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: