH + PbO + O = Na₄PbO₄ + 2H₂O.

An experiment made in this connection showed that calcium plumbate is formed just as readily from slaked lime and litharge as the sodium plumbates above. Litharge is an intermediate product, produced in large quantities in lead works, and must in any case be brought back into the process. If, then, the litharge is roasted at a low temperature with slaked lime, the roasting of the galena could perhaps be entirely avoided by introducing that ore together with calcium plumbate into the converter, after the latter had once been heated up. Mr. Huppertz undertook the further development of this process, but I have no information on the later experimental results, as he placed himself in communication with neighboring lead works for the purpose of continuing his investigation, and has not since then given me any precise data. I will therefore confine myself to the statement that the fundamental idea for the experiments, which Mr. Huppertz undertook at my suggestion, was the following:

To dispense with the roasting of the galena, which is necessary according to Huntington and Heberlein; in other words, to convert the galena by direct blast, with the addition of calcium plumbate, the latter being produced from the litharge which is an unavoidable intermediate product in the metallurgy of lead and silver. (Borchers, “Elektrometallurgie,” 3d edition, 1902-1903, p. 467.)

This alone would, of course, have meant a considerable simplification of the roast, but the problem of the roasting of galena has been solved in a better way by A. Savelsberg, of Ramsbeck, Westphalia, who has determined the conditions for directly converting the galena with the addition of limestone and water and without previous roasting. He has communicated the following information regarding these conditions:

In order that, in blowing the air through the mixture of ore and limestone, an alteration of the mixture may not take place owing to the lighter particles of the limestone being carried away, it is necessary (quite at variance with the processes in use hitherto, in which for the sake of economy stress is laid on the precaution of charging the ore as dry as possible into the apparatus) to add a considerable quantity of water to the charge before introducing it into the converter. The water serves this purpose perfectly, also preventing any change in the mixture of ore and limestone, which invariably occurs if the ore is used dry. The water, moreover, exerts a very beneficial action in the process, inasmuch as it aids materially in the formation and temporary retention of sulphuric acid, which latter then, by its oxidizing action, greatly enhances the reaction and consequently the desulphurization of the ore. Furthermore, the water tends to moderate the temperature in the charge by absorbing heat in its volatilization.

In carrying out the process the converter must not be filled entirely all at once, but first only in part, additional layers being charged in gradually in the course of the operation. In this way a uniform progress of the reaction in the mass is secured.

The following mode of procedure is advantageously adopted: A small quantity of glowing fuel (coal, coke, etc.) is introduced into the converter, which is provided at the bottom with a grate (perforated sheet iron), the grate being first covered with a thin layer of crushed limestone in order to protect it from the action of the red-hot coals and ore. Upon this red-hot fuel a uniform layer of the wetted mixture of crude ore and limestone is placed. When the surface of the first layer has acquired a uniform red heat, a fresh layer is charged on, and this is continued, layer by layer, until the converter is quite full. While the layers are still being put on, the blast is passed in at quite a low pressure, and only when the converter is entirely filled is the whole force of the blast, at a rather greater pressure, turned on. There then sets in a kind of slag formation, which, however, is preceded by a very vigorous desulphurization. After the termination of the process, which can be recognized by the fact that vapors cease to be evolved, and that the surface of the ore becomes hard, the converter is tipped over, and the desulphurized mass drops out as a solid cone of slag, which is then suitably broken up for the subsequent smelting in the shaft furnace.

Savelsberg explains the reaction of this process as follows:

“1. The particles of limestone act mechanically, gliding in between the particles of lead ore and separating them from one another. In this way a premature sintering is prevented, and the whole mass is rendered loose and porous.

“2. The limestone moderates the reaction temperature produced in the combustion of the sulphur, so that the fusion of the galena, the formation of dust and the separation of metallic lead are avoided, or at least kept within the limits permissible. The lowering of the temperature of reaction is due partly to the decomposition of the limestone into caustic lime and carbon dioxide, in which heat is absorbed, and partly to the consumption of the quantity of heat which is necessary in the further progress of the operation for the formation of a slag from the gangue of the ore and the lead oxide produced.

“3. The limestone gives rise to chemical reactions. By its decomposition it produces lime, which, at the moment of its formation, is converted into calcium sulphate at the expense of the sulphur in the ore. The calcium sulphate at the time of slag formation is converted into silicate by the silica present, sulphuric acid being evolved. The limestone therefore assists directly and forcibly in the desulphurization of the ore, causing the formation of sulphuric acid at the expense of the sulphur in the ore, the sulphuric acid then acting as a strong oxidizing agent toward the sulphur in the ore.”

The most conclusive proof for the correctness of the opinion which I expressed above, that it is very important to create at the beginning of the operation the conditions for the formation of as much sulphate as possible, has been furnished by Carmichael and Bradford. They recommend that gypsum be added to the charge in place of limestone. At one of the works of the Broken Hill Proprietary Company (where their process has been carried on successfully, and where lead ores very rich in zinc had to be worked up) the dehydrated gypsum was mixed with an equal quantity of concentrate and three times the quantity of slime from the lead ore-dressing plant, as in the table given herewith:

─────────────────┬────────┬─────────────┬──────────┬──────── │ OF THE │ OF THE │ OF THE │ OF THE CONTENTS │ SLIME │ CONCENTRATE │ CALCIUM │ WHOLE │ │ │ SULPHATE │ CHARGE ─────────────────┼────────┼─────────────┼──────────┼──────── Galena │ 24 │ 70 │ │ 29 Zinc blende │ 30 │ 15 │ │ 21 Pyrites │ 3 │ │ │ 2 Ferric oxide │ 4 │ │ │ 2.5 Ferrous oxide │ 1 │ │ │ 1 Manganous oxide │ 6.5 │ │ │ 5 Alumina │ 5.5 │ │ │ 3 Lime │ 3.5 │ │ 4.1 │ 10 Silica │ 23 │ │ │ 14 Sulphur trioxide │ │ │ 59 │ 12 ─────────────────┴────────┴─────────────┴──────────┴────────

The charge is mixed, with addition of water, in a suitable pug-mill. The mass is then, while still wet, broken up into pieces 50 mm. (2 in.) in diameter, which are then allowed to dry on a floor in contact with air; in doing so they set hard, owing to the rehydration of the gypsum.

As in the case of the Savelsberg process, the converters are heated with a small quantity of coal, are filled with the material prepared in the manner above described, and the charge is blown, regulating the blast in such manner that, after the moisture present has been dissipated, a gas of about 10 per cent. SO₂ content is produced, which is worked up for sulphuric acid in a system of lead chambers.

The reactions are in this case the same as in the Savelsberg process, for here also calcium sulphate is formed transitorily, which, like other sulphates, reacts partly with sulphides, partly with silica.

Where gypsum is available and cheap, the Carmichael-Bradford process must be given preference; in all other cases unquestionably the Savelsberg process is superior, owing to its great simplicity.

LIME-ROASTING OF GALENA

BY W. MAYNARD HUTCHINGS

(_October 21, 1905_)

Much interest attaches to the paper by Professor Borchers, recently presented in the _Engineering and Mining Journal_ (Sept. 2, 1905) on “New Methods of Desulphurizing Galena,” together with an editorial on “Lime-Roasting of Galena”; it is a curious coincidence that the same issue contained also an article on the “Newer Treatment of Broken Hill Sulphides,” in which is shown the importance of the new methods as a contribution to actual practice.

For some years it had been a source of surprise to me that a new process, so interesting and so successful as the Huntington-Heberlein treatment of sulphide ores, should have received scarcely any notice or discussion. This lack, however, now appears to be remedied. The suggestion that the subject should be discussed in the _Journal_ is good, as is also that of the designation “Lime-Roasting” for a type-name. Such observations and experiments on the subject as I have had occasion to record have, for many years, figured in my note-books under that heading.

Whatever may be the final results of the later processes, now before the metallurgical world or still to come, there can be no doubt whatever that full and exclusive credit must be given to Huntington and Heberlein, not only for first drawing attention to the use of lime, but also for working out and introducing practically the process. It has been a success from the first; and so far as part of it is concerned, it seems to be an absolute and fundamental necessity which later inventors can neither better nor set aside. The other processes, since patented, however good they may be, are simply grafts on this parent stem.

It is, however, quite certain that Huntington and Heberlein, in the theoretical explanation of the process, failed to understand the most important reactions. Their attributing the effect to the formation and action of calcium peroxide affords a sad case of _a priori_ assumption devoid of any shred of evidence. As Professor Borchers points out, calcium peroxide, so difficult to produce and so unstable when formed, is an absolute and absurd impossibility under the conditions in question. Probably many rubbed their eyes with astonishment on reading that part of the patent on its first appearance, and hastened to look up the chemical authorities to refresh their minds, lest something as to the nature of calcium peroxide might have escaped them.

Fortunately the patent law is such that there was no danger of a really good and sound invention being invalidated by a wrong theoretical explanation by its originators. But, nevertheless, it was a misfortune that the inventors did not understand their own process. Had they known, they could have added a few more words to their patent-claims and rendered the Carmichael patent an impossibility.

Professor Borchers appears to consider that the active agent in the new process is calcium plumbate. That this compound may play a part at some stage of the process may be true; this long ago suggested itself to some others. We may yet expect to hear that the experiments undertaken by Professor Borchers himself, and by others at his instigation (in which calcium plumbate is separately prepared and then brought into action with lead sulphide), have given good results. But it does not appear so far that there is any real proof that calcium plumbate is formed in the Huntington-Heberlein or other similar processes; and it is difficult to see at what stage or how it would be produced under the conditions in question. This is a point which research may clear up, but it should not be taken for granted at this stage. Indeed, it seems to me that the results obtained may be fairly well explained without calling calcium plumbate into play at all.

Of course the action of lime in contact with lead sulphide excited interest many years before the new process came into existence. My own attention to it dates back more than a dozen years before that time (I was in charge of works where I found the old “Flintshire process” still in use).

Percy pointed out, in his work on lead smelting, that on the addition of slaked lime to the charge, at certain stages, to “stiffen it up,” the mixture could be seen to “glow” for a time. When I myself saw this phenomenon, I commenced to make some observations and experiments. Also (as others probably had done), I had observed that charges of lead with calcareous gangue are roasted more rapidly and better than others, and to an extent which could not be wholly explained by simple physical action of the lime present.

Simple experiments made in assay-scorifiers in a muffle, on lime roasting, are very striking, and I think quite explain a good part of what takes place up to a certain stage in the processes now under consideration. I tried them a number of years ago, on many sorts of ore, and again more recently, when studying the working of the new patents. For illustration, I will take one class of ore (Broken Hill concentrate), using a sample assaying; Pb, 58 per cent.; Fe, 3.6 per cent.; S, 14.6 per cent.; SiO₂, 3 per cent. The ore contained some pyrite. If two scorifiers are charged, one with the finely powdered ore alone, and one with the ore intimately mixed with, say, 10 per cent. of pure lime, and placed side by side just within a muffle at low redness, the limed charge will soon be seen to “glow.” Before the simple ore charge shows any sign of action, the limed charge rapidly ignites all over, like so much tinder, and heats up considerably above the surrounding temperature, at the same time increasing noticeably in bulk. This lasts for some time, during which hardly any SO₂ passes off. After the violent glowing is over, the charge continues to calcine quietly, giving off SO₂, but is still far more active than its neighbor. If, finally, the fully roasted charge is taken out, cooled and rubbed down, it proves to contain no free lime at all, but large quantities of calcium sulphate can be dissolved out by boiling in distilled water. For instance, in one example where weighed quantities were taken of lime and the ore mentioned, the final roasted material was shown to contain nearly 23 per cent. of CaSO₄; the quantity actually extracted by water was 20.2 per cent. Further tests show that the insoluble portion still contains calcium sulphate intimately combined with lead sulphate, but not extractable by water.

There is no doubt that when lead sulphide (or other sulphide) is heated with lime, with free access of air, the lime is rapidly and completely converted into sulphate. The strong base, lime, apparently plays the part of “catalyzer” in the most vigorous manner, the first SO₂ evolved being instantly oxidized and combined with the lime to sulphate, with so strong an evolution of heat that the operation spreads rapidly and still goes on energetically, even if the scorifier is taken out of the muffle. Also, the “catalytic” action starts the oxidation of the sulphides at a far lower temperature than is required when they are roasted alone.

If, in place of lime, we take an equivalent weight of pure calcium carbonate and intimately mix it with ore, we obtain just the same action, only it takes a little longer to start it. Once started, it is almost as vigorous and rapid, and with the same results. It does not seem correct to assume (as is usually done) that the carbonate has first to be decomposed by heat, the lime then coming into action. The reaction commences in so short a time and while the charge is still so cool, that no appreciable driving off of CO₂ by heat only can have taken place. The main liberation of the CO₂ occurs during the vigorous exothermic oxidation of the mixture, and is coincident with the conversion of the CaO into CaSO₄.

If, in place of lime or its carbonate, we use a corresponding quantity of pure calcium sulphate and mix it with the ore, we see very energetic roasting in this case also, with copious evolution of sulphur dioxide, only it is much more energetic and rapid and occurs at a lower temperature than in the case of a companion charge of ore alone.

It is very easily demonstrated that the CaSO₄ in contact with the still unoxidized ore (whether it has been introduced ready made or has been formed from lime or limestone added) greatly assists the further roasting, in acting as a “carrier” and enabling calcination to take place more rapidly and easily and at a lower temperature than would otherwise be the case.

The result of these experiments (whether we mix the ore with CaO, CaCO₃, or CaSO₄) is that we arrive with great ease and rapidity at a nearly dead-sweet roast. The lime is converted into sulphate, and the lead partly to sulphate and partly to oxide. Two examples out of several, both from the above ore, gave results as follows:

No. 1—Roasted with 20 per cent. CaCO₃ (= 11.2 per cent. CaO); sulphide sulphur, 0.02 per cent.; sulphate sulphur, 9.30 per cent.; total sulphur, 9.32 per cent.

No. 8—Roasted with 27.2 per cent. CaSO₄ (= 11 per cent. CaO); sulphide sulphur, 0.05 per cent.; sulphate sulphur, 11.28 per cent.; total sulphur, 11.33 per cent.

If these calcined products are now intimately mixed with additional silica (in about the proportions used in the Huntington-Heberlein process) and strongly heated, fritting is brought about and the sulphur content is reduced by the decomposition of the sulphates by the silica. Thus, the resultant material of experiment No. 1, above, when treated in this manner with strong heat for three hours, was sintered to a mass which was quite hard and stony when cold, and which contained 6.75 per cent. of total sulphur. Longer heating drives out more sulphur, but a very long time is required; in furnaces, and on a large scale, it is with great difficulty and cost that a product can be obtained comparable with that which is rapidly and cheaply turned out from the “converters” of the new process.

To return to the Huntington-Heberlein process, working, for example, on an ore more or less like the one given above, we may assume that, during the comparatively short preliminary roast, the lime is all rapidly converted into CaSO₄ and that some PbSO₄ is also formed (but not much, as the mixture to be transferred from the furnace to the converter requires not less than 6 to 8 per cent. of sulphur to be still present as sulphide, in order that the following operation may work at its best). As the blast permeates the mass, oxidation is energetic; no doubt that CaSO₄ here plays a very important part as a carrier of oxygen, in the same manner as we can see it act on a scorifier or on the hearth of a furnace.

What the later reactions are does not seem so clear. They are quite different from those on the scorifier or on the open hearth of a furnace, and result in the rapid formation (in successive layers of the mixture, from the bottom upward) of large amounts of lead oxide, fluxing the silica and other constituents to a more or less slaggy mass, which decomposes the sulphates and takes up the CaO into a complex and easily fused silicate. It is true that, as a whole, the contents of a well-worked converter are never very hot, but locally (in the regions where the progressive reaction and decomposition from below upward is going on) the temperature reached is considerable. This formation of lead oxide is so pronounced at times that one may see in the final product considerable quantities of pure uncombined litharge.

When the work is successful, the mass discharged from the converters is a basic silicate of PbO, CaO, and oxides of other metals present, and nearly all the sulphates have disappeared. A large piece of yellow product (which was taken from a well-worked converter) contained only 1.1 per cent. of total sulphur.

It may be that calcium plumbate is formed and plays a part in these reactions; but its presence would be difficult to prove, and its formation and existence during these stages would not be easy to explain. Neither does it seem necessary, as the whole thing appears to be capable of explanation without it.

While the mixture in the converter is still dry and loose, energetic oxidation of the sulphides goes on, with the intervention of the CaSO₄ as a carrier. As soon as the heat rises sufficiently, fluxing commences in a given layer and sulphates are decomposed. The liberated sulphuric anhydride, at the locally high temperature and under the existing conditions, will act with the greatest possible vigor on the sulphides in the adjacent layers; these layers will then in their turn flux and act on those above them, till the whole charge is worked out. The column of ore is of considerable hight, requiring a blast of 1½ lb., or perhaps more, in the larger converters now used. This pressure of the oxidizing blast (and of the far more powerfully oxidizing sulphuric anhydride, continuously being liberated within the mass of ore, locally very hot) constitutes a totally different set of conditions from those obtained on the hearth of a furnace with the ore in thin layers, where it is neither so hot nor under any pressure. It is to these conditions, in which we have the continued intense action of red-hot sulphuric anhydride under a considerable pressure (together with the earlier action of the CaSO₄), that the remarkable efficiency of the process seems to me to be due.

In the Carmichael process, the preliminary roast is done away with, CaSO₄ being added directly instead of having to be formed during the operation from CaO and the oxidized sulphur of the ore. The charge in the converter has to be started by heat supplied to it, and the work then goes forward on the same lines as in the Huntington-Heberlein process, so that we may assume that the reactions are the same and come under the same explanation.

Carmichael was quick to see what was really an important part and a correct explanation of the original process. He was not misled by wrong theory about any mythical calcium peroxide, and so he obtained his patent for the use of CaSO₄ and the dispensing of the roast in a furnace.

This process would always be limited in its application by the comparative rarity of cheap supplies of gypsum, but it appears to be a great success at Broken Hill; there it is not only of importance in working the leady ores, but also for making sulphuric acid for the new treatment of mixed sulphides by the Delprat and Potter methods. For this purpose, the use of CaSO₄ will have the additional advantage that the mixture to be worked in the converter will contain not only the sulphur of the ore, but also that of the added gypsum; on decomposition, it will yield stronger gases for the lead chambers of the acid plant.

Finally comes the Savelsberg patent, which is the simplest of all; not only (like the Carmichael process) avoiding the preliminary roast with its extra plant, but also not requiring the use of ready-made CaSO₄, as it uses raw ore and limestone directly in the converter. I have no knowledge as to actual results of this process; and, so far as I am aware, nothing on the subject has been published. But Professor Borchers evidently has some information about it, and regards it as the most successful of the methods of carrying out the new ideas. On the face of it, there seems no reason why it should not attain all the results desired, as the chemical and physical actions of the CaO, and of the CaSO₄ formed from it, should come into play in the same manner and in the same order as in the original process; as it is carried out in the identical converter used by Huntington and Heberlein, the final reactions (as suggested above) will take place under the same conditions as to continuous decomposition _under considerable heat and pressure_, which I regard as the most vital part of the whole matter.

It is well to emphasize again the fact that the idea, and the means of obtaining these vital conditions, owe their origination to Huntington and Heberlein.

THEORETICAL ASPECTS OF LEAD-ORE ROASTING[22]

BY C. GUILLEMAIN

(March 10, 1906)

It is well known that the process of roasting lead ores in reverberatory furnaces proceeds in various ways according to the composition of the ore in question. Thus in roasting a sulphide lead ore rich in silica, one of the reactions is:

PbS + 3O = PbO + SO₂.

But this reaction is incomplete, for the gases which pass on in the furnace are rich in SO₂ and in SO₃. And so it is found that whatever lead oxide is formed passes over almost immediately into lead sulphate, according to the reaction:

PbO + SO₂ + O = PbSO₄.

This reaction is the chief one which takes place. Whether the silicious gangue serves as a catalyzer for the sulphur dioxide, or whether it serves merely to keep the galena open to the action of the gases, the end result of the roast is usually the formation of lead sulphate according to the above reaction.

In the case of an ore rich in galena, a slow roast is essential, for it is desired to have the following reaction take place during the latter part of the roast:

PbS + 3PbSO₄ = 4PbO + 4SO₂.

Now, if the heating were too rapid, not enough lead sulphate would be found to react with the unaltered galena. The quick roasting of a rich ore would result in the early sintering of the charge, and sintering prevents the further formation of lead sulphate. Whether this sintering (which takes place so easily and which is so harmful in the latter part of the process) is due to the low melting point of the lead sulphide, whether the heat evolved by the reaction

PbS + 3O = PbO + SO₂

is sufficient to melt the lead sulphide, or whether other thermochemical effects (notably the preliminary sulphatizing of the lead sulphide) come into play, must for the present be undecided. Suffice it to say that the sintering of the charge works against a good roast.

In the Tarnowitz process a definite amount of lead sulphide is converted into lead sulphate by a preliminary roast. The sulphate then reacts with the unaltered lead sulphide, and metallic lead is set free, thus:

PbS + PbSO₄ = 2Pb + 2SO₂.

But when a very little of the sulphide has been transformed into sulphate, and when there is so little of the latter present that only a small amount of lead sulphide can be reduced to metallic lead, the mass of ore begins to sinter and grow pasty. Very little lead could be formed were it not for the addition of crushed lime to the charge just before the sintering begins. This lime breaks up the charge and cools it, prevents any sintering, and allows the continued formation of lead sulphate.

It scarcely can be held that the lime has any chemical effect in forming lead sulphate, or in forming a hypothetical compound of lead and calcium. Even if such theories were tenable from a physico-chemical point of view, they would be lessened in importance by the fact that other substances, such as purple ore or puddle cinder, act just as well as the lime.

There are now to be mentioned several new processes of lead-ore roasting whose operations fall so far outside the common ideas on the subject that their investigation is full of interest. For a long time the attempt had been made to produce lead directly by blowing air through lead sulphide in a manner analogous to the production of bessemer steel or the converting of copper matte. In the case of the lead sulphide, the oxidation of the sulphur was to furnish the heat necessary to carry on the process.

After many attempts along this line, Antonin Germot has perfected a method wherein, by blowing air through molten galena, metallic lead is obtained.[23] About 60 per cent. of a previously melted charge of galena is sublimed as lead sulphide, and the rest remains behind as metallic lead. The disadvantages of the process are the difficulties of collecting all of the sublimate and of working it up. Moreover, it is impossible as yet to secure two products of which one is silver-free and the other silver-bearing. The silver values are in both the metallic lead and in the sublimed lead sulphide.

While the process just described answers for pure galena, it fails with ores which contain about 10 per cent. of gangue. In the case of such ores, they form a non-homogeneous mass when melted, and the blast penetrates the charge with difficulty. If the pressure is increased the air forces itself out through tubes and canals which it makes for itself, and the charge freezes around these passages.

Messrs. Huntington and Heberlein have gone a little farther. Although they are unable to obtain metallic lead directly, they prepare the ore satisfactorily for smelting in the blast furnace, after their roasting is completed. The inventors found that if lead sulphide is mixed with crushed lime, heated with access of air, and then charged into a converter and blown, the sulphur is completely removed in the form of sulphur dioxide. The charge, being divided by the lime, remains open uniformly to the passage of air, and sinters only when the sulphur is eliminated.

The inventors announce, as the theory of their process, that at 700 deg. C. the lime forms a dioxide of calcium (CaO₂) which at 500 deg. C. breaks down into lime (CaO) and nascent oxygen. This nascent oxygen oxidizes the lead sulphide to lead sulphate according to the reaction:

PbS + 4O = PbSO₄.

Furthermore it is claimed that the heat evolved by this last reaction is large enough to start and keep in operation a second reaction, namely

PbS + PbSO₄ = 2PbO + 2SO₂.

The theory, as just mentioned, cannot be accepted, and some of the reasons leading to its rejection will be given.

It is well established that the simple heating of lime with access of air will not result in further oxidation of the calcium. The dioxide of calcium cannot be formed even by heating lime to incandescence in an atmosphere of oxygen, nor by fusing lime with potassium chlorate. Moreover, calcium stands very near barium in the periodic system. And as the dioxide of barium is formed at a low temperature and breaks up on continued heating, it seems absurd to suppose that the dioxide of calcium would act in exactly the opposite manner. Moreover, a consideration of the thermo-chemical effects will disclose more inconsistencies in the ideas of the inventors. The breaking up of CaO₂ into CaO and O is accompanied by the evolution of 12 cal. The reaction of the oxygen (thus supposed to be liberated) upon the lead sulphide is strongly exothermic, giving up 195.4 cal. So much heat is produced by these two reactions that, if the ideas of the inventors were true, the further breaking up of the calcium dioxide would stop, as the whole charge would be above 500 deg. C. It appears, then, that the explanations suggested by Messrs. Huntington and Heberlein are untrue.

In the usual roasting process, as carried out in reverberatory furnaces, it is well established that the gangue, and whatever other substances are added to the ore, prevent mechanical locking up of charge particles, since they stop sintering. It is not at all improbable that in the new roasting process the chief, if not the only, part played by the lime is the same as that played by the gangue in reverberatory-furnace roasting. A few observations leading to this belief will be given.

It is known that other substances will answer just as well as lime in this new roasting process. Such substances are manganese and iron oxides. Not only these two substances, but in fact any substance which answers the purpose of diminishing the local strong evolution of heat, due to the reaction:

PbS + 3O = PbO + SO₂,

serves just as well as the lime. This fact is proved by exhaustive experiments in which mixtures of lead sulphide on the one hand, and quartz, crushed lead slags, iron slags, crushed iron ores, crushed copper slags, etc., on the other hand, were used for blowing. All these substances are such that any chemical action, analogous to the splitting up of CaO₂, or the formation of plumbates as suggested by Dr. Borchers, cannot be imagined. The time is not yet ripe, without more experiments on the subject, to assert conclusively that there is no acceleration of the process due to the formation of plumbates through the agency of lime. But the facts thus far secured point out that such reactions are, at least, not of much importance.

Theoretical considerations point out that it ought to be possible to avoid the injurious local increase of temperature during the progress of this new roasting process, without having to add any substance whatever. To explain: The first reaction taking place in the roasting is

PbS + 3O = PbO + SO₂ + 99.8 cal.

Now the heat thus liberated may be successfully dispersed if there is, in simultaneous progress, the endothermic reaction:

PbS + 3PbSO₄ = 4PbO + 4SO₂ - 187 cal.

Hence if there could be obtained a mixture of lead sulphide and of lead sulphate in the proportions demanded by the above reaction, then such a mixture ought to be blown successfully to lead oxide without the addition of any other substance. Such a process has, in fact, been carried out. The original galena is heated until the required amount of lead sulphate has been formed. Then the mixture of lead sulphide and of lead sulphate is transferred to a converter and blown successfully without the addition of any other substance.

The adaptability of an ore to the process just mentioned depends on the cost of the preliminary roast and the thoroughness with which it must be done. As is known, when lead sulphide is heated with access of air, it is very easy to form sintered incrustations of lead sulphate. If these incrustations are not broken up, or if their formation is not prevented by diligent rabbling, the further access of air to the mass is prevented and the oxidation of the charge stops. If ores with such incrustations are placed in the converter without being crushed, they remain unaltered by the blowing. If the incrustations are too numerous the converting becomes a failure.

It has been found that the adoption of mechanical roasting furnaces prevents this. Such furnaces appear to stop the frequent failures of the blowing which are due to the lack of care on the part of the workmen during the preliminary roasting. Moreover, in such mechanical furnaces a more intimate mixture of the sulphide with the sulphate is obtained, and the degree of the sulphatizing roast is more easily controlled.

As a summary of the facts connected with this new blowing process, it may be stated that the best method of working can be determined upon and adopted if one has in mind the fact that the amount of substance (lime) to be added is dependent on: 1, the amount of sulphur present; 2, the forms of oxidation of this sulphur; 3, the amount of gangue in the ore; 4, the specific heats of the gangue and of the substance added; 5, the degree of the preparatory roasting and heating.

For example, with concentrates which run high in sulphur, there is required either a large amount of additional material, or a long preliminary roast. The specific heat of the added material must be high, and the heat evolved by the oxidation of the sulphur in the preliminary roast must be dispersed. Oftentimes it is necessary to cool the charge partially with water before blowing. On the other hand, if the ore runs low in sulphur, the preliminary roast must be short, and the temperature necessary for starting the blowing reactions must be secured by heating the charge out of contact with air. Not only must no flux be added, but oftentimes some other sulphides must be supplied in order that the blowing may be carried out at all.

The opportunity for the acquisition of more knowledge on this subject is very great. It lies in the direction of seeing whether or not the strong local evolution of heat cannot be reduced by blowing with gases poor in oxygen rather than with air. Mixtures of filtered flue gases and of air can be made in almost any proportion, and such mixtures would have a marked effect upon the possibility of regulating the progress of the oxidation of the various ores and ore-mixtures which are met with in practice.

METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE[24]

BY F. O. DOELTZ

(January 27, 1906)

In his British patent,[25] for desulphurizing sulphide ores, A. D. Carmichael states that a mixture of lead sulphide and calcium sulphate reacts “at dull red heat, say about 400 deg. C.,” forming lead sulphate and calcium sulphide, according to the equation:

PbS + CaSO₄ = PbSO₄ + CaS.

Judging from thermo-chemical data, this reaction does not seem probable. According to Roberts-Austen,[26] the heats of formation (in kilogram-calories) of the different compounds in this equation are as follows: PbS = 17.8; CaSO₄ = 318.4; PbSO₄ = 216.2; CaS = 92. Hence we have the algebraic sum:

-17.8 - 318.4 + 216.2 + 92 = -28.0 cal.

As the law of maximum work does not hold, experiment only can decide whether this decomposition takes place or not. The following experiments were made:

_Experiment 1._—Coarsely crystalline and specially pure galena was ground to powder. Some gypsum was powdered, and then calcined. The powdered galena and calcined gypsum were mixed in molecular proportions (PbS + CaSO₄), and heated for 1½ hours to 400 deg. C., in a stream of carbon dioxide in a platinum resistance furnace. The temperature was measured with a Le Chatelier pyrometer. The material was allowed to cool in a current of carbon dioxide.

The mixture showed no signs of reaction. Under the magnifying glass the bright cube-faces of galena could be clearly distinguished. If any reaction had taken place, in accordance with the equation given above, no bright faces of galena would have remained.

_Experiment 2._—A similar mixture was slowly heated, also in the electric furnace, to 850 deg. C., in a stream of carbon dioxide, and was kept at this temperature for one hour.

It was observed that some galena sublimed without decomposition, being redeposited at the colder end of the porcelain boat (7 cm. long), in the form of small shining crystals. The residue was a mixture of dark particles of galena and white particles of gypsum, in which no evidence of any reaction was visible under the microscope. That galena sublimes markedly below its melting point has already been noted by Lodin.[27]

_Experiment 3._—In order to determine whether the inverse reaction takes place, for which the heat of reaction is + 28.0 cal., the following equations are given:

PbSO₄ + CaS = PbS + CaSO₄; - 216.2 - 92 + 17.8 + 318.4 = 28.

A mixture of lead sulphate and calcium sulphide was heated in a porcelain crucible in a benzine-bunsen flame (Barthel burner). The materials were supplied expressly “for scientific investigation” by the firm, C. A. F. Kahlbaum.

The white mixture turned dark and presently assumed the color which would correspond to its conversion into lead sulphide and calcium sulphate. This experiment is easy to perform.

_Experiment 4._—The same materials, lead sulphate and calcium sulphide, were mixed in molecular ratio (PbSO₄ + CaS), and were heated for 30 minutes to 400 deg. C., on a porcelain boat in the electric furnace, in a current of carbon dioxide. The mixture was allowed to cool in a stream of carbon dioxide, and was withdrawn from the furnace the next day (the experiment having been made in the evening).

The mixture showed a dark coloration, similar to that of the last experiment; but a few white particles were still recognizable. The material in the boat smelled of hydrogen sulphide.

_Experiment 5._—A mixture of pure galena and calcined gypsum, in molecular ratio (PbS + CaSO₄), was placed on a covered scorifier and introduced into the hot muffle of a petroleum furnace, at 700 to 800 deg. C. The temperature was then raised to 1100 deg. C.

From 5 g. of the mixture a dark-gray porous cake weighing 3.7g. was thus obtained. There was some undecomposed gypsum present, recognizable under the magnifying glass. No metallic lead had separated out. When hot hydrochloric acid was poured over the mixture, it evolved hydrogen sulphide. The fracture of the cake showed isolated shining spots. The supposition that it was melted or sublimed galena was confirmed by the aspect of the cake when cut with a knife; the surface showed the typical appearance of the cut surface of melted galena. On cutting, the cake was found to be brittle, with a tendency to crumble. On boiling with acetic acid, a little lead went into solution. Wetting with water did not change the color of the crushed cake.

_Experiment 6._—In his experiments for determining the melting point of galena, Lodin[28] found that, in addition to its sublimation at a comparatively low temperature, the galena also undergoes oxidation if carbon dioxide is used as the “neutral” atmosphere. Lodin was therefore compelled to use a stream of nitrogen in his determination of the melting point of galena. Now the temperature of experiment 2 (850 deg. C.), described heretofore, is not as high as the melting point of galena (which lies between 930 and 940 deg. C.); therefore experiment 2 was repeated in a stream of nitrogen, so as to insure a really neutral atmosphere. A mixture of galena and calcined gypsum in molecular ratio (PbS + CaSO₄) was heated to 850 deg. C., was kept at this temperature for one hour, and allowed to cool, the entire operation being carried out in a stream of nitrogen.

Again, galena had sublimed away from the hotter end of the porcelain boat (6.5 cm. long), and had been partially deposited in the form of small crystals of lead sulphide at the colder end. The material in the boat consisted of a mixture of particles having the dark color of galena, and others with the white color of gypsum, the original crystals of gypsum and the bright surfaces of the lead sulphide being distinctly recognizable under the magnifying glass. The loss in weight was 1.9 per cent.

_Experiment 7._—For the same reason as in 2, experiment 5 was also repeated, using a current of nitrogen. A mixture of galena and calcined gypsum, in molecular ratio (PbS + CaSO₄) was heated in a porcelain boat to 1030 deg. C., in a platinum-resistance furnace, and allowed to cool, being surrounded by a stream of nitrogen during the whole period.

Some sublimation of lead sulphide again took place. The mixture was seen to consist of white particles of gypsum, and others dark, like galena. The loss in weight was 3.5 per cent. The mixture had sintered together slightly; with hot hydrochloric acid, it evolved hydrogen sulphide. On boiling with acetic acid, a little lead (only a trace) went into solution. There was, therefore, practically no lead oxide present; no metallic lead had separated out.

_Experiment 8._—In experiment 3, lead sulphate and calcium sulphide were mixed roughly and by hand (i.e., not weighed out in molecular ratio); in this experiment such a mixture of lead sulphate and calcium sulphide in molecular ratio (PbSO₄ + CaS) was heated in a porcelain crucible in a benzine-bunsen flame. It presently turned dark, and a dark gray product was obtained, as in the former experiment.

_Experiment 9._—In a mixture of lead sulphate and sodium sulphide in molecular ratio (PbSO₄ + Na₂S), the constituents react directly on rubbing together in a porcelain mortar. The mass turns dark gray, with formation of lead sulphide and sodium sulphate.

If a similar mixture is heated, it also turns dark gray. On lixiviation with water, a solution is obtained which gives a dense white precipitate with barium chloride.

_Experiment 10._—If lead sulphate and calcium sulphide are rubbed together in a mortar, the mass turns a grayish-black.

_Conclusion._—From these experiments I infer that the reaction

PbS + CaSO₄ = PbSO₄ + CaS

does not take place, but, on the contrary, that when lead sulphate and calcium sulphide are brought together, the tendency is to form lead sulphide and calcium sulphate.

Nevertheless, on heating a mixture of galena and gypsum in contact with air, lead sulphate will be formed along with lead oxide; not, however, owing to any double decomposition of the galena with the gypsum, but rather to the formation of lead sulphate from lead oxide and sulphuric acid produced by catalysis, thus:

PbO + SO₂ + O = PbSO₄.

This is the well-known process which always takes place in roasting galena, the explanation of which was familiar to Carl Friedrich Plattner. That the presence of gypsum has any chemical influence on this process seems to be out of the question according to the above experiments.

THE HUNTINGTON-HEBERLEIN PROCESS

BY DONALD CLARK

(October 20, 1904)

The process was patented in 1897, and is based on the fact that galena can be desulphurized by mixing it with lime and blowing a current of air through the mixture. If the temperature is dull red at the start, no additional source of heat is necessary, because the reaction causes a great rise in temperature. The chemistry of the process cannot be said at present to have been worked out in detail.

The reactions given by the patentees are not satisfactory, since calcium dioxide is formed only at low temperatures and is readily decomposed on gently warming it; lead oxide, however, combines with oxygen under suitable conditions at a temperature not exceeding 450 deg. C. and forms a higher oxide, and it is probable that this unites with the lime to form calcium plumbate. The reaction between sulphides and lime when intimately mixed and heated may be put down as

CaO + PbS = CaS + PbO.

In contact with the air the calcium sulphide oxidizes to sulphite, then to sulphate, then reacts with lead oxide, giving calcium plumbate and sulphur dioxide,

CaSO₄ + PbO = CaPbO₃ + SO₂.

Further, calcium sulphate will also react with galena, giving calcium sulphide and lead sulphate; the calcium sulphide is oxidized, by air blown through, to calcium sulphate again, the ultimate reaction being

CaSO₄ + PbS + O = CaPbO₃ + SO₂.

In all cases the action is oxidizing and desulphurizing. It was found that oxides of iron and manganese will, to a certain extent, serve the same purpose as lime, and on application to complex ores, especially those containing much blende, that these may be desulphurized as well as galena. In the case of zinc sulphide the decomposition is probably due to the interaction of sulphide and sulphate.

ZnS + 3ZnSO₄ = 4ZnO + 4SO₂.

The process has now been adopted by the Broken Hill Proprietary Company at its works at Port Pirie, the Tasmanian Smelting Company, Zeehan, the Fremantle Smelting Works, West Australia, and the Sulphide Corporation’s works at Cockle Creek, New South Wales.

The operations carried on at the Tasmania Smelting Works comprise mixing pulverized limestone, galena and slag-making materials and introducing the mixture either into hand-rabbled reverberatories or mechanical furnaces with rotating hearths. After a roast, during which the materials have become well mixed and most of the limestone converted into sulphate and about half of the sulphur expelled, the granular product is run while still hot into the Huntington-Heberlein converters. These consist of inverted sheet-iron cones, hung on trunnions, the diameter being 5 ft. 6 in. and the depth 5 ft. A perforated plate or colander is placed as a diaphragm across the apex of the cone, the small conical space below serving as a wind-box into which compressed air is forced. A hood above the converter serves to carry away waste gases. As soon as the vessel is filled, air under a pressure of 17 oz. is forced through the mass, which rapidly warms up, giving off sulphur dioxide abundantly. The temperature rises and the mixture fuses, and in from two to four hours the action is complete. The sulphur is reduced from 10 to 1 per cent., and the whole mass is fritted and fused together. The converter is emptied by inverting it, when the sintered mass falls out and is broken up and sent to the smelters. There are 12 converters, of the size indicated, for the two mechanical furnaces, of 15 ft. diameter. Larger converters of the same type were erected to deal with the product from the hand-rabbled roasters.

At Cockle Creek, New South Wales, the galena concentrate is reduced to 1.5 mm., more than 60 per cent. of the material being finer; the limestone is crushed down to from 10 to 16 mesh; silica is also added, if it does not exist in the ore, so that, excluding the lead, the rest of the bases will be in such proportion as to form a slag running about 20 per cent. silica. The mixture may contain from 25 to 50 per cent. lead, and from 6 to 9 per cent. lime; if too much lime is added the final product is powdery, instead of being in a fused condition. This is given a preliminary roast in a Godfrey furnace.

The Godfrey furnace is characterized by a rotating, circular hearth and a low dome-shaped roof. Ore is fed through a hopper at the center and deflected outward by blades attached to a fixed radial arm. At each revolution the ore is turned over and moved outward, the mount of deflection of the blades, which are adjustable, and rate of rotation of the hearth, determining the output.

The hot semi-roasted ore is discharged through a slot at the circumference of the roaster. This may contain from 12 to 6.5 per cent. of sulphur, but from 6.5 to 8 per cent. is held to be the most suitable quantity for the subsequent operations. Thorough mixing is of the utmost importance, for if this is not done the mass will “volcano” in the converter; that is, channels will form in the mass through which the gases will escape, leaving lumps of untouched material alongside. The action can be started if a little red-hot ore is run into the converter and cold ore placed above it; the whole mass will become heated up, and the products will fuse, and sinter into a homogeneous mass showing none of the original ingredients. At Cockle Creek the time taken is stated to be five hours; a small air-pressure is turned on at first, and ultimately it is increased to 20 oz.

Operations at Port Pirie are conducted on a much larger scale. A mixture of pulverized galena, powdery limestone, ironstone and sand is fed into Ropp furnaces, of which there are five, by means of a fluted roll placed at the base of a hopper. Each roaster deals with 100 tons of the mixture in 24 hours. About 50 per cent. of the sulphur is eliminated from the ore by the Ropps (the galena in this case being admixed with a large amount of blende, there being only 55 per cent. of lead and 10 per cent. of zinc in the concentrate produced at the Proprietary mine). The hot ore from the roasters is trucked to the converters, there being 17 of these ranged in line. The converters here are large segmental cast-iron pots hung on trunnions; each is about 8 ft. diameter and 6 ft. deep, and holds an 8-ton charge. At about two feet from the bottom an annular perforated plate fits horizontally; a shallow frustrum of a cone, also perforated, rests on this; while a plate with a few perforations closes the top of the frustrum. The whole serves as a wind-box. A conical hood with flanged edges rests on the flanged edges of the converter, giving a close joint. This hood is provided with doors which allow the charge to be barred if necessary. A pipe about 1 ft. 9 in. diameter, fitted with a telescopic sliding arrangement, allows for the raising or lowering of the hood by block and tackle, and thus enables the converter to be tilted up and its products emptied. The cast-iron pots stand very well; they crack sometimes, but they can be patched up with an iron strap and rivets. Only two pots have been lost in 18 months.

Air enters at a pressure of about 24 oz. and the time taken for conversion is about four hours. The sulphur contents are reduced to about three per cent. It is found that the top of the charge is not so well converted as the interior. There is practically no loss of lead or silver due to volatilization and very little due to escape of zinc. It has also been found that practically all the limestone fed into the Ropp is converted into calcium sulphate; also that a considerable portion of lead becomes sulphate, and it is considered that lead sulphate is as necessary for the process as galena.

The value of the process may be judged from the fact that better work is now done with 8 blast furnaces than was done with 13 before the process was adopted. In addition to the sintered product from the Huntington-Heberlein pots, sintered slime, obtained by heap roasting, and flux consisting of limestone and ironstone, are fed into the furnaces, which take 2000 long tons per day of ore, fluxes and fuel. The slags now being produced average: SiO₂, 25 to 26 per cent.; FeO, 1 to 3 per cent.; MnO, 5 to 5.5; CaO, 15.5 to 17; ZnO, 13; Al₂O₃, 6.5; S, 3 to 4; Pb, by wet assay, 1.2 to 1.5 per cent.; and Ag, 0.7 oz. per ton. Although this comparatively large quantity of sulphur remains, yet no matte is formed.

THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHÜTTE[29]

BY A. BIERNBAUM

(September 2, 1905)

Nothing, for some time past, has caused such a stir in the metallurgical treatment of lead ores, and produced such radical changes at many lead smelting works, as the introduction of the Huntington-Heberlein process. This process (which it may be remarked, incidentally, has given rise to the invention of several similar processes) represents an important advance in lead smelting, and, now that it has been in use for some time at the Friedrichshütte, near Tarnowitz, in Upper Silesia, and has there undergone further improvement in several respects, a comparison of this process with the earlier roasting process is of interest.

At the above-mentioned works, up to 1900 the lead ore was treated exclusively (1) by smelting in reverberatory furnaces (Tarnowitzeröfen), and (2) by roasting in reverberatory-sintering furnaces roasted material in the shaft furnace. The factor which determined whether the treatment was to be effected in the reverberatory-smelting or in the roasting-sintering furnace was the percentage of lead and zinc in the ores; those comparatively rich in lead and poor in zinc being worked up in the former, with partial production of pig-lead; while those poorer in lead and richer in zinc were treated in the latter. About two-fifths of the lead ores annually worked up were charged into the reverberatory-smelting furnaces, and three-fifths into the sintering furnaces.

In 1900 there were available 10 reverberatory-smelting and nine sintering furnaces. These were worked exclusively by hand.

The sintered product of the roasting furnaces, and the gray slag from the reverberatory-smelting furnaces, were transferred to the shaft furnaces for further treatment, and were therein smelted together with the requisite fluxes. Eight such furnaces (8 m. high, and 1.4 m., 1.6 m., and 1.8 m. respectively in diameter at the tuyeres), partly with three and partly with five or eight tuyeres, were at that time in use.

Now that the Huntington-Heberlein process has been completely installed, the reverberatory-smelting furnaces have been shut down entirely, and the sintering furnaces also for the most part; all kinds of lead ore, with a single exception, are worked up by the Huntington-Heberlein process, irrespective of the contents of lead and zinc. An exceedingly small proportion of the ore treated, viz., the low-grade concentrate (Herdschlieche) containing 25 to 35 per cent. Pb, is still roasted in the old sintering furnace, together with various between-products (such as dust, fume, scaffoldings, and matte); these are scorified by the aid of the high percentage of silica in the material.

For roasting lead ores at the present time there are six round mechanical roasters of 6 m. diameter, one of 8 m. diameter, and two ordinary, stationary Huntington-Heberlein furnaces. The latter (which represent the primitive Huntington-Heberlein furnaces, requiring manual labor) have recently been shut down, and will probably never be used again. In the mechanical Huntington-Heberlein furnace, roasting of lead ore is carried only to such a point that a small portion of the lead sulphide is converted into sulphate. The desulphurization of the ore is completed in the so-called converter (made of iron, pear-shaped or hemispherical in form) in which the charge, up to this stage loosely mixed, is blown to a solid mass.

Owing to the ready fusibility of this product (which still contains, as a rule, up to 1.5 per cent. sulphur as sulphide), it is possible to use shaft furnaces of rather large dimensions; therefore a round shaft furnace (2.4 m. diameter at the tuyeres, 7 m. high, and furnished with 15 tuyeres) was built. In this furnace nearly the whole of the roasted ore from the Huntington-Heberlein converters is now smelted, some of the smaller shaft furnaces being used occasionally. The introduction of the new process has caused no noteworthy change in the subsequent treatment of the work-lead.

In the following study I shall discuss the treatment of a given annual quantity of ore (50,000 tons), which is the actual figure at the Friedrichshütte at the present time.

1. _Roasting Furnaces._—A reverberatory-smelting furnace used to treat 5 tons of ore in 24 hours; a roasting-sintering furnace, 8 tons. Assuming the ratios previously stated, the annual treatment by the former process would be 20,000 tons, and by the latter 30,000 tons. On the basis of 300 working days per year, and no prolonged stoppages for furnace repairs (though considering the high temperatures of these furnaces this record would hardly be expected), there would be required:

20,000 ÷ (5 × 300) = 13.3 (or 13 to 14 reverberatory furnaces). 30,000 ÷ (8 × 300) = 12.5 (or 12 to 13 sintering furnaces).

The capacity of a stationary Huntington-Heberlein furnace is 18 tons; hence in order to treat the same quantity of ores there would be required:

50,000 ÷ (18 × 300) = 9.3 (or 9 to 10 Huntington-Heberlein furnaces).

With the revolving-hearth roasters (of 6 m. diameter) working a total charge of at least 27 tons of ore, there would be required:

50,000 ÷ (27 × 300) = 6.1 (or 6 to 7 roasters).

Still better results are obtained with the 8 m. round roaster, which has been in operation for some time; in this, 55 tons of ore can be roasted daily. Three such furnaces would therefore suffice for working up the whole of the ore charged per annum.

Now, making due provision for reserve furnaces, to work up 50,000 tons of ore would require:

Reverberatory (15) and sintering furnaces (15) 30 Stationary Huntington-Heberlein furnaces 12 6 m. revolving-hearth furnaces 8 8 m. revolving-hearth furnaces 4

Similar relations hold good regarding the number of workmen attending the furnaces, there being required, daily, six men for the reverberatory furnace; eight men for the sintering furnace; ten men for the stationary; and six men for the mechanical Huntington-Heberlein furnace; or, for 14 reverberatory furnaces, daily, 84 men; for sintering furnaces, daily, 104 men; total, 188 men. While for 10 stationary Huntington-Heberlein furnaces, 100 men are required; and for 7 mechanical Huntington-Heberlein furnaces, daily, 42 men. It is expected that only 14 men (working in two shifts) will be required to run the new installation with 8 m. round roasters.

It is true that the exclusion of human labor here has been carried to an extreme. The roasters and converters will be charged exclusively by mechanical means; thus every contact of the workmen with the lead-containing material is avoided until the treatment of the roasted material in the converters is completed.

From the data given above, the capacity of each individual workman is readily determined, as follows: With the reverberatory-smelting furnace, each man daily works up 0.83 tons; with the sintering furnace, 1 ton; with the stationary Huntington-Heberlein furnace, 1.8 tons; with the 6 m. revolving-hearth furnace, 4.5 tons; and with the 8 m. revolving-hearth furnace, 11.8 tons.

A significant change has also taken place in coal consumption. Thus, when working with the reverberatory and sintering furnaces in order to attain the requisite temperature of 1000 deg. C., there was required not only a comparatively high-grade coal, but also a large quantity of it. A reverberatory furnace consumed about 503 kg., a sintering furnace about 287 kg., of coal per ton of ore. For roasting the ore in the stationary and also in the mechanical Huntington-Heberlein furnaces, a lower temperature (at most 700 deg. C.) is sufficient, as the roasting proper of the ore is effected in the converters, and the sulphur furnishes the actual fuel. For this reason, the consumption of coal is much lower. The comparative figures per ton of ore are as follows: In the reverberatory furnace, 50.3 per cent.; in the sintering furnace, 28.7 per cent.; in the stationary Huntington-Heberlein furnace, 10.3 per cent.; and in the Huntington-Heberlein revolving-hearth furnace, 7.3 per cent.

But there is another technical advantage of the Huntington-Heberlein process which should be mentioned. It is well known that the volatilization of lead at high temperatures is an exceedingly troublesome factor in the running of a lead-smelting plant; the recovery of the valuable fume is difficult, and requires condensing apparatus, to say nothing of the unhealthful character of the volatile lead compounds. This volatilization is of course particularly marked at the high temperatures employed when working with reverberatory-smelting furnaces; the same is true, in a somewhat less degree, of the sintering furnaces. In consequence of the markedly lower temperature to which the charge is heated in the Huntington-Heberlein furnace, and also of the peculiar mode of completing the roast in blast-converters, the production of fume is so reduced that the difference between the values recovered in the old and the new processes is very striking. Whereas, in 1900, in working up 12,922 tons of ore in the reverberatory-smelting furnace, and 14,497 tons in the sintering furnace (27,419 tons in all), there was recovered 2470 tons (or 9 per cent.) as fume from the condensers and smoke flues, the quantity of fume recovered, in 1903, fell to 879 tons (or 1.8 per cent.), out of the 48,208 tons of ore roasted, and this notwithstanding the fact that in the meantime fume-condensing appliances had been considerably expanded and improved, whereby the collection was much more efficient.

Lastly, the zinc content of the ores no longer exerts the same unfavorable influence as in the old process (wherein it was advisable to subject ore containing much blende to a final washing before proceeding to the actual metallurgical treatment). In the new process, the ores are simply roasted without regard to their zinc content. In this connection it has been found that a considerable proportion of the zinc passes off with the fume, and that the roasted material usually contains a quantity of zinc so small that it no longer causes any trouble in the shaft furnace. It may also be mentioned here that the ore-dressing plants recently installed in the mines of Upper Silesia have resulted in a more perfect separation of the blende.

_Shaft Furnaces._—The finished product from the Huntington-Heberlein blast-converters is of a porous character, and already contains a part of the flux materials (such as limestone, silica and iron) which are required for the shaft-furnace charge. It is just these two characteristics of the roasted product (its porous nature, on the one hand, leading to its more perfect reduction by the furnace gases; and, on the other hand, the admixture of fluxes in the molten condition, resulting in a more complete utilization of the temperature), which, together with its higher lead and lower zinc content, determine its ready fusibility. If we further consider that it is possible in the new process to make the total charge of the shaft furnace richer in lead than formerly (two-thirds of the total charge as against one-third), and that a higher blast pressure can be used without danger, it follows immediately that the capacity of a shaft furnace is much greater by the new process than by the old method of working. The daily production of the shaft furnaces on the old and the new process is as shown in the table given herewith:

─────────────┬─────────────────────────┬─────────┬──────────────────── │ │ CHARGE │ WORK-LEAD TYPE OF SHAFT│ CHARACTER OF CHARGE │ PER DAY,│ PRODUCED FURNACE │ │ TONS │ PER DAY, TONS ─────────────┼─────────────────────────┼─────────┼──────────────────── 3 tuyeres │{ Gray slag from } │ 36 │ 6 to 7 } │{ reverberatory } │ │ } │{ furnaces and } │ │ } Low- │{ sintered concentrate } │ │ }pressure │ │ │ } Blast 8 tuyeres │ ” ” │ 36 to 38│ 6 to 8 } │ │ │ } 3 tuyeres │{ Roasted product of } │ 36 │ 11 to 12 } │{ Huntington-Heberlein } │ │ │{ process } │ │ │ │ │ 8 tuyeres │ ” ” │ 65 to 72│ 24 to 26 } High- │ │ │ }pressure 15 tuyeres │ ” ” │ 270 │ 90 to 100 } Blast ─────────────┴─────────────────────────┴─────────┴────────────────────

It should be noted that the figure given for the furnace with 15 tuyeres represents the average for 1904; this average is lowered by the circumstance that during this period there was frequently a deficiency of roasted material, and the furnace had to work with low-pressure blast. A truer impression can be gained from the month of March, 1905, for instance, during which time this furnace worked under normal conditions; the results are as follows:

The average for March, 1905, was: Ore charged, 8,269.715 tons; coke, 652.441 tons; total, 8,922.156 tons. Or, in 24 hours: Ore charged, 266.765 tons; coke, 21.046 tons; total, 287.811 tons. The production of work-lead was 3,133.245 tons, or 101.069 tons per day.

The maximum production of roasted ore was 210 tons, on June 30, 1905, when the total charge was: Ore, 327.38 tons; coke, 25.2 tons; total, 352.58 tons. The quantity of work-lead produced on that day was 120.695 tons, while the largest quantity previously produced in one day was 124.86 tons. It should also be mentioned that the lead tenor of the slag is almost invariably below 1 per cent.; it usually lies between 0.3 and 0.5 per cent.

As in the case of the roasting furnaces, the productive capacity of the shaft furnace also comes out clearly if we figure the number of furnaces required, on the basis of an annual consumption of 50,000 tons of ore. If we consider 1 ton of the roasted material as equivalent to 1 ton of ore (which is about right in the case of the Huntington-Heberlein material, but is rather a high estimate in the case of the product of the sintering furnace), then, in the old process (where one-third of the charge was lead-bearing material), 12 tons could be smelted daily. There would therefore be needed at least:

50,000 ÷ (12 × 300) = 14 three-tuyere shaft furnaces.

Since, as already mentioned, the lead-bearing part of the charge constitutes two-thirds of the whole in the Huntington-Heberlein process, the number of shaft furnaces of different types, as compared with the foregoing, would figure out:

3-tuyere shaft furnace, with product of sintering furnace, 50,000 ÷ (12 × 300) = 14 furnaces;

3-tuyere shaft furnace, with product of Huntington-Heberlein furnace, 50,000 ÷ (24 × 300) = 7 furnaces;

8-tuyere shaft furnace, with product of Huntington-Heberlein furnace, 50,000 ÷ (48 × 300) = 3.4 (say 4) furnaces;

15-tuyere shaft furnace, with product of Huntington-Heberlein furnace, 50,000 ÷ (180 × 300) = 1 furnace.

Running regularly and without interruption, the large shaft furnace is therefore fully capable of coping with the Huntington-Heberlein roasted material at the present rate of production.

As regards the number of workmen and the product turned out per man, no such marked difference is produced by the introduction of the Huntington-Heberlein process in the case of the shaft furnace as there was noted for the roasting operation. This is chiefly due to the fact that the work which requires the more power (such as charging of the furnaces, conveying away the slag and pouring the lead) can be executed only in part by mechanical means. Nevertheless, it will be seen from the table given herewith that, on the one hand, the number of men required for the charge worked up is smaller; and, on the other, the product turned out per man has risen somewhat.

─────────┬─────────┬────────┬──────────┬────────┬─────────────┬─────── TYPE OF │CHARACTER│ CHARGE │NUMBER OF │ CHARGE │DAILY OUTPUT │OUTPUT SHAFT │OF CHARGE│PER DAY,│FURNACEMEN│PER MAN,│OF WORK-LEAD,│PER MAN, FURNACE │ │ TONS │ │ TONS │ TONS │ TONS ─────────┼─────────┼────────┼──────────┼────────┼─────────────┼─────── 3 tuyere│ A │ 36 │ 6 │ 6.0 │ 6 │ 1.0 8 tuyere│ B │ 38 │ 6 │ 6.3 │ 8 │ 1.3 3 tuyere│ C │ 36 │ 6 │ 6.0 │ 12 │ 2.0 8 tuyere│ D │ 72 │ 12 │ 6.0 │ 26 │ 2.1 15 tuyere│ E │ 270 │ 34 │ 7.9 │ 90 │ 2.6 ─────────┴─────────┴────────┴──────────┴────────┴─────────────┴───────

┌──────────┬──────────────────────────────────────┐ │ CHARACTER│ CHARACTER OF CHARGE │ │ OF CHARGE│ │ │ CODE │ │ ├──────────┼──────────────────────────────────────┤ │ A │ Sintered concentrate and gray slag │ │ B │ from reverberatory furnace. │ │ B │ Gray slag from reverberatory furnace.│ │ C │ Huntington-Heberlein product. │ │ D │ Huntington-Heberlein product. │ │ E │ Huntington-Heberlein product. │ └──────────┴──────────────────────────────────────┘

A slight difference only is produced by the new process in the consumption of coke; the economy is a little over 1 per cent., the coke consumed being reduced from 9.39 per cent. to 8.17 per cent. of the total charge. But with the high price of coke, even this small difference represents a considerable lowering of the cost of production.

With the great increase in the blast pressure, it would be supposed that the losses in fume would be much greater than with the former method of working. But this is not the case; on the contrary, all experience so far shows that there is much less fume developed. In 1904, for instance, the shaft-furnace fume recovered in the condensing system amounted to only 1.06 per cent. of the roasted material, or 0.64 per cent. of the total charge, as against 2.03 and 1.0 per cent., respectively, in former years. The observations made on the quantity of flue dust carried away with the gases escaping into the air through the stack showed that it is almost nil.

Now, from the loss in fume being slight, from the tenor of lead in the slag being low, and, on the one hand, from the quantity of lead-matte produced being much less than before, while on the other the losses in roasting the ore are greatly reduced—from all these considerations, it is clear that the total yield must have been much improved. As a matter of fact, the yield of lead and silver has been increased by at least 6 to 8 per cent.

_Economic Results._—As regards the economical value of the new process, for obvious reasons no data can be furnished of the exact expenditure, i.e., the actual total cost of roasting and smelting the ore. But this at least is placed beyond doubt by what has been developed above, namely, that considerable saving must be effected in the roasting, and especially in the smelting, as compared with the former mode of working. If we take into account only the economy which is gained in wages through the increase in the material which one workman can handle, and that resulting from the reduced consumption of coal and coke, these alone will show sufficiently that an important diminution of working cost has taken place. The objection which might be raised, that the saving effected by reducing manual labor may be neutralized by the expense of mechanical power (actuating the roasters, furnishing the compressed blast, etc.), cannot be regarded as justified, as the cost of mechanical work is comparatively low. Thus, for instance, the large 8 m. furnace and the small, round furnaces require 15 h.p. if worked by electricity. According to an exact calculation, the cost (to the producer) of the h.p. hour, inclusive of machinery, figures out to 3.6 pfennigs (0.9c.); hence the daily expense for running the revolving-hearth furnaces amounts to: 15 × 3.6 pfg. × 24 = 12.96 marks ($3.42). As the seven furnaces together work up: (6 × 27) + 55 = 217 tons of ore, the cost per ton of ore is about 0.06 mark (1.5c.).

The requisite blast is produced by means of single-compression Encke blowers, of which one is quite sufficient when running at full load, and then consumes 34 h.p. The daily expenses are accordingly: 34 × 3.6 pfg. × 24 = 29.28 marks ($7.32); or per ton of ore, 29.28 ÷ 217 = 0.14 mark (3.5c.). Therefore the total expense for the mechanical work in roasting the ore amounts to 0.06 + 0.14 = 0.20 mark (5c.).

However, the cost of roasting is much more affected by the expense for keeping the furnaces in repair; another important factor is the acquisition and maintenance of the tools. Both in the case of the sintering and also the reverberatory-smelting furnace, the cost of keeping in repair was high; the consumption of iron was especially large, owing to the rapid wear of the tools. This was not surprising, considering that a notably higher temperature prevailed in the reverberatory and sintering furnaces than in the new roasters, in which the temperature strictly ought not to rise above 700 deg. C. But in the old type of furnace the high temperature and the constant working with the iron tools caused their rapid wear, thus creating a large item for iron and steel and smith work. In the new process (and more especially in the revolving-hearth roasters) this disadvantage does not arise. In this case there is practically no work on the furnace, and the wear and tear of iron is small. Also, the cost of keeping the furnaces in repair when working regularly is small as compared with the old process. In the year 1900, for instance, the cost of maintenance and tools for the reverberatory and sintering furnaces came to 20,701.93 marks ($5,175.48) for treating 27,419.75 tons of ore. Per ton of ore, this represents 0.75 mark (19c.). In the year 1903, on the other hand, only 9,074.17 marks ($2,268.54) were expended, although 48,208 tons of ore were worked up in the three stationary and six mechanical Huntington-Heberlein furnaces. The cost of maintenance was, therefore, in this case 0.18 mark (4.5c.) per ton of ore.

In the cost of smelting in the shaft furnace, only a slight difference in favor of the Huntington-Heberlein process is found if the estimate is based on the total charge; but a marked difference is shown if it is referred to the lead-bearing portion of the charge, or to the work-lead produced. Thus the cost of maintenance and total cost of smelting, figured for one ton of ore, without taking into account general expenses, have been tabulated as follows:

────────────────────────────┬──────────────────────────────── │REDUCTION IN EXPENSES PER TON OF ├────────┬──────────┬──────────── │ TOTAL │ LEAD ORE │ WORK-LEAD │ CHARGE │ │ ────────────────────────────┼────────┼──────────┼──────────── (_a_) Cost of maintenance │ 0.01M │ 0.38M │ 0.67M │(0.25c) │ (9.5c) │ (16.75c) │ │ │ (_b_) Total cost of smelting│ 0.20M │ 6.46M │ 11.48M │ (5c) │ ($1.615) │ ($2.87) ────────────────────────────┴────────┴──────────┴────────────

The marked reduction in the expenses, as referred to the lead-ore and the work-lead produced, is determined (as was pointed out above) by the greater lead content of the charge, and by the larger yield of lead consequent thereon. The advantage of longer smelting campaigns (which ultimately were mostly prolonged to one year) also makes itself felt; it would be still more marked, if the shaft furnace (which was still in working condition after it was blown out) had been run on for some time longer.

Finally, if we examine the question of the space taken up by the plant (which, owing to the scarcity of suitably located building sites, would have been important at the Friedrichshütte at the time when the quantity of ore treated was suddenly doubled), here again we shall recognize the great advantage which this establishment has gained from the Huntington-Heberlein process.

As was calculated above, there would have been required 15 reverberatory and 15 sintering furnaces to cope with the quantity of ore treated. As a reverberatory requires, in round numbers, 120 sq. m. (1290 sq. ft.), and a sintering furnace 200 sq. m. (2153 sq. ft.); and as fully 100 sq. m. (1080 sq. ft.) must be allowed for each furnace for a dumping ground, therefore the 15 reverberatory furnaces would have required an area of 15 × 120 + 15 × 100 = 3300 sq. m.; the 15 sintering furnaces would have required 15 × 200 + 15 × 100 = 4500 sq. m.; in all 3300 + 4500 = 7800 sq. m. (83,960 sq. ft.). The 12 stationary Huntington-Heberlein furnaces (built together two and two) would take up a space of 6 × 200 + 12 × 100 = 2400 sq. m. (25,830 sq. ft.). Similarly, 8 small furnaces would require 8 × 100 + 8 × 100 = 1600 sq. m. (17,222 sq. ft.); while for the new installation of four 8-meter revolving-hearth furnaces and 10 large converters, only 1320 sq. m. (14,120 sq. ft.) have been allowed.

For shaft furnaces with three or eight tuyeres, which were run with low-pressure blast for the material roasted on the old plan, the total area built upon was 18 × 16.5 = 297 sq. m.; while a further area of 18 × 14 = 250 sq. m. was hitherto provided, and was found sufficient for dumping slag when working regularly. Therefore, the installation of shaft furnaces formerly in existence, after requisite enlargement to 14 furnaces, would have demanded a space of 7 × 297 + 7 × 250 = 3829 sq. m. (42,215 sq. ft.). If four of the small shaft furnaces had been reconstructed for eight tuyeres, and run with Huntington-Heberlein roasted material, using high-pressure blast, the area occupied would have been reduced to 2 × 297 + 2 × 250 sq. m. = 1094 sq. m. (11,776 sq. ft.).

Still more favorable are the conditions of area required in the case of the large shaft furnace. This furnace stands in a building covering an area of 350 sq. m. (3767 sq. ft.), which is more than sufficient room. The slag-yard (situated in front of this building, and amply large enough for 36 hours’ run) has an area of 250 sq. m. (2691 sq. ft.); thus the space occupied by the large shaft furnace, including a yard of 170 sq. m. (1830 sq. ft.), is in all 780 sq. m. (8396 sq. ft.).

After completion of the new roasting plant and the large shaft furnace in connection with it, there would be occupied 1320 + 780 = 2100 sq. m. (2260 sq. ft.); and if the system of reverberatory and sintering furnaces had been continued (with the requisite additions thereto and to the old shaft-furnace system), there would have been required 11,629 sq. m. (125,214 sq. ft.). In the estimate above given no regard has been paid to any of the auxiliary installations (dust chambers, etc.), which, just as in the case of the old process, would have had to be provided on a large scale.

It is of course self-evident that both the principal and the auxiliary installations in the old process would not only have involved a high first cost, but would also, on account of their extensive dimensions, have caused considerably greater annual expense for maintenance.

THE HUNTINGTON-HEBERLEIN PROCESS FROM THE HYGIENIC STANDPOINT[30]

BY A. BIERNBAUM

(October 14, 1905)

With regard to the hygienic improvements which the Huntington-Heberlein process offers, we must first deal with the questions: What were the sources of danger in the old process, and in what way are these now diminished or eliminated? The only danger which enters into consideration is lead-poisoning, other influences detrimental to health being the same in one process as the other.

With the reverberatory-smelting and roasting-sintering furnaces, the chief danger of lead-poisoning lies in the metallic vapor evolved during the withdrawal of the roasted charge from the furnace. It is true that appliances may be provided, by which these vapors are drawn off or led back into the furnace during this operation; but, even working with utmost care, it is impossible to insure the complete elimination of lead fumes, especially in wheeling away the pots filled with the red-hot sintered product. Moreover, the work at the reverberatory-smelting and roasting-sintering furnaces involves great physical exertion, wherefore the respiratory organs of the workmen are stimulated to full activity, while the exposure to the intense heat causes the men to perspire freely. Hence, as has been established medically, the absorption of the poisonous metallic compounds (which are partially soluble in the perspiration) into the system is favored both by inhalation of the lead vapor and by its penetration into the pores of the skin, opened by the perspiration.

A further danger of lead-poisoning was occasioned by the frequently recurring work of clearing out the dust flues. The smoke from the reverberatory-smelting furnace especially contained oxidized lead compounds, which on absorption into the human body might readily be dissolved by the acids of the stomach, and thus endanger the health of the workmen.

In the Huntington-Heberlein furnaces, on the other hand, although the charge is raked forward and turned over by hand, it is not withdrawn, as in the old furnaces, by an opening situated next to the fire, but is emptied at a point opposite into the converters which are placed in front of the furnace. Moreover, the converters are filled with the charge at a much lower temperature. Inasmuch as this charge has already cooled down considerably, there can be practically no volatilization of lead. The small quantity of gas which may nevertheless be evolved is drawn off by fans through hoods placed above the converters.

A further improvement, from the hygienic point of view, is in the use of the mechanical furnaces, from which the converters can be filled automatically (almost without manual labor, and with absolute exclusion of smoke). The converters are then placed on their stands and blown. This work also is carried out under hoods, as gas-tight as possible, furnished with a few closable working apertures. During the blowing of the material, the work of the attendant consists solely in keeping up the charge by adding more cold material and filling any holes that may be formed. It does not entail nearly as much physical strain as the handling of the heavy iron tools and the continued exposure of the workmen to the hottest part of the furnace, which the former roasting process involved.

Some experiments carried out with larger converters (of 4 and 10 ton capacity) have indicated the direction in which the advantages mentioned above may probably be developed to such a point that the danger of lead-poisoning need hardly enter into consideration. Both the charging of the revolving-hearth furnaces and the filling of the converters are to be effected mechanically. Furthermore, in the case of the large converters the filling up of holes becomes unnecessary, and no manual work of any kind is required during the whole time of blowing. The converters can be so perfectly enclosed in hoods that the escape of gases into the working-rooms becomes impossible, and lead-poisoning of the men can occur only under quite unusual circumstances.

The beneficial influence on the health of the workmen attending on the roasting furnaces, occasioned by the introduction of the Huntington-Heberlein process, can be seen from the statistics of sickness from lead-poisoning for the years 1902 to 1904, as given herewith:

─────────┬──────┬──────┬──────────────────────────────┬─────────────── │ │ │ LEAD-POISONING │ CASES │ │ ├─────────────┬────────────────┤ CONTRACTED │ │ │NO. OF CASES │DAYS OF SICKNESS│AT REVER.│ AT ─────────┼──────┼──────┼─────┬───────┼───────┬────────┤ AND │H. H. SICKNESS METHOD OF│ YEAR │NO. OF│TOTAL│PER 100│ TOTAL │PER 100 │ SINT. │ FUR. WORKING │ │ MEN │ │PERSONS│ │PERSONS │ FUR. │ ─────────┼──────┼──────┼─────┼───────┼───────┼────────┼─────────┼───── │ │ │ │ │ │ │ │ Old │{ 1902│ 93 │ 15 │ 16.1 │ 246 │ 264.5 │ 11 │ 4 │{ 1903│ 86 │ 12 │ 13.9 │ 222 │ 258.1 │ 7 │ 5 │ │ │ │ │ │ │ │ H.-H. │ 1904│ 87 │ 8 │ 9.2 │ 242 │ 278.2 │ 6 │ 2 ─────────┴──────┴──────┴─────┴───────┴───────┴────────┴─────────┴─────

This shows a gratifying decrease in the number of cases, namely, from 16.1 to 9.2 per cent.; this decrease would have been still greater if Huntington-Heberlein furnaces had been in use exclusively. However, most of the time two or three sintering furnaces were fired for working up by-products, 16 to 18 men being engaged on that work. The Huntington-Heberlein furnaces alone (at which, in the year 1904, 69 men in all were occupied) show only 2.9 per cent. of cases. That the number of days of illness was not reduced is due to the fact that the cases among the gang of men working at the sintering furnaces were mostly of long standing and took some time to cure.

The noxious effects upon the health of the workmen in running the shaft furnaces are due to the fumes from the products made in this operation, such as work-lead, matte and slag, which flow out of the furnace at a temperature far above their melting points. Even with the old method of running the shaft furnaces the endeavor has always been to provide as efficiently as possible against the danger caused by this volatilization, and, wherever feasible, to install safety appliances to prevent the escape of lead vapors into the work-rooms; but these measures could not be made as thorough as in the case of the Huntington-Heberlein process.

The principal work in running the shaft furnaces, aside from the charging, consists in tapping the slag and pouring out the work-lead. Other unpleasant jobs are the barring down (which in the old process had to be done frequently) and the cleaning out of the furnace after blowing out.

In the old process the slag formed in the furnace flows out continuously through the tap-hole into iron pots placed in front of the spout. A number of such pots are so arranged on a revolving table that as soon as one is filled the next empty can be brought up to the duct; thus the slag first poured in has time to cease fuming and to solidify before it is removed. The vapors arising from the slag as it flows out are conveyed away through hoods. At the same time with the slag, lead matte also issues from the furnace. Now the greater the quantity of lead matte, the more smoke is also produced; and, with the comparatively high proportion of lead matte resulting from the old process, the quantity of smoke was so great that the ventilation appliances were no longer sufficient to cope with it, thus allowing vapors to escape into the work-room.

The work-lead collects at the back of the furnace in a well, from which it is from time to time ladled into molds placed near by. If the lead is allowed to cool sufficiently in the well, it does not fume much in the ladling out. But when the furnace runs very hot (which sometimes happens), the lead also is hotter and is more inclined to volatilize. In this event the danger of lead-poisoning is very great, for the workman has to stand near the lead sump.

A still greater danger attends the work of barring down and cleaning out the furnace. The barring down serves the purpose of loosening the charge in the zone of fusion; at the same time it removes any crusts formed on the sides of the furnace, or obstructions stopping up the tuyeres. With the old furnaces, and their strong tendency to crust, this work had to be undertaken almost every day, the men being compelled to work for rather a long time and often very laboriously with the heavy iron tools in the immediate neighborhood of the glowing charge, the front of the furnace being torn open for this purpose. In this operation they were exposed without protection to the metallic vapors issuing from the furnace, inasmuch as the ventilating appliances had to be partially removed during this time, in order to render it at all possible to do the work.

In a similar manner, but only at the time of shutting down a shaft furnace, the cleaning out (that is to say, the withdrawing of no longer fused but still red-hot portions of the charge left in the furnace) is carried out. In this process, however, the glowing material brought out could be quenched with cold water to such a point that the evolution of metallic vapors could be largely avoided.

Lastly, the mode of charging of the shaft furnace is also to be regarded as a cause of poisoning, inasmuch as it is impossible to avoid entirely the raising of dust in the repeated act of dumping and turning over the materials for smelting, in preparing the mix, and in subsequently charging the furnace.

By the introduction of the Huntington-Heberlein process, all these disadvantages, both in the roasting operation and in running the shaft furnaces, are in part removed altogether, in part reduced to such a degree that the danger of injury is brought to a minimum.

In furnaces in which the product of the Huntington-Heberlein roast is smelted, the slag is tapped only periodically at considerable intervals; and, as there is less lead matte produced than formerly, the quantity of smoke is never so great that the ventilating fan cannot easily take care of it. There is therefore little chance of any smoke escaping into the working-room.

As the production of work-lead, especially in the case of the large shaft furnace, is very considerable, so that the lead continually flows out in a big stream into the well, the hand ladling has to be abandoned. Therefore the lead is conducted to a large reservoir standing near the sump, and is there allowed to cool below its volatilizing temperature. As soon as this tank is full, the lead is tapped off and (by the aid of a swinging gutter) is cast into molds ready for this purpose. Both the sump and the reservoir-tank are placed under a fume-hood. The swinging gutter is covered with sheet-iron lids while tapping, so that any lead volatilized is conveyed by the gutter itself to a hood attached to the reservoir; thus the escape of metallic vapors into the working space is avoided, as far as possible.

This method of pouring does not entail the same bodily exertion as the ladling of the lead; moreover, as it requires but little time, it gives the workmen frequent opportunity to rest.

But one of the chief advantages of the Huntington-Heberlein process lies in the entire omission of the barring down. If the running of the shaft furnace is conducted with any degree of care, disorders in the working of the furnace do not occur, and one can rely on a perfectly regular course of the smelting process day after day. No formation of any crusts interfering with the operation of the furnace has been recorded during any of the campaigns, which have, in each case, lasted nearly a year.

As regards the cleaning out of the furnace, this cannot be avoided on blowing out the Huntington-Heberlein shaft furnace; but at most it occurs only once a year, and can be done with less danger to the workmen, owing to the better equipment.

Further, the charge is thrown straight into the furnace (in the case of the large shaft furnace); thus the repeated turning over of the smelting material, as formerly practised, becomes unnecessary, and the deleterious influence of the unavoidable formation of dust is much diminished.

The accompanying statistics of sickness due to lead-poisoning in connection with the operation of the shaft furnace (referring to the same period of time as those given above for the roasting furnaces) confirm the above statements.

────┬──────────┬──────────────────────────────────────────── │ │ LEAD-POISONING—SHAFT FURNACES │ ├─────────────────────┬────────────────────── YEAR│NO. OF MEN│ CASES │ DAYS OF ILLNESS │ ├─────┬───────────────┼─────┬──────────────── │ │TOTAL│PER 100 PERSONS│TOTAL│PER 100 PERSONS ────┼──────────┼─────┼───────────────┼─────┼──────────────── 1902│ 250 │ 58 │ 23.2 │ 956 │ 382.4 1903│ 267 │ 59 │ 22.1 │1044 │ 391.0 1904│ 232 │ 24 │ 10.3 │ 530 │ 228.4 ────┴──────────┴─────┴───────────────┴─────┴────────────────

If it were possible to make the necessary distinctions in the case of the large shaft furnace, the diminution in sickness from lead-poisoning would be still more apparent; for, among the furnace attendants proper, there has been no illness; all cases of poisoning have occurred among the men who prepare the charge, who break up the roasted material, and others who are occupied with subsidiary work. Some of these are exposed to illness through their own fault, owing to want of cleanliness, or to neglect of every precautionary measure against lead-poisoning.

Thus far we have dealt only with the advantages and improvements of the Huntington-Heberlein process; we will now, in conclusion, consider also its disadvantages.

The chief drawback of the new process lies in the difficulty of breaking up the blocks of the roasted product from the converters, a labor which, apart from the great expense involved, is also unhealthy for the workmen engaged thereon. Seemingly this evil is still further increased by working with larger charges in the 10 ton converters, as projected; but in this case it is proposed to place the converters in an elevated position, and to cause the blocks to be shattered by their fall from a certain hight, so that further breaking up will require but little work. Trials made in this direction have already yielded satisfactory results, and seem to promise that the disadvantage will in time become less important.

Another unpleasant feature is the presence (in the waste gases from the converters) of a higher percentage of sulphur dioxide, the suppression of which, if it is feasible at all, might be fraught with trouble and expense.

That the roaster gases from the reverberatory-smelting and sintering furnaces did not show such a high percentage of sulphur dioxide must be ascribed chiefly to the circumstance that the roasting was much slower, and that the gases were largely diluted with air already at the point where they are formed, as the work must always be done with the working-doors open. In the Huntington-Heberlein process, on the other hand, the aim is to prevent, as far as possible, the access of air from outside while blowing the charge. The more perfectly this is effected, and the greater the quantity of ore to be blown in the converters, the higher will also be the percentage of sulphur dioxide in the waste gases. This circumstance has not only furnished the inducement, but it has rendered it possible to approach the plan of utilizing the sulphur dioxide for the manufacture of sulphuric acid. If this should be done successfully (which, according to the experiments carried out, there is reasonable ground to expect), the present disadvantage might be turned into an advantage. This has the more significance because an essential constituent of the lead ore—the sulphur—will then no longer, as hitherto, have to be regarded as wholly lost.[31]

THE HUNTINGTON-HEBERLEIN PROCESS

BY THOMAS HUNTINGTON AND FERDINAND HEBERLEIN

(May 26, 1906)

This process for roasting lead sulphide ores has now fairly established itself in all parts of the world, and is recognized by metallurgical engineers as a successful new departure in the method of desulphurization. It offers the great advantage over previous methods of being a more scientific application of the roasting reactions (of the old well-used formulæ PbS + 3O = PbO + SO₂ and PbS + PbSO₄ + 2O = 2PbO + 2SO₂) and admits of larger quantities being handled at a time, so that the use of fuel and labor are in proportion to the results achieved, and also there is less waste all around in so far as the factors necessary for the operation—fuel, labor and air—can be more economically used. The workman’s time and strength are not employed in laboriously shifting the ore from one part of the furnace to another with a maximum amount of exertion and a minimum amount of oxidation. The fuel consumed acts more directly upon the ore during the first part of the process in the furnace and its place is taken by the sulphur itself during the final and blowing stage, so that during the whole series of operations more concentrated gases are produced and consequently the large excess of heated air of the old processes is avoided to such an extent that the gases can be used for the production of sulphuric acid.

With a modern well-constructed plant practically all the evils of the old hand-roasting furnaces are avoided, and besides the notable economy achieved by the H.-H. process itself, the health and well-being of the workmen employed are greatly advanced, so that where hygienic statistics are kept it is proved that lead-poisoning has greatly diminished. It is only natural, therefore, that the H.-H. process should have been a success from the start, popular alike with managers and workmen once the difficulties inseparable from the introduction of any new process were overcome.

Simple as the process now appears, however, it is the result of many years of study and experiment, not devoid of disappointments and at times appearing to present a problem incapable of solution. The first trials were made in the smelting works at Pertusola, Italy, as far back as 1889, where considerable sums were devoted every year to this experimental work and lead ore roasting was almost continuously on the list of new work from 1875 on.

It may be interesting to mention that at this time the Montevecchio ores (containing about 70 per cent. lead and about 15 per cent. sulphur, together with a certain amount of zinc and iron) were considered highly refractory to roast, and the only ores approved of by the management of the works at this date were the Monteponi and San Giovanni first-class ores (containing about 80 per cent. lead), and the second-class carbonates (with at least 60 per cent. lead and 5 per cent. sulphur). It must be noted that a modified Flintshire reverberatory process was in use in the works, which could deal satisfactorily only with this class of ore, so that, as these easy ores diminished in quantity every year and their place was taken by the “refractory” Montevecchio type, the roasting problem was always well to the front at the Pertusola works.

It may be asserted that almost every known method of desulphurization was examined and experimented upon on a large scale. Gas firing was exclusively used on certain classes of ores for several years with considerable success, and revolving furnaces of the Brückner type—gas fired—were also tried. Although varying degrees of success were obtained, no really great progress was made in actual desulphurization; methods were cheapened and larger quantities handled at a time, but the final product—whether sintered or in a pulverulent state—seldom averaged much under 5 per cent. sulphur, while the days of the old “gray slags” (1 per cent. to 2 per cent. sulphur) from the reverberatories totally disappeared, together with the class of ores which produced them.

During the long period of these experiments in desulphurization various facts were established:

(1) That sulphide of lead—especially in a pulverulent state—could not be desulphurized in the same way as other sulphides, such as sulphides of iron, copper, zinc, etc., because if roasted in a mechanical furnace the temperature had to be kept low enough to avoid premature sintering, which would choke the stirrers and cause trouble by the ore clogging on the sides and bottom of the furnace. If, however, the ore was roasted in a “dry state” at low temperature, a great deal of sulphur remained in the product as sulphate of lead, which was as bad for the subsequent blast-furnace work as the sulphide of lead itself. When air was pressed through molten galena—in the same way as through molten copper matte—a very heavy volatilization of lead took place, while portions of it were reduced to metal or were contained as sulphide in the molten matte, so that a good product was not obtained.

(2) That no complete dead roast of lead ores could be obtained unless the final product was thoroughly smelted and agglomerated.

(3) That a well roasted lead ore could be obtained by oxidizing the PbS with compressed air, after the ore had been suitably prepared.

(4) That metal losses were mainly due to the excessive heat produced in the oxidation of PbS to PbO, and other sulphides present in the ore.

It was by making use of these facts that the H.-H. roasting process was finally evolved, and by carefully applying its principles it is possible to desulphurize completely the ore to a practically dead roast of under 1 per cent. sulphur; in practice, however, such perfection is unnecessary and a well agglomerated product with from 2 to 2.5 per cent. sulphur is all that is required. During some trials in Australia, where a great degree of perfection was aimed at, a block of over 2000 tons of agglomerated, roasted ore was produced containing 1 per cent. sulphur (as sulphide); as the ores contained an average of about 10 per cent. Zn, this was a very fine result from a desulphurization point of view, but it was not found that this 1 per cent. product gave any better results in the subsequent smelting in the blast furnace than later on a less carefully prepared material containing 2.5 per cent. sulphur.

In the early stages of experiment the great difficulty was to obtain agglomeration without first fusing the sulphides in the ore, and turning out a half-roasted product full of leady matte. Simple as the thing now is, it seemed at times impossible to avoid this defect, and it was only by a careful study of the effects of an addition of lime, Fe₂O₃ or Mn₂O₃, and their properties that the right path was struck. Before the introduction of the H.-H. process lime was only used in the reverberatory process (Flintshire and Tarnowitz) to stiffen the charge, but as Percy tells us that after its addition the charge was glowing, it must have had a chemical as well as a mechanical effect. In recognition of this fact fine caustic lime or crushed limestone was mixed with the ore _before_ charging it into the furnace and exposing it to an oxidizing heat.

It was thought probable that a dioxide of lime might be temporarily formed, which in contact with PbS would be decomposed immediately after its formation, or that the CaO served as _Contactsubstanz_ in the same way as spongy platinum, metallic silver, or oxide of iron. As CaSO₄ and not CaSO₃ is always found in the roasted ore, this may prove that CaO is really a contact substance for oxygen (see W. M. Hutchings, _Engineering and Mining Journal_, Oct. 21, 1905, Vol. LXXX, p. 726). The fact that the process works equally well with Fe₂O₃ instead of CaO speaks against the theory of plumbate of lime. Whatever theory may be correct, the fact remains that CaO assists the roasting process and that by its use the premature agglomeration of the sulphide ore is avoided. A further advantage of lime is that it keeps the charge more porous and thus facilitates the passage of the air.

The shape and size of the blowing apparatus best adapted for the purpose in view occupied many months; starting from very shallow pans or rectangular boxes several feet square with a few inches of material over a perforated plate, it gradually resolved itself into the cone-shaped receptacle—holding about a ton of ore—as first introduced together with the process. In later years and in treating larger quantities a more hemispherical form has been adopted, containing up to 15 tons of ore.

It is probable about eight years were employed in actually working out the process before it was introduced on any large scale at Pertusola, but by the end of 1898 the greater part of the Pertusola ores were treated by the process. Its first introduction to any other works was in 1900, when it was started outside its home for the first time at Braubach (Germany). Since then its application has gradually extended, proceeding from Europe to Australia and Mexico and finally to America and Canada, where recognition of its merits was more tardy than elsewhere. It is now practically in general use all over the world and is recognized as a sound addition to metallurgical progress. It is doubtless only a step in the right direction and with its general use a better knowledge of its principles will prevail, so that its future development in one direction or another, as compared with present results, may show some further progress.

The present working of the H.-H. process still follows practically the original lines laid down, and by preliminary roasting in a furnace with lime, oxide of iron, or manganese (if not already contained in the ore), prepares the ore for blowing in the converter. Mechanical furnaces have been introduced to the entire exclusion of the old hand-roasters, and the size of the converters has been gradually increased from the original one-ton apparatus successively to 5, 7 and 10 ton converters; at present some for 15 tons are being built in Germany and will doubtless lead to a further economy.

The mechanical furnace at present most in use is a single-hearth revolving furnace with fixed rabbles, the latest being built with a diameter of 26½ ft. and a relatively high arch to ensure a clear flame and rapid oxidation of the ore. The capacity of these furnaces varies, of course, with the nature of the ores to be treated, but with ordinary lead ores (European and Australian practice) of from 50 per cent. to 60 per cent. lead and 14 per cent, to 18 per cent. sulphur, the average capacity may be taken at about 50 to 60 tons of crude ore per day of 24 hours. The consumption of coal with a well-constructed furnace is very low and is always under 8 per cent.—6 per cent. being perhaps the average. These furnaces require very little attention, being automatic in their charging and discharging arrangements.

The ore on leaving the furnace is charged into the converters by various mechanical means (Jacob’s ladders, conveyors, etc.). The converter charge usually consists of some hot ore direct from the furnace, on top of which ore is placed which has been cooled down by storage in bins or by the addition of water. The converter is generally filled in two charges of five tons each, and the blowing time should not be more than 4 to 6 hours. The product obtained should be porous and well agglomerated, but easily broken up, tough melted material being due to an excess of silica and too much lead sulphide. Attention, therefore, to these two points (good preliminary roasting and correction of the charge by lime) obviates this trouble. This roasted ore should not contain more than about 1.5 to 2 per cent. sulphur, and in a modern blast furnace gives surprisingly good results, the matte-fall being in most cases reduced to nothing, and the capacity of the furnace is largely increased, while the slags are poorer.

If the converter charge has been properly prepared, the blowing operation proceeds with the greatest smoothness and requires very little attention on the part of the workmen, the heat and oxidation rise gradually from the bottom and volatilization losses remain low, so that it is possible, if desired, to produce hot concentrated sulphurous gases suitable for the manufacture of sulphuric acid.

Besides the actual economy obtained in roasting ores by the process, a great feature of its success has been the remarkable improvement in smelting and reducing the roasted ore as compared with previous experience. This is due to the nature of the roasted material, which, besides being much poorer in sulphur than was formerly the case, is thoroughly porous and well agglomerated and contains—if the original mixture is properly made—all the necessary slagging materials itself, so that it practically becomes a case of smelting slags instead of ore, and to an expert the difference is evident.

Experience has shown that on an average the improvement in the capacity of the blast furnace may be taken at about 50 to 100 per cent., so that in works using the H.-H. process—after its complete introduction—about half the blast furnaces formerly necessary for the same tonnage were blown out. The matte-fall has become a thing of the past, so that, except in those cases where some matte is required to collect the copper contained in the ores, lead matte has disappeared and the quantity of flue dust as well as the lead and silver losses have been greatly reduced.

Referring to the latest history of the H.-H. process, and the theory of direct blowing, it may be remarked—putting aside all legal questions—that the idea, metallurgically speaking, is attractive, as it would seem that by eliminating one-half of the process and blowing the ores direct without the expense of a preliminary roast a considerable economy should be effected. Upon examination, however, this supposed economy and simplicity is not at all of such great importance, and in many cases, without doubt, would be retrogressive in lead ore smelting rather than progressive. When costs of roasting in a furnace are reduced to such a low figure as can be obtained by using 50 ton furnaces and 10 to 15 ton converters, there is very little margin for improvement in this direction. From the published accounts of the Tarnowitz smelting works (the _Engineering and Mining Journal_, Sept. 23, 1905, Vol. LXXX, p. 535) the cost of mechanical preliminary roasting cannot exceed 25c. per ton, so that even assuming direct blowing were as cheap as blowing a properly prepared material, the total economy would only be the above figure, viz., 25c.; but this is far from being the case.

Direct blowing of a crude ore is considerably more expensive than dealing with the H.-H. product, because of necessity the blowing operation must be carried out slowly and with great care so as to avoid heavy metal losses, and whereas a pre-roasted ore can be easily blown in four hours and one man can attend to two or three 10 ton converters, the direct blowing operation takes from 12 to 18 hours and requires the continual attention of one man. In the first case the cost of labor would be: One man at say $3 for 50 tons (at least), i.e., 6c. per ton, and in the second case one man at $3 for 10 tons (at the best), i.e., 30c., a difference in favor of pre-roasting of 24c., so that any possible economy would disappear. Furthermore, as the danger of blowing upon crude sulphides for 12 or 18 hours is greater as regards metal losses than a quick operation of four hours, it is very likely that instead of an economy there would be an increase in cost, owing to a greater volatilization of metals.

These remarks refer to ordinary lead ores with say 50 per cent. lead and about 14 per cent. sulphur. With ores, however, such as are generally treated in the United States the advantages of pre-roasting are still more evident. These ores contain about 10 to 15 per cent. lead, 30 to 40 per cent. sulphur, 20 to 30 per cent. iron, 10 per cent. zinc, 5 per cent. silica, and lose the greater part of the pyritic sulphur in the preliminary roasting, leaving the iron in the form of oxide, which in the subsequent blowing operation acts in the same way as lime. For this reason the addition of extra fluxes, such as limestone, gypsum, etc., to the original ore is not necessary and only a useless expense.

In certain exceptional cases and with ores poor in sulphur, direct blowing might be applicable, but for the general run of lead ores no economy can be expected by doing away with the preliminary roast.

MAKING SULPHURIC ACID AT BROKEN HILL

(August 11, 1904)

The Broken Hill Proprietary Company has entered upon the manufacture of sulphuric acid on a commercial scale. The acid is practically a by-product, being made from the gases emanating from the desulphurization of the ores, concentrates, etc., by the Carmichael-Bradford process. The acid can be made at a minimum of cost, and thus materially enhances the value of the process recently introduced for the separation of zinc blende from the tailings by flotation. The following particulars are taken from a recently published description of the process: The ores, concentrates, slimes, etc., as the case may be, are mixed with gypsum, the quantity of the latter varying from 15 to 25 per cent. The mixture is then granulated to the size of marbles and dumped into a converter. The bottom of the charge is heated from 400 to 500 deg. C. It is then subjected to an induced current of air, and the auxiliary heat is turned off. The desulphurization proceeds very rapidly with the evolution of heat and the gases containing sulphurous anhydride. The desulphurization is very thorough, and no losses occur through volatilization. The sulphur thus rendered available for acid making is rather more than is contained in the ore, the sulphur in the agglomerated product being somewhat less than that accounted for by the sulphur contained in the added gypsum. Thus from one ton of 14 per cent. sulphide ore it is possible to make about 12 cwt. of chamber acid, fully equaling 7 cwt. of strong acid.

The plant at present in use, which comprises a lead chamber of 40,000 cu. ft., can turn out 35 tons of chamber acid per week. This plant is being duplicated, and it has also been decided to erect a large plant at Port Pirie for use in the manufacture of superphosphates. It is claimed that the production of sulphuric acid from ores containing only 14 per cent. of sulphur establishes a new record.

THE CARMICHAEL-BRADFORD PROCESS

BY DONALD CLARK

(November 3, 1904)

Subsequent to the introduction of the Huntington-Heberlein process in Australia, Messrs. Carmichael and Bradford, two employees of the Broken Hill Proprietary Company, patented a process which bears their name. Instead of starting with lime, or limestone and galena, as in the Huntington-Heberlein process, they discovered that if sulphate of lime is mixed with galena and the temperature raised, on blowing a current of air through the mixture the temperature rises and the mass is desulphurized. The process would thus appear to be a corollary of the original one, and the reactions in the converter are identical. Owing to the success of the acid processes in separating zinc sulphide from the tailing at Broken Hill, it became necessary to manufacture sulphuric acid locally in large quantity. The Carmichael-Bradford process has been started for the purpose of generating the sulphur dioxide necessary, and is of much interest as showing how gases rich enough in SO₂ may be produced from a mixture containing only from 13 to 16 per cent. sulphur.

Gypsum is obtained in a friable state within about five miles from Broken Hill. This is dehydrated, the CaSO, 2H₂O being converted into CaSO₄ on heating to about 200 deg. C. The powdered residue is mixed with slime produced in the milling operations and concentrate in the proportion of slime 3 parts, concentrate 1 part, and lime sulphate 1 part. The proportions may vary to some extent, but the sulphur contents run from 13 to 16 or 17 per cent. The average composition of the ingredients is as given in the table on the next page.

These materials are moistened with water and well mixed by passing them through a pug-mill. The small amount of water used serves to set the product, the lime sulphate partly becoming plaster of paris, 2CaSO, H₂O. While still moist the mixture is broken into pieces not exceeding two inches in diameter and spread out on a drying floor, where excess of moisture is evaporated by the conjoint action of sun and wind.

─────────────────┬─────┬───────────┬────────┬──────── │SLIME│CONCENTRATE│CALCIUM │AVERAGE │ │ │SULPHATE│ ──────────────────┼─────┼───────────┼────────┼──────── Galena │ 24 │ 70 │ │ 29 Blende │ 30 │ 15 │ │ 21 Pyrite │ 3 │ │ │ 2 Ferric oxide │ 4 │ │ │ 2.5 Ferrous oxide │ 1 │ │ │ 1 Manganous oxide │ 6.5│ │ │ 5 Alumina │ 5.5│ │ │ 3 Lime │ 3.5│ │ 41 │ 10 Silica │ 23 │ │ │ 14 Sulphur trioxide │ │ │ 59 │ 12 ─────────────────┴─────┴───────────┴────────┴────────

The pots used are small conical cast-iron ones, hung on trunnions, and of the same pattern as used in the Huntington-Heberlein process. Three of these are set in line, and two are at work while the third is being filled. These pots have the same form of conical cover leading to a telescopic tube, and all are connected to the same horizontal pipe leading to the niter pots. Dampers are provided in each case. A small amount of coal or fuel is fed into the pots and ignited by a gentle blast; as soon as a temperature of about 400 to 500 deg. C. is attained the dried mixture is fed in, until the pot is full; the cover is closed down and the mass warms up. Water is first driven off, but after a short time concentrated fumes of sulphur dioxide are evolved. The amount of this gas may be as much as 14 per cent., but it is usually kept at about 10 per cent., so as to have enough oxygen for the conversion of the dioxide to the trioxide. The gases are led over a couple of niter pots and thence to the usual type of lead chamber having a capacity of 40,000 cu. ft. Chamber acid alone is made, since this requires to be diluted for what is known as the saltcake process.

The plant has now been in operation for some time and, it is claimed, with highly successful results. The product tipped out of. the converter is similar to that obtained in the Huntington-Heberlein process, and is at once fit for the smelters, the amount of sulphur left in it being always less than that originally introduced with the gypsum; analysis of the desulphurized material shows usually from 3 to 4 per cent. sulphur.

THE CARMICHAEL-BRADFORD PROCESS

BY WALTER RENTON INGALLS

(October 28, 1905)

As described in United States patent No. 705,904, issued July 29, 1902, lead sulphide ore is mixed with 10 to 35 per cent. of calcium sulphate, the percentage varying according to the grade of the ore. The mixture is charged into a converter and gradually heated externally until the lower portion of the charge, say one-third to one-fourth, is raised to a dull-red heat; or the reactions may be started by throwing into the empty converter a shovelful of glowing coal and turning on a blast of air sufficient to keep the coal burning and then feeding the charge on top of the coal. This heating effects a reaction whereby the lead sulphide of the ore is oxidized to sulphate and the calcium sulphate is reduced to sulphide. The heated mixture being continuously subjected to the blast of air, the calcium sulphide is re-oxidized to sulphate and is thus regenerated for further use. This reaction is exothermic, and sufficient heat is developed to complete the desulphurization of the charge of ore by the concurrent reactions between the lead sulphate (produced by the calcium sulphate) and portions of undecomposed ore, sulphurous anhydride being thus evolved. The various reactions, which are complicated in their nature, continue until the temperature of the charge reaches a maximum, by which time the charge has shrunk considerably in volume and has a tendency to become pasty. This becomes more marked as the production of lead oxide increases, and as the desired point of desulphurization is attained the mixture fuses; at this stage the calcium sulphide which is produced from the sulphate cannot readily oxidize, owing to the difficulty of coming into actual contact with the air in the pasty mass, but, being subjected to the strong oxidizing effect of the metallic oxide, it is converted into calcium plumbate, while sulphurous anhydride is set free. The mass then cools, as the exothermic reactions cease, and can be readily removed to a blast furnace for smelting.

The reactions above described are as outlined in the original American patent specification. Irrespective of their accuracy, the Carmichael-Bradford process is obviously quite similar to the Huntington-Heberlein, and doubtless owes its origin to the latter. The difference between them is that in the Huntington-Heberlein process the ore is first partially roasted with addition of lime, and is then converted in a special vessel. In the Carmichael-Bradford process the ore is mixed with gypsum and is then converted directly. The greatest claim for originality in the Carmichael-Bradford process may be considered to lie in it as a method of desulphurizing gypsum, inasmuch as not only is the sulphur of the ore expelled, but also a part of the sulphur of the gypsum; and the sulphur is driven off as a gas of sufficiently high tenor of sulphur dioxide to enable sulphuric acid to be made from it economically. Up to the present time the Carmichael-Bradford process has been put into practical use only at Broken Hill, N. S. W.

The Broken Hill Proprietary Company first conducted a series of tests in a converter capable of treating a charge of 20 cwt. These tests were made at the smelting works at Port Pirie. Exhaustive experiments made on various classes of ores satisfactorily proved the general efficacy of the process. The following ores were tried in these preliminary experiments, viz.:

First-grade concentrate containing: Pb, 60 per cent.; Zn, 10 per cent.; S, 16 per cent.; Ag, 30 oz.

Second-grade concentrate containing: Pb, 45 per cent.; Zn, 12.5 per cent.; S, 14.5 per cent.; Ag, 22 oz.

Slime containing: Pb, 21 per cent.; Zn, 17 per cent.; S, 13 per cent.; Ag, 18 oz.

Lead-copper matte containing: Fe, 42 per cent.; Pb, 17 per cent.; Zn, 13.3 per cent.; Cu, 2.4 per cent.; S, 23 per cent.; Ag, 25 oz.

Other mattes, of varying composition up to 45 per cent. Pb and 100 oz. Ag, were also tried.

The results from these preliminary tests were so gratifying that a further set of tests was made on lead-zinc slime, with a view of ascertaining whether any volatilization losses occurred during the desulphurization. This particular material was chosen because of its accumulation in large proportions at the mine, and the unsatisfactory result of the heap roasting which has recently been practised. The heap roasting, although affording a product containing only 7 per cent. S, which is delivered in lump form and therefore quite suitable for smelting, resulted in a high loss of metal by volatilization (17 per cent. Pb, 5 per cent. Ag).

The result of nine charges of the slime treated by the Carmichael-Bradford process was as follows:

─────────────────┬──────┬─────────────────────┬─────────────────────── │ │ ASSAYS │ CONTENTS │ Cwt. ├────┬──────┬────┬────┼─────┬─────┬────┬────── │ │Pb% │Ag oz.│Zn% │ S% │ Pb │ Ag. │ Zn │ S │ │ │ │ │ │cwt. │ oz. │cwt.│cwt. ├──────┼────┼──────┼────┼────┼─────┼─────┼────┼────── Raw slime │128.1 │21.3│ 18.0 │16.8│13.1│27.28│115.3│26.2│16.78 Raw gypsum │ 54.9 │ │ │ │ │ │ │ │ 9.88 ├──────┤ │ │ │ ├─────┼─────┼────┼────── Total │183.0 │ │ │ │ │27.28│115.3│25.2│26.66 ──────────────────┼──────┼────┼──────┼────┼────┼─────┼─────┼────┼────── Sintered material│109.88│20.7│ 17.2 │ │4.80│22.74│ 94.5│ │ 5.27 Middling │ 14.47│17.7│ 15.7 │ │6.20│ 2.56│ 11.3│ │ 0.89 Fines │ 11.12│19.0│ 14.8 │ │7.50│ 2.11│ 8.2│ │ 0.83 ├──────┼────┼──────┼────┼────┼─────┼─────┼────┼────── Total │135.47│ │ │ │5.17│27.41│113.0│ │ 6.99 ─────────────────┴──────┴────┴──────┴────┴────┴─────┴─────┴────┴──────

These results indicated practically no volatilization of lead and silver during the treatment, the lead showing a slight increase, viz., 0.47 per cent., and the silver 1.13 per cent. loss. A desulphurization of 70.4 per cent. was effected. A higher desulphurization could have been effected had this been desired. In the above tabulated results, the term “middling” is applied to the loose fritted lumps lying on the top of the charge: these are suitable for smelting, the fines being the only portion which has to be returned.

In order to test the practicability of making sulphuric acid, a plant consisting of three large converters of capacity of five tons each, together with a lead chamber 100 ft. by 20 ft. by 20 ft., was then erected at Broken Hill, together with a dehydrating furnace, pug-mill, and granulator. These converters are shown in the accompanying engravings.

A trial run was made with 108 tons of concentrate of the following composition: 54 per cent. lead; 1.9 per cent. iron; 0.9 per cent. manganese; 9.4 per cent. zinc; 14.6 per cent. sulphur; 19.2 per cent. insoluble residue, and 24 oz. silver per ton.

The converter charge consisted of 100 parts of the concentrate and 25 parts of raw gypsum, crushed to pass a 1 in. hole and retained by a 0.25 in. hole, the material finer than 0.25 in. (which amounted to 5 per cent. of the total) being returned to the pug-mill. After desulphurization in the converter, the product assayed as follows: 48.9 per cent. lead; 1.80 per cent. iron; 0.80 per cent. manganese; 7.87 per cent. zinc; 3.90 per cent. sulphur; 1.02 per cent. alumina; 5.80 per cent. lime; 21.75 per cent. insoluble residue; 8.16 per cent. undetermined (oxygen as oxides, sulphates, etc.); total, 100 per cent. Its silver content was 22 oz. per ton. The desulphurized ore weighed 10 per cent. more than the raw concentrate. During this run 34 tons of acid were made.

A trial was then made on 75 tons of slime of the following composition: 18.0 per cent. lead; 16.6 per cent. zinc; 6.0 per cent. iron; 2.5 per cent. manganese; 3.2 per cent. alumina; 2.1 per cent. lime; 38.5 per cent. insoluble residue; total, 100 per cent. Its silver content was 19.2 oz. per ton.

The converter charge in this case consisted of 100 parts of raw slime and 30 parts of gypsum. The converted material assayed as follows: 16.1 per cent. lead; 14.0 per cent. zinc; 3.6 per cent. sulphur; 5.42 per cent. iron; 2.25 per cent. manganese; 4.10 per cent. alumina; 8.60 per cent. lime; 39.80 per cent. insoluble residue; 6.13 per cent. undetermined (oxygen, etc.); total, 100 per cent.; and silver, 17.5 oz. per ton. The increase in weight of desulphurized ore over that of the raw ore was 11 per cent. During this run 22 tons of acid were manufactured.

The analysis of the gypsum used in each of the above tests (at Broken Hill) was as follows: 76.1 per cent. CaSO₄, 2H₂O; 0.5 per cent. Fe₂O₃; 4.5 per cent. Al₂O₃; 18.9 per cent. insoluble residue.

The plant was then put into continuous operation on a mixture of three parts slime and one of concentrate, desulphurizing down to 4 per cent. S, and supplying 20 tons of acid per week, and additions were made to the plant as soon as possible. The acid made at Broken Hill has been used in connection with the Delprat process for the concentration of the zinc tailing. At Port Pirie, works are being erected with capacity for desulphurization of about 35,000 tons per annum, with an acid output of 10,000 tons. This acid is to be utilized for the acidulation of phosphate rock.

The cost of desulphurization of a ton of galena concentrate by the Carmichael-Bradford process, based on labor at $1.80 per 8 hours, gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb., is estimated as follows:

0.25 ton of gypsum $0.60 Dehydrating and granulating gypsum .48 Drying mixture of ore and gypsum .12 Converting .24 Spalling sintered material .12 0.01 ton coal .08 ——-——- Total $1.64

The lime in the sintered product is credited at 12c., making the net cost $1.52 per ton (2240 lb.) of ore.

The plant required for the Carmichael-Bradford process can be described with sufficient clearness without drawings, except the converters. The ore (concentrate, slime, etc.) to be desulphurized is delivered at the top of the mill by cars, conveyors, or other convenient means, and dumped into a bin. Two screw feeders placed inside the bin supply the mill with ore, uniformly and as fast as it is required. These feeders deliver the ore into a chute, which directs it into a vertical dry mixer.

A small bin, on the same level as the ore-bin, receives the crude gypsum from cars. Thence it is fed automatically to a disintegrator, which pulverizes it finely and delivers it into a storage bin underneath. This disintegrator revolves at about 1700 r.p.m. and requires 10 h.p. The body of the machine is cast iron, fitted with renewable wearing plates (made of hard iron) in the grinding chamber. The revolving parts consist of a malleable iron disc in which are fixed steel beaters, faced on the grinding surface with highly tempered steel. The bin that receives the floured gypsum contains a screw conveyor similar to those in the ore-bin, and dumps the material into push conveyors passing into the dehydrating furnace. They carry the crushed gypsum along at a speed of about 1 ft. per minute, and allow about 20 ft. to dehydrate the gypsum. This speed can, of course, be regulated to suit requirements.

The dehydrated gypsum runs down a chute into an elevator boot, and is elevated into a bin which is on the same level as the ore-bin. This bin also contains a screw conveyor, like that in the ore-bin. The speed of delivery is regulated to deliver the right proportion of dehydrated gypsum to the mixer.

The mixer is of the vertical pattern and receives the sulphide ore and dehydrated gypsum from the screw feeders. In it are set two flat revolving cones running at different speeds, thus ensuring a thorough mixture of the gypsum and ore. The mixed material drops from the cones upon two baffle plates, and is wetted just before entering the pug-mill. The pug-mill is a wrought-iron cylinder of ¼ in. plate about 2 ft. 6 in. diameter and 6 or 8 ft. long, and has the mixer fitted to the head. It contains a 3 ft. wrought-iron spiral with propelling blades, which forces the plastic mixture through ¾ in. holes in the cover. The material comes out in long cylindrical pieces, but is broken up and formed into marble-shaped pieces on dropping into a revolving trommel.

The trommel is about 5 ft. long, 2 ft. in diameter at the small end and about 4 ft. at the large end. It revolves about a wrought-iron spindle (2½ in. diameter) carrying two cast-iron hubs to which are fitted arms for carrying the conical plate ⅛ in. thick. About 18 in. of the small end of the cone is fitted with wire gauze, so as to prevent the material as it comes out of the pug-mill from sticking to it. The trommel is driven by bevel gearing at 20 to 25 r.p.m. The granulated material formed in the trommel is delivered upon a drying conveyor.

The conveyor consists of hinged wrought-iron plates flanged at the side to keep the material from running off. It is driven from the head by gearing, at a speed of 1 ft. per minute, passing through a furnace 10 ft. long to dry and set the granules of ore and gypsum. This speed can, of course, be regulated to suit requirements. The granulated material, after leaving the furnace, is delivered to a single-chain elevator, traveling at a speed of about 150 ft. per minute. It drops the material into a grasshopper conveyor, driven by an eccentric, which distributes the material over the length of a storage bin. From this bin the material is directed into the converters by means of the chutes, which have their bottom ends hinged so as to allow for the raising of the hood when charging the converters.

The converters are shown in the accompanying engravings, but they may be of slightly different form from what is shown therein, i.e., they may be more spherical than conical. They will have a capacity of about four tons, being 6 ft. in diameter at the top, 4 ft. in diameter at the false bottom, and about 5 ft. deep. They are swung on cast-iron trunnions bolted to the body and turned by means of a hand-wheel and worm (not shown). They are carried on strong cast-iron standards fitted with bearings for trunnions, and all necessary brackets for tilting gear. The hood has a telescopic funnel which allows it to be raised or lowered, weights being used to balance it. At the apex of the cone a damper is provided to regulate the draft. A 4 in. hole in the pot allows the air from the blast-pipe, 18 in. in diameter, to enter under the false perforated bottom, the connection between the two being made by a flexible pipe and coupling. Two Baker blowers supply the blast for the converters. The material, after being sintered, is tipped on the floor in front of the converters and is there broken up to any suitable size, and thence dispatched to the smelters.

The necessary power for a plant with a capacity of 150 tons of ore per day will be supplied by a 50 h.p. engine.

THE SAVELSBERG PROCESS

BY WALTER RENTON INGALLS

(December 9, 1905)

There are in use at the present time three processes for the desulphurization of galena by the new method, which has been referred to as the “lime-roasting of galena.” The Huntington-Heberlein and the Carmichael-Bradford processes have been previously described. The third process of this type, which in some respects is more remarkable than either of the others, is the invention of Adolf Savelsberg, director of the smeltery at Ramsbeck, Westphalia, Germany, which is owned by the Akt. Gesell. f. Bergbau, Blei. u. Zinkhüttenbetrieb zu Stolberg u. in Westphalen. The process is in use at the Ramsbeck and Stolberg lead smelteries of that company. It is described in American patent No. 755,598, issued March 22, 1904 (application filed Dec. 18, 1903). The process is well outlined in the words of the inventor in the specification of that patent:

“The desulphurizing of certain ores has been effected by blowing air through the ore in a chamber, with the object of doing away with the imperfect and costly process of roasting in ordinary furnaces; but hitherto it has not been possible satisfactorily to desulphurize lead ores in this manner, as, if air be blown through raw lead ores in accordance with either of the processes used for treating copper ores, for example, the temperature rises so rapidly that the unroasted lead ore melts and the air can no longer act properly upon it, because by reason of this melting the surface of the ores is considerably decreased, the greater number of points or extent of surface which the raw ore originally presented to the action of the oxygen of the air blown through being lost, and, moreover, the further blowing of air through the molten mass of ore produces metallic lead and a plumbiferous slag (in which the lead oxide combines with the gangue) and also a large amount of light dust, consisting mainly of sublimated lead sulphide. Huntington and Heberlein have proposed to overcome these objections by adopting a middle course, consisting in roasting the ores with the addition of limestone for overcoming the ready fusibility of the ores, and then subjecting them to the action of the current of air in the chamber; but this process is not satisfactory, because it still requires the costly previous operation in a roasting furnace.

“My invention is based on the observation which I have made that if the lead ores to be desulphurized contain a sufficient quantity of limestone it is possible, by observing certain precautions, to dispense entirely with the previous roasting in a roasting furnace, and to desulphurize the ores in one operation by blowing air through them. I have found that the addition of limestone renders the roasting of the lead ore unnecessary, because the limestone produces the following effects:

“The particles of limestone act mechanically by separating the particles of lead ore from each other in such a way that premature agglomeration is prevented and the whole mass is loosened and rendered accessible to air; and, moreover, the limestone moderates the high reaction temperature resulting from the burning of the sulphur, so that the liquefaction of the galena, the sublimation of lead sulphide, and the separation of metallic lead are avoided. The moderation of the temperature of reaction is caused by the decomposition of the limestone into caustic lime and carbon dioxide, whereby a large amount of heat becomes latent. Further, the decomposition of the limestone causes chemical reactions, lime being formed, which at the moment of its formation is partly converted into sulphate of lime at the expense of the sulphur contained in the ore, and this sulphate of lime, when the scorification takes place, is transformed into calcium silicate by the silicic acid, the sulphuric acid produced thereby escaping. The limestone also largely contributes to the desulphurization of the ore, as it causes the production of sulphuric acid at the expense of the sulphur of the ore, which sulphuric acid is a powerful oxidizing agent. If, therefore, a mixture of raw lead ore and limestone (which mixture must, of course, contain a sufficient amount of silicic acid for forming silicates) be introduced into a chamber and a current of air be blown through the mixture, and at the same time the part of the mixture which is near the blast inlet be ignited, the combustion of the sulphur will give rise to very energetic reactions, and sulphurous acid, sulphuric acid, lead oxide, sulphates and silicates are produced. The sulphurous acid and the carbon dioxide escape, while the sulphuric acid and sulphates act in their turn as oxidizing agents on the undecomposed galena. Part of the sulphates is decomposed by the silicic acid, thereby liberating sulphuric acid, which, as already stated, acts as an oxidizing agent. The remaining lead oxide combines finally with the gangue of the ore and the non-volatile constituents of the flux (the limestone) to form the required slag. These several reactions commence at the blast inlet at the bottom of the chamber, and extend gradually toward the upper portion of the charge of ore and limestone. Liquefaction of the ores does not take place, for although a slag is formed it is at once solidified by the blowing in of the air, the passages formed thereby in the hardening slag allowing of the continued passage therethrough of the air. The final product is a silicate consisting of lead oxide, lime, silicic acid, and other constituents of the ore, which now contains but little or no sulphur and constitutes a coherent solid mass, which, when broken into pieces, forms a material suitable to be smelted.

“The quantity of limestone required for the treatment of the lead ores varies according to the constitution of the ores. It should, however, amount generally to from 15 to 20 per cent. As lead ores do not contain the necessary amount of limestone as a natural constituent, a considerable amount of limestone must be added to them, and this addition may be made either during the dressing of the ores or subsequently.

“For the satisfactory working of the process, the following precautions are to be observed: In order that the blowing in of the air may not cause particles of limestone to escape in the form of dust before the reaction begins, it is necessary to add to the charge before it is subjected to the action in the chamber a considerable amount of water—say 5 per cent. or more. This water prevents the escape of dust, and it also contributes considerably to the formation of sulphuric acid, which, by its oxidizing action, promotes the reaction, and, consequently, also the desulphurization. It is advisable, in conducting the operation, not to fill the chamber with the charge at once, but first only partly to fill it and add to the charge gradually while the chamber is at work, as by this means the reaction will take place more smoothly in the mass.

“It is advantageous to proceed as follows: The bottom part of a chamber of any suitable form is provided with a grate, on which is laid and ignited a mixture of fuel (coal, coke, or the like) and pieces of limestone. By mixing the fuel with pieces of limestone the heating power of the fuel is reduced and the grate is protected, while at the same time premature melting of the lower part of the charge is prevented; or the grate may be first covered with a layer of limestone and the fuel be laid thereon, and then another layer of limestone be placed on the fuel. On the material thus placed in the chamber, a uniform charge of lead ore and limestone—say about 12 in. high—is placed, this having been moistened as previously explained. Under the influence of the air-blast and the heat, the reactions hereinbefore described take place. When the upper surface of the first layer becomes red-hot, a further charge is laid thereon, and further charges are gradually introduced as the surface of the preceding charge becomes red-hot, until the chamber is full. So long as charges are still introduced a blast of air of but low pressure is blown through; but when the chamber is filled a larger quantity of air at a higher pressure is blown through. The scorification process then takes place, a very powerful desulphurization having preceded it. During the scorification the desulphurization is completed.

“When the process is completed, the chamber is tilted and the desulphurized mass falls out and is broken into small pieces for smelting.”

The drawing on page 190, Fig. 17, shows a side view of the apparatus used in connection with the process, which will be readily understood without special description. The dotted lines show the pot in its emptying position. The series of operations is clearly illustrated in Figs. 18-20, which are reproduced from photographs.

This process has now been in practical use at Ramsbeck for three years, where it is employed for the desulphurization of galena of high grade in lead, with which are mixed quartzose silver ore (or sand if no such ore be available), and calcareous and ferruginous fluxes. A typical charge is 100 parts of lead ore, 10 parts of quartzose silver ore, 10 parts of spathic iron ore, and 19 parts of limestone. A thorough mixture of the components is essential; after the mixture has been effected, the charge is thoroughly wetted with about 5 per cent. of water, which is conceived to play a threefold function in the desulphurizing operation, namely: (1) preservation of the homogeneity of the mixture during the blowing; (2) reduction of temperature during the process; and (3) formation of sulphuric acid in the process, which promotes the desulphurization of the ore.

The moistened charge is conveyed to the converters, into which it is fed in thin layers. The converters are hemispherical cast-iron pots, supported by trunnions on a truck, as shown in the accompanying engravings. Except for this method of support, which renders the pot movable, the arrangement is quite similar to that which is employed in the Huntington-Heberlein process. The pots which are now in use at Ramsbeck have capacity for about 8000 kg. of charge, but it is the intention of the management to increase the capacity to 10,000 or 12,000 kg. Previously, pots of only 5000 kg. capacity were employed. Such a pot weighed 1300 kg., exclusive of the truck. The air-blast was about 7 cu. m. (247.2 cu. ft.) per min., beginning at a pressure of 10 to 20 cm. of water (2¾ to 4½ oz.) and rising to 50 to 60 cm. (11½ to 13½ oz.) when the pot was completely filled with charge. The desulphurization of a charge is completed in 18 hours. A pot is attended by one man per shift of 12 hours; this is only the attention of the pot proper, the labor of conveying material to it and breaking up the desulphurized product being extra. One man per shift should be able to attend to two pots, which is the practice in the Huntington-Heberlein plants.

When the operation in the pot is completed, the latter is turned on its trunnions, until the charge slides out by gravity, which it does as a solid cake. This is caused to fall upon a vertical bar, which breaks it into large pieces. By wedging and sledging these are reduced to lumps of suitable size for the blast furnace. When the operation has been properly conducted the charge is reduced to about 2 or 3 per cent. sulphur. It is expected that the use of larger converters will show even more favorable results in this particular.

As in the Huntington-Heberlein and Carmichael-Bradford processes, one of the greatest advantages of the Savelsberg process is the ability to effect a technically high degree of desulphurization with only a slight loss of lead and silver, which is of course due to the perfect control of the temperature in the process. The precise loss of lead has not yet been determined, but in the desulphurization of galena containing 60 to 78 per cent. lead, the loss of lead is probably not more than 1 per cent. There appears to be no loss of silver.

The process is applicable to a wide variety of lead-sulphide ores. The ore treated at Ramsbeck contains 60 to 78 per cent. lead and about 15 per cent. of sulphur, but ore from Broken Hill, New South Wales, containing 10 per cent. of zinc has also been treated. A zinc content up to 7 or 8 per cent. in the ore is no drawback, but ores carrying a higher percentage of zinc require a larger addition of silica and about 5 per cent. of iron ore in order to increase the fusibility of the charge. The charge ordinarily treated at Ramsbeck is made to contain about 11 per cent. of silica. The presence of pyrites in the ore is favorable to the desulphurization. Dolomite plays the same part in the process that limestone does, but is of course less desirable, in view of the subsequent smelting in the blast furnace. The ore is best crushed to about 3 mm. size, but good results have been obtained with ore coarser in size than that. However, the proper size is somewhat dependent upon the character of the ore. The blast pressure required in the converter is also, of course, somewhat dependent upon the porosity of the charge. Fine slimes are worked up by mixture with coarser ore.

In making up the charge, the proportion of limestone is not varied much, but the proportions of silica and iron must be carefully modified to suit the ore. Certain kinds of ore have a tendency to remain pulverulent, or to retain balls of unsintered, powdered material. In such cases it is necessary to provide more fusible material in the charge, which is done by varying the proportions of silica and iron. The charge must, moreover, be prepared in such a manner that overheating, and consequently the troublesome fusion of raw galena, will be avoided.

The essential difference between the Huntington-Heberlein and Savelsberg processes is the use in the former of a partially desulphurized ore, containing lime and sulphate of lime; and the use in the latter of raw ore and carbonate of lime. It is claimed that the latter, which loses its carbon dioxide in the converter, necessarily plays a different chemical part from that of quicklime or gypsum. Irrespective of the reactions, however, the Savelsberg process has the great economic advantage of dispensing with the preliminary roasting of the Huntington-Heberlein process, wherefore it is cheaper both in first cost of plant and in operation.

THE LIME-ROASTING OF GALENA[32]

BY WALTER RENTON INGALLS

During the last two years, and especially during the last six months, a number of important articles upon the new methods for the desulphurization of galena have been published in the technical periodicals, particularly in the _Engineering and Mining Journal_ and in _Metallurgie_. I proposed for these methods the type-name of “lime-roasting of galena,” as a convenient metallurgical classification,[33] and this term has found some acceptance. The articles referred to have shown the great practical importance of these new processes, and the general recognition of their metallurgical and commercial value, which has already been accorded to them. It is my present purpose to review broadly the changes developed by them in the metallurgy of lead, in which connection it is necessary to refer briefly to the previous state of the art.

The elimination of the sulphur content of galena has been always the most troublesome part of the smelting process, being both costly in the operation and wasteful of silver and lead. Previous to the introduction of the Huntington-Heberlein process at Pertusola, Italy, it was effected by a variety of methods. In the treatment of non-argentiferous galena concentrate, the smelting was done by the roast-reduction method (roasting in reverberatory furnace and smelting in blast furnace); the roast-reaction method, applied in reverberatory furnaces; and the roast-reaction method, applied in Scotch hearths.[34] Precipitation smelting, simple, had practically gone out of use, although its reactions enter into the modern blast-furnace practice, as do also those of the roast-reaction method.

In the treatment of argentiferous lead ores, a combination of the roast-reduction, roast-reaction and precipitation methods had been developed. Ores low in lead were still roasted, chiefly in hand-worked reverberatories (the mechanical furnaces not having proved well adapted to lead-bearing ores), while the high loss of lead and silver in sinter-or slag-roasting of rich galenas had caused those processes to be abandoned, and such ores were charged raw into the blast furnace, the part of their sulphur which escaped oxidation therein reappearing in the form of matte. In the roast-reduction smelting of galena alone, however, there was no way of avoiding the roasting of the whole, or at least a very large percentage of the ore, and in this roasting the ore had necessarily to be slagged or sintered in order to eliminate the sulphur to a satisfactory extent. This is exemplified in the treatment of the galena concentrate of southeastern Missouri at the present time.

Until the two new Scotch-hearth plants at Alton and Collinsville, Ill., were put in operation, the three processes of smelting the southeastern Missouri galena were about on an equal footing. Their results per ton of ore containing 65 per cent. lead were approximately as follows[35]:

──────────────────┬──────────────┬──────────── METHOD │ COST │ EXTRACTION ──────────────────┼──────────────┼─────────── Reverberatory │ $6.50-7.00 │ 90-92% Scotch hearth │ 5.75-6.50 │ 87-88% Roast-reduction │ 6.00-7.00 │ 90-92% ──────────────────┴──────────────┴───────────

The new works employ the Scotch-hearth process, with bag-houses for the recovery of the fume, which previously was the weak point of this method of smelting.[36] This improvement led to a large increase in the recovery of lead, so that the entire extraction is now approximately 98 per cent. of the content of the ore, while on the other hand the cost of smelting per ton of ore has been reduced through the increased size of these plants and the introduction of improved means for handling ore and material. The practice of these works represents the highest efficiency yet obtained in this country in the smelting of high-grade galena concentrate, and probably it cannot be equaled even by the Huntington-Heberlein and similar processes. The Scotch-hearth and bag-house process is therefore the one of the older methods of smelting which will survive.

In the other methods of smelting, a large proportion of the cost is involved in the roasting of the ore, which amounts in hand-worked reverberatory furnaces to $2 to $2.50 per ton. Also, the larger proportion of the loss of metal is suffered in the roasting of the ore, this loss amounting to from 6 to 8 per cent. of the metal content of such ore as is roasted. The loss of lead in the combined process of treatment depends upon the details of the process. The chief advantage of lime-roasting in the treatment of this class of ore is in the higher extraction of metal which it affords. This should rise to 98 per cent. That figure has been, indeed, surpassed in operations on a large scale, extending over a considerable period.

In the treatment of the argentiferous ores of the West different conditions enter into the consideration. In the working of those ores, the present practice is to roast only those which are low in lead, and charge raw into the blast furnace the rich galenas. The cost of roasting is about $2 to $2.50 per ton; the cost of smelting is about $2.50 per ton. On the average about 0.4 ton of ore has to be roasted for every ton that is smelted. The cost of roasting and smelting is therefore about $3.50 per ton. In good practice the recovery of silver is about 98 per cent. and of lead about 95 per cent., reckoned on basis of fire assays.

In treatment of these ores, the lime-roasting process offers several advantages. It may be performed at less than the cost of ordinary roasting.[37] The loss of silver and lead during the roasting is reduced to insignificant proportion. The sulphide fines which must be charged raw into the blast furnace are eliminated, inasmuch as they can be efficiently desulphurized in the lime-roasting pots without significant loss; all the ore to be smelted in the blast furnace can be, therefore, delivered to it in lump form, whereby the speed of the blast furnace is increased and the wind pressure required is decreased. Finally, the percentage of sulphur in the charge is reduced, producing a lower matte-fall, or no matte-fall whatever, with consequent saving in expense of retreatment. In the case of a new plant, the first cost of construction and the ground-space occupied are materially reduced. Before discussing more fully the extent and nature of these savings, it is advisable to point out the differences among the three processes of lime-roasting that have already come into practical use.

In the Huntington-Heberlein process, the ore is mixed with suitable proportions of limestone and silica (or quartzose ore) and is then partially roasted, say to reduction of the sulphur to one half. The roasting is done at a comparatively low temperature, and the loss of metals is consequently small. The roasted ore is dampened and allowed to cool. It is then charged into a hemispherical cast-iron pot, with a movable hood which covers the top and conveys off the gases. There is a perforated grate in the bottom of the pot, on which the ore rests, and air is introduced through a pipe entering the bottom of the pot, under the grate. A small quantity of red-hot calcines from the roasting furnaces is thrown on the grate to start the reaction; a layer of cold, semi-roasted ore is put upon it, the air blast is turned on and reaction begins, which manifests itself by the copious evolution of sulphur fumes. These consist chiefly of sulphur dioxide, but they contain more or less trioxide, which is evident from the solution of copperas that trickles from the hoods and iron smoke-pipes, wherein the moisture condenses. As the reaction progresses, and the heat creeps up, more ore is introduced, layer by layer, until the pot is full. Care is taken by the operator to compel the air to pass evenly and gently through the charge, wherefore he is watchful to close blow-holes which develop in it. At the end of the operation, which may last from four to eighteen hours, the ore becomes red-hot at the top. The hood is then pushed up, and the pot is turned on its trunnions, by means of a hand-operated wheel and worm-gear, until the charge slides out, which it does as a solid, semi-fused cake. The pot is then turned back into position. Its design is such that the air-pipe makes automatic connection, a flanged pipe cast with the pot settling upon a similarly flanged pipe communicating with the main, a suitable gasket serving to make a tight joint. The pots are set at an elevation of about 12 ft. above the ground, so that when the charge slides out the drop will break it up to some extent, and it is moreover caused to fall on a wedge, or similar contrivance, to assist the breakage. After cooling it is further broken up to furnace size by wedging and sledging; the lumps are forked out, and the fines screened and returned to a subsequent charge for completion of their desulphurization.

The Savelsberg process differs from the Huntington-Heberlein in respect to the preliminary roasting, which in the Savelsberg process is omitted, the raw ore, mixed with limestone and silica, being charged directly into the converter. The Savelsberg converter is supported on a truck, instead of being fixed in position, but otherwise its design and management are quite similar to those of the Huntington-Heberlein converter. In neither case are there any patents on the converters. The patents are on the processes. In view of the litigation that has already been commenced between their respective owners, it is interesting to examine the claims.

The Huntington-Heberlein patent (U. S. 600,347, issued March 8, 1898, applied for Dec. 9, 1896) has the following claims:

1. The herein-described method of oxidizing sulphide ores of lead preparatory to reduction to metal, which consists in mixing with the ore to be treated an oxide of an alkaline-earth metal, such as calcium oxide, subjecting the mixture to heat in the presence of air, then reducing the temperature and finally passing air through the mass to complete the oxidation of the lead, substantially as and for the purpose set forth.

2. The herein-described method of oxidizing sulphide ores of lead preparatory to reduction to metal, which consists in mixing calcium oxide or other oxide of an alkaline-earth metal with the ore to be treated, subjecting the mixture in the presence of air to a bright-red heat (about 700 deg. C.), then cooling down the mixture to a dull-red heat (about 500 deg. C.), and finally forcing air through the mass until the lead ore, reduced to an oxide, fuses, substantially as set forth.

3. The herein-described method of oxidizing lead sulphide in the preparation of the same for reduction to metal, which consists in subjecting the sulphide to a high temperature in the presence of an oxide of an alkaline-earth metal, such as calcium oxide, and oxygen, and then lowering the temperature substantially as set forth.

Adolf Savelsberg, in U. S. patent 755,598 (issued March 22, 1904, applied for Dec. 18, 1903) claims:

1. The herein-described process of desulphurizing lead ores, which consists in mixing raw ore with limestone and then subjecting the mixture to the simultaneous application of heat and a current of air in sufficient proportions to substantially complete the desulphurization in one operation, substantially as described.

2. The herein-described process of desulphurizing lead ores, which process consists in first mixing the ores with limestone, then moistening the mixture, then filling it without previous roasting into a chamber, then heating it and treating it by a current of air, as and for the purpose described.

3. The herein-described process of desulphurizing lead ores, which consists in mixing raw ores with limestone, then filling the mixture into a chamber, then subjecting the mixture to the simultaneous application of heat and a current of air in sufficient proportions to substantially complete the desulphurization in one operation, the mixture being introduced into the chamber in partial charges introduced successively at intervals during the process, substantially as described.

4. The herein-described process of desulphurizing lead ores, then moistening the mixture, then filling it without previous roasting into a chamber, then heating it and treating it by a current of air, the mixture being introduced into the chamber in partial charges introduced successively at intervals during the process, as and for the purpose described.

5. The herein-described process of desulphurizing lead ores, which process consists in first mixing the ores with sufficient limestone to keep the temperature of the mixture below the melting-point of the ore, then filling the mixture into a chamber, then heating said mixture and treating it with a current of air, as and for the purpose described.

6. The herein-described process of desulphurizing lead ores, which process consists in first mixing the ores with sufficient limestone to mechanically separate the particles of galena sufficiently to prevent fusion, and to keep the temperature below the melting-point of the ore by the liberation of carbon dioxide, then filling the mixture into a chamber, then heating said mixture and treating it with a current of air, as and for the purpose described.

The Carmichael-Bradford process differs from the Savelsberg by the treatment of the raw ore mixed with gypsum instead of limestone, and differs from the Huntington-Heberlein both in respect to the use of gypsum and the omission of the preliminary roasting. The Carmichael-Bradford process has not been threatened with litigation, so far as I am aware. The claims of its original patent read as follows[38]:

1. The process of treating mixed sulphide ores, which consists in mixing with said ores a sulphur compound of a metal of the alkaline earths, starting the reaction by heating the same, thereby oxidizing the sulphide and reducing the sulphur compound of the alkali metal, passing a current of air to oxidize the reduced sulphide compound of the metal of the alkalies preparatory to acting upon a new charge of sulphide ores, substantially as and for the purpose set forth.

2. The process of treating mixed sulphide ores, which consists in mixing calcium sulphate with said ores, starting the reaction by means of heat, thereby oxidizing the sulphide ores, liberating sulphurous-acid gas and converting the calcium sulphate into calcium sulphide and oxidizing the calcium sulphide to sulphate preparatory to treating a fresh charge of sulphide ores, substantially as and for the purpose set forth.

The process described by W. S. Bayston, of Melbourne (Australian patent No. 2862), appears to be identical with that of Savelsberg.

Irrespective of the validity of the Savelsberg and Carmichael-Bradford patents, and without attempting to minimize the ingenuity of their inventors and the importance of their discoveries, it must be conceded that the merit for the invention and introduction of lime-roasting of galena belongs to Thomas Huntington and Ferdinand Heberlein. The former is an American, and this is the only claim that the United States can make to a share in this great improvement in the metallurgy of lead. It is to be regretted, moreover, that of all the important lead-smelting countries in the world, America has been the most backward in adopting it.

The details of the three processes and the general results accomplished by them have been rather fully described in a series of articles recently published in the _Engineering and Mining Journal_. There has been, however, comparatively little discussion as to costs; and unfortunately the data available for analysis are extremely scanty, due to the secrecy with which the Huntington-Heberlein process, the most extensively exploited of the three, has been veiled. Nevertheless, I may attempt an approximate estimation of the various details, taking the Huntington-Heberlein process as the basis.

The ore, limestone and silica are crushed to pass a four-mesh screen. This is about the size to which it would be necessary to crush as preliminary to roasting in the ordinary way, wherefore the only difference in cost is the charge for crushing the limestone and silica, which in the aggregate may amount to one-sixth of the weight of the raw sulphide and may consequently add 2 to 2.5c. to the cost of treating a ton of ore. The mixing of ore and fluxes may be costly or cheap, according to the way of doing it. If done in a rational way it ought not to cost more than 10c. per ton of ore, and may come to less. The delivery of the ore from the mixing-house to the roasting furnaces ought to be done entirely by mechanical means, at insignificant cost.

The Heberlein roasting furnace, which is used in connection with the H.-H. process, is simply an improvement on the old Brunton calciner—a circular furnace, with revolving hearth. The construction of this furnace, according to American designs, is excellent. The hearth is 26 ft. in diameter; it is revolved at slow speed and requires about 1.5 h.p. A flange at the periphery of the hearth dips into sand in an annular trough, thus shutting off air from the combustion chamber, except through the ports designed for its admittance. The mechanical construction of the furnace is workmanlike, and the mechanism under the hearth is easy of access and comfortably attended to.

A 26 ft. furnace roasts about 80,000 lb. of charge per 24 hours. In dealing with an ore containing 20 to 22 per cent. of sulphur, the latter is reduced to about 10 to 11 per cent., the consumption of coal being about 22.5 per cent. of the weight of the charge. The hearth efficiency is about 150 lb. per sq. ft., which in comparison with ordinary roasting is high. The coal consumption, however, is not correspondingly low. Two furnaces can be managed by one man per 8 hour shift. On the basis of 80 tons of charge ore per 24 hours, the cost of roasting should be approximately as follows:

Labor—3 men at $2.50 $ 7.50 Coal—18 tons at $2 36.00 Power 3.35 Repairs 3.35 —————— Total $50.20 = 63c. per ton.

In the above estimate repairs have been reckoned at the same figure as is experienced with Brückner cylinders, and the cost of power has been allowed for with fair liberality. The estimated cost of 63c. per ton is comparable with the $1.10 to $1.45 per ton, which is the result of roasting in Brückner cylinders in Colorado, reducing the ore to 4.5-6 per cent. sulphur.

The Heberlein furnace is built up to considerable elevation above the ground level, externally somewhat resembling the Pearce turret furnace. This serves two purposes: (1) it affords ample room under the hearth for attention to the driving mechanism; and (2) it enables the ore to be discharged by gravity into suitable hoppers, without the construction of subterranean gangways. The ore discharges continuously from the furnace, at dull-red heat, into a brick bin, wherein it is cooled by a water-spray. Periodically a little ore is diverted into a side bin, in which it is kept hot for starting a subsequent charge in the converter.

The cooled ore is conveyed from the receiving bins at the roasting furnaces to hopper-bins above the converters. If the tramming be done by hand the cost, with labor at 25c. per hour, may be approximately 12.5c. per ton of ore, but this should be capable of considerable reduction by mechanical conveyance.

The converters are hemispherical pots of cast iron, 9 ft. in diameter at the top, and about 4 ft. in depth. They are provided with a circular, cast-iron grate, which is ¾ in. thick and 6 ft. in diameter and is set and secured horizontally in the pot. This grate is perforated with holes ¾ in. in diameter, 2 in. apart, center to center, and is similar to the Wetherill grate employed in zinc oxide manufacture. The pot itself is about 2½ in. thick at the bottom, thinning to about 1½ in. at the rim. It is supported on trunnions and is geared for convenient turning by hand. The blast pipe which enters the pot at the bottom is 6 in. in diameter.

Two roasting furnaces and six converters are rated nominally as a 90 ton plant. This rating is, however, considerably in excess of the actual capacity, at least on certain ores. The time required for desulphurization in the converter apparently depends a good deal upon the character of the ore. The six converters may be arranged in a single row, or in two rows of three in each. They are set so that the rim of the pot, when upright, is about 12 ft. above the ground level. A platform gives access to the pots. One man per shift can attend to two pots. His work consists in charging them, which is done by gravity, spreading out the charge evenly in the pot, closing any blow-holes which may develop, and at the end of the operation raising the hood (which covers the pot during the operation) and dumping the pot. The work is easy. The conditions under which it is done are comfortable, both as to temperature and atmosphere. Reports have shown a great reduction in liability to lead-poisoning in the works where the H.-H. process has been introduced.

A new charge is started by kindling a small wood or coal fire on the grate, then throwing in a few shovelfuls of hot calcines, and finally dropping in the regular charge of damp ore (plus the fluxes previously referred to). The charge is introduced in stages, successive layers being dropped in and spread out as the heat rises. At the beginning the blast is very low—about 2 oz. It is increased as the hight of the ore in the pot rises, finally attaining about 16 oz. The operation goes on quietly, the smoke rising from the surface evenly and gently, precisely as in a well-running blast furnace. While the charge is still black on top, the hand can be held with perfect comfort, inside of the hood, immediately over the ore. This explains, of course, why the volatilization of silver and lead is insignificant. There is, moreover, little or no loss of ore as dust, because the ore is introduced damp, and the passage of the air through it is at low velocity. In the interior of the charge, however, there is high temperature (evidently much higher than has been stated in some descriptions), as will be shown further on. The conditions in this respect appear to be analogous to those of the blast furnace, which, though smelting at a temperature of say 1200 deg. C. at the level of the tuyeres, suffers only a slight loss of silver and lead by volatilization.

At the end of the operation in the H.-H. pot, the charge is dull red at the top, with blow-holes, around which the ore is bright red. Imperfectly worked charges show masses of well-fused ore surrounded by masses of only partially altered ore, a condition which may be ascribed to the irregular penetration of air through the charge, affording good evidence of the important part which air plays in the process. A properly worked charge is tipped out of the pot as a solid cake, which in falling to the ground breaks into a few large pieces. As they break, it appears that the interior of the charge is bright red all through, and there is a little molten slag which runs out of cavities, presumably spots where the chemical action has been most intense. When cold, the thoroughly desulphurized material has the appearance of slag-roasted galena. Prills of metallic lead are visible in it, indicating reaction between lead sulphide and lead sulphate.

The columns of the structure supporting the pots should be of steel, since fragments of the red-hot ore dumped on the ground are likely to fall against them. To hasten the cooling of the ore, water is sometimes played on it from a hose. This is bad, since some is likely to splash into the still inverted pot, leading to cracks. The cracked pots at certain works appear to be due chiefly to this cause, in the absence of which the pots ought to last a long time, inasmuch as the conditions to which they are subjected during the blowing process are not at all severe. When the ore is sufficiently cold it is further broken up, first by driving in wedges, and finally by sledging down to pieces of orange size, or what is suitable for the blast furnace. These are forked out, leaving the fine ore, which comes largely from the top of the charge and is therefore only partially desulphurized. The fines are, therefore, re-treated with a subsequent charge. The quantity is not excessive; it may amount to 7 or 8 per cent. of the charge.

The breaking up of the desulphurized ore is one of the problems of the process, the necessity being the reduction of several large pieces of fused, or semi-fused, material weighing two or three tons each. When done by hand only, as is usually (perhaps always) the practice, the operation is rather expensive. It would appear, however, to be not a difficult matter to devise some mechanical aids for this process—perhaps to make it entirely mechanical. When done by hand, a 6-pot plant requires 6 men per shift sledging and forking. With 8-hour shifts, this is 18 men for the breaking of about 60 tons of material, which is about 3⅓ tons per man per 8 hours. With labor at 25c. per hour, the cost of breaking the fused material comes to 60c. per ton. It may be remarked, for comparison, that in breaking ore as it ordinarily comes, coarse and fine together, a good workman would normally be expected to break 5 to 5.5 tons in a shift of 8 hours.

The ordinary charge for the standard converter is about 8 tons (16,000 lb.) of an ore weighing 166 lb. per cu. ft. With a heavier ore, like a high-grade galena, the charge would weigh proportionately more. The time of working off a charge is decidedly variable. Accounts of the operation of the process in Australia tell of charge-workings in 3 to 5 hours, but this does not correspond with the results reported elsewhere, which specify times of 12 to 18 hours. Assuming an average of 16 hours, which was the record of one plant, six converters would have capacity for about 72 tons of charge per 24 hours, or about 58 tons of ore, the ratio of ore to flux being 4:1. The loss in weight of the charge corresponds substantially to the replacement of sulphur by oxygen, and the expulsion of carbon dioxide. The finished charge contains on the average from 3 to 5 per cent. sulphur. This is about the same as the result achieved in good practice in roasting lead-bearing ores in hand-worked reverberatory furnaces, but curiously the H.-H. product, in some cases at least, does not yield any matte, to speak of, in the blast furnace; the product delivered to the latter being evidently in such condition that the remaining sulphur is almost completely burned off in the blast furnace. This is an important saving effected by the process. In calculating the value of an ore, sulphur is commonly debited at the rate of 25c. per unit, which represents approximately the cost of handling and reworking the matte resulting from it. The practically complete elimination of matte-fall rendered possible by the H.-H. process may not be, however, an unmixed blessing. There may be, for example, a small formation of lead sulphide which causes trouble in the crucible and lead-well, and results in furnace difficulties and the presentation of a vexatious between-product.

It may now be attempted to summarize the cost of the converting process. Assuming the case of an ore assaying lead, 50 per cent.; iron, 15; sulphur, 22; silica, 8, and alumina, etc., 5, let it be supposed that it is to be fluxed with pure limestone and pure quartz, with the aim to make a slag containing silica, 30; ferrous oxide, 40; and lime, 20 per cent. A ton of ore will make, in round numbers, 1000 lb. of slag, and will require 344 lb. of limestone and 130 lb. of quartz, or we may say roughly one ton of flux must be added to four tons of ore, wherefore the ore will constitute 80 per cent. of the charge. In reducing the charge to 3 per cent. sulphur it will lose ultimately through expulsion of sulphur and carbon dioxide (of the limestone) about 20 per cent. in weight, wherefore the quantity of material to be smelted in the blast furnace will be practically equivalent to the raw sulphide ore in the charge for the roasting furnaces; but in the roasting furnace the charge is likely to gain weight, because of the formation of sulphates. Taking the charge, which I have assumed above, and reckoning that as it comes from the roasting furnace it will contain 10 per cent. sulphur, all in the form of sulphate, either of lead or of lime, and that the iron be entirely converted to ferric oxide, in spite of the expulsion of the carbon dioxide of the limestone and the combustion of a portion of the sulphur of the ore as sulphur dioxide, the charge will gain in weight in the ratio of 1:1.19. This, however, is too high, inasmuch as a portion of the sulphur will remain as sulphide while a portion of the iron may be as ferrous oxide. The actual gain in weight will consequently be probably not more than one-tenth. The following theoretical calculation will illustrate the changes:

─────────────────────┬──────────────────────┬───────────────────────── RAW CHARGE │ SEMI-ROASTED CHARGE │ FINISHED CHARGE ─────────────────────┼──────────────────────┼───────────────────────── {1000 lb. Pb │ {1154 lb. PbO │ { 1154 lb. PbO { 300 lb. Fe │ { 428 lb. Fe₂O₃ │ { 428 lb. Fe₂O₃(?) Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂ { 100 lb. Al₂O₃,│ { 100 lb. Al₂O₃, │ { 100 lb. Al₂O₃, etc. │ etc. │ etc. { 440 lb. S │ { 300 lb. S │ { 68 lb. S │ │ { 130 lb. SiO₂ │ { 130 lb. SiO₂ │ { 130 lb. SiO₂ Flux { 344 lb. CaCO₃ │ Flux { 193 lb. CaO │ Flux { 193 lb. CaO │ 450 lb. O │ ———— │ ———— │ ———— 2474 lb. │ 2915 lb. │ 2233 lb. │ │ │ 10% S. │ 3% S. ─────────────────────┴──────────────────────┴─────────────────────────

Ratios:

2474:2915 :: 1:1.18. 2915:2233 :: 1:0.76⅔. 2474:2233 :: 1:0.90.

It may be assumed that for every ton of charge (containing about 80 per cent. of ore) there will be 1.1 ton of material to go to the converter, and that the product of the latter will be 0.9 of the weight of the original charge of raw material.

Each converter requires 400 cu. ft. of air per minute. The blast pressure is variable, as different pots are always at different stages of the process, but assuming the maximum of 16 oz. pressure, with a blast main of sufficient diameter (at least 15 in.) and the blower reasonably near the battery of pots, the total requirement is 21 h.p. The cost of converting will be approximately as follows:

Labor, 3 foremen at $3.20 $ 9.60 “ 9 men at $2.50 22.50 Power, 21 h.p. at 30c 6.30 Supplies, repairs and renewals 5.00 —————— Total $43.40 = 60c. per ton of charge.

The cost of converting is, of course, reduced directly as the time is reduced. The above estimate is based on unfavorable conditions as to time required for working a charge.

The total cost of treatment, from the initial stage to the delivery of the desulphurized ore to the blast furnaces, will be, per 2000 lb. of charge, approximately as follows:

Crushing 1.0 ton at 10c $0.10 Mixing 1.0 ton at 10c .10 Roasting 1.0 ton at 63c .63 Delivering 1.1 ton to converters at 12c .13 Converting 1.1 ton at 60c .66 Breaking 0.9 ton at 60c .54 ——-——- Total $2.16

The cost per ton of ore will be 2.16 ÷ 0.80 = $2.70. Making allowance for the crushing of the ore, which is not ordinarily included in the cost of roasting, and possibly some overestimates, it appears that the cost of desulphurization by this method, under the conditions assumed in this paper, is rather higher than in good practice with ordinary hand-worked furnaces, but it is evident that the cost can be reduced to approximately the same figure by introduction of improvements, as for example in breaking the desulphurized ore, and by shortening the time of converting, which is possible in the case of favorable ores. The chief advantage must be, however, in the further stage of the smelting. As to this, there is the evidence that the Broken Hill Proprietary Company was able to smelt the same quantity of ore in seven furnaces, after the introduction of the Huntington-Heberlein process, that formerly required thirteen. A similar experience is reported at Friedrichshütte, Silesia.

This increase in the capacity of the blast furnace is due to three things: (1) In delivering to the furnace a charge containing a reduced percentage of fine ore, the speed of the furnace is increased, i.e., more tons of ore can be smelted per square foot of hearth area. (2) There is less roasted matte to go into the charge. (3) Under some conditions the percentage of lead in the charge can be increased, reducing the quantity of gangue that must be fluxed.

It is difficult to generalize the economy that is effected in the blast-furnace process, since this must necessarily vary within wide limits because of the difference in conditions. An increase of 60 to 100 per cent. in blast-furnace capacity does not imply a corresponding reduction in the cost of smelting. The fuel consumption per ton of ore remains the same. There is a saving in the power requirements, because the smelting can be done with a lower blast pressure; also, a saving in the cost of reworking matte. There will, moreover, be a saving in other labor, in so far as portions thereof are not already performed at the minimum cost per ton. The net result under American conditions of silver-lead smelting can only be determined closely by extensive operations. That there will be an important saving, however, there is no doubt.

The cost of smelting a ton of charge at Denver and Pueblo, exclusive of roasting and general expense, is about $2.50, of which about $0.84 is for coke and $1.66 for labor, power and supplies. General expense amounts to about $0.16 additional. If it should prove possible to smelt in a given plant 50 per cent. more ore than at present without increase in the total expense, except for coke, the saving per ton of charge would be 70c. That is not to be expected, but the half of it would be a satisfactory improvement. With respect to sulphur in the charge, the cost is commonly reckoned at 25c. per unit. As compared with a charge containing 2 per cent. of sulphur there would be a saving rising toward 50c. per ton as the maximum. It is reasonable to reckon, therefore, a possible saving of 75c. per ton of charge in silver-lead smelting, no saving in the cost of roasting, and an increase of about 3 per cent. in the extraction of lead, and perhaps 1 per cent. in the extraction of silver, as the net results of the application of the Huntington-Heberlein process in American silver-lead smelting.

On a charge averaging 12 per cent. lead and 33 oz. silver per ton, an increase of 3 per cent. in the extraction of lead and 1 per cent. in the extraction of silver would correspond to 25c. and 35c. respectively, reckoning lead at 3.5c. per lb., and silver at 60c. per oz. In this, however, it is assumed that all lead-bearing ores will be desulphurized by this process, which practically will hardly be the case. A good deal of pyrites, containing only a little lead, will doubtless continue to be roasted in Brückner cylinders, and other mechanical furnaces, which are better adapted to the purpose than are the lime-roasting pots. Moreover, a certain proportion of high-grade lead ore, which is now smelted raw, will be desulphurized outside of the furnace, at additional expense. It is comparatively simple to estimate the probable benefit of the Huntington-Heberlein process in the case of smelting works which treat principally a single class of ore, but in such works as those in Colorado and Utah, which treat a wide variety of ores, we must anticipate a combination process, and await results of experience to determine just how it will work out. It should be remarked, moreover, that my estimates do not take into account the royalty on the process, which is an actual debit, whether it be paid on a tonnage basis or be computed in the form of a lump sum for the license to its use.

However, in view of the immense tonnage of ore smelted annually for the extraction of silver and lead, it is evident that the invention of lime-roasting by Huntington and Heberlein was an improvement of the first order in the metallurgy of lead.

In the case of non-argentiferous galena, containing 65 per cent. of lead (as in southeastern Missouri), comparison may be made with the slag-roasting and blast-furnace smelting of the ore. Here, no saving in cost of roasting may be reckoned and no gain in the speed of the blast furnaces is to be anticipated. The only savings will be in the increase in the extraction of lead from 92 to 98 per cent., and the elimination of matte-roasting, which latter may be reckoned as amounting to 50c. per ton of ore. The extent of the advantage over the older method is so clearly apparent that it need not be computed any further. In comparison with the Scotch-hearth bag-house method of smelting, however, the advantage, if any, is not so certain. That method already saves 98 per cent. of the lead, and on the whole is probably as cheap in operation as the Huntington-Heberlein could be under the same conditions. The Huntington-Heberlein method has replaced the old roast-reaction method at Tarnowitz, Silesia, but the American Scotch-hearth method as practised near St. Louis is likely to survive.

A more serious competitor will be, however, the Savelsberg process, which appears to do all that the Huntington-Heberlein process does, without the preliminary roasting. Indeed, if the latter be omitted (together with its estimated expense of 63c. per ton of charge, or 79c. per ton of ore), all that has been said in this paper as to the Huntington-Heberlein process may be construed as applying to the Savelsberg. The charge is prepared in the same way, the method of operating the converters is the same, and the results of the reactions in the converters are the same. The litigation which is pending between the two interests, Messrs. Huntington and Heberlein claiming that Savelsberg infringes their patents, will be, however, a deterrent to the extension of the Savelsberg process until that matter be settled.

The Carmichael-Bradford process may be dismissed with a few words. It is similar to the Savelsberg, except that gypsum is used instead of limestone. It is somewhat more expensive because the gypsum has to be ground and calcined. The process works efficiently at Broken Hill, but it can hardly be of general application, because gypsum is likely to be too expensive, except in a few favored localities. The ability to utilize the converter gases for the manufacture of sulphuric acid will cut no great figure, save in exceptional cases, as at Broken Hill, and anyway the gases of the other processes can be utilized for the same purpose, which is in fact being done in connection with the Huntington-Heberlein process in Silesia.

The cost of desulphurizing a ton of galena concentrate by the Carmichael-Bradford process is estimated by the company controlling the patents as follows, labor being reckoned at $1.80 per eight hours, gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb.:

0.25 ton of gypsum $0.60 Dehydrating and granulating gypsum .48 Drying mixture of ore and gypsum .12 Converting 0.24 Spalling sintered material .12 0.01 ton coal .08 ——-——- Total $1.64

The value of the lime in the sintered product is credited at 12c., making the net cost $1.52 per 2240 lb. of ore.

The cost allowed for converting may be explained by the more rapid action that appears to be attained with the ores of Broken Hill than with some ores that are treated in North America, but the low figure estimated for spalling the sintered material appears to be highly doubtful.

The theory of the lime-roasting processes is not yet well established. It is recognized that the explanation offered by Huntington and Heberlein in their original patent specification is erroneous. There is no good evidence in their process, or any other, of the formation of the higher oxide of lime, which they suggest.

At the present time there are two views. In one, formulated most explicitly by Professor Borchers, there is formed in this process a plumbate of calcium, which is an active oxidizing agent. A formation of this substance was also described by Carmichael in his original patent, but he considered it to be the final product, not the active oxidizing agent.

In the other view, the lime, or limestone, serves merely as a diluent of the charge, enabling the air to obtain access to the particles of galena, without liquefaction of the latter. The oxidation of the lead sulphide is therefore effected chiefly by the air, and the process is analogous to what takes place in the bessemer converter or in the Germot process of smelting, or perhaps more closely to what might happen in an ordinary roasting furnace, provided with a porous hearth, through which the air supply would be introduced. Roasting furnaces of that design have been proposed, and in fact such a construction is now being tested for blende roasting in Kansas.

Up to the present time, the evidence is surely too incomplete to enable a definite conclusion to be reached. Some facts may, however, be stated.

There is clearly reaction to a certain extent between lead sulphide and lead sulphate, as in the reverberatory smelting furnace, because prills of metallic lead are to be observed in the lime-roasted charge.

There is a formation of sulphuric acid in the lime-roasting, upon the oxidizing effect of which Savelsberg lays considerable stress, since its action is to be observed on the iron work in which it condenses.

Calcium sulphate, which is present in all of the processes, being specifically added in the Carmichael-Bradford, evidently plays an important chemical part, because not only is the sulphur trioxide expelled from the artificial gypsum, but also it is to a certain extent expelled from the natural gypsum, which is added in the Carmichael-Bradford process; in other words, more sulphur is given off by the charge than is contained by the metallic sulphides alone.

Further evidence that lime does indeed play a chemical part in the reaction is presented by the phenomena of lime-roasting in clay dishes in the assay muffle, wherein the air is certainly not blown through the charge, which is simply exposed to superficial oxidation as in ordinary roasting.

The desulphurized charge dropped from the pot is certainly at much below the temperature of fusion, even in the interior, but we have no evidence of the precise temperature condition during the process itself.

Pyrite and even zinc blende in the ore are completely oxidized. This, at least, indicates intense atmospheric action.

The papers by Borchers,[39] Doeltz,[40] Guillemain,[41] and Hutchings[42] may profitably be studied in connection with the reactions involved in lime-roasting. The conclusion will be, however, that their precise nature has not yet been determined. In view of the great interest that has been awakened by this new departure in the metallurgy of lead, it is to be expected that much experimental work will be devoted to it, which will throw light upon its principles, and possibly develop it from a mere process of desulphurization into one which will yield a final product in a single operation.