Chapter IX: Part V: Lime-Roasting of Galena (2)

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

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

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

METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE[24]

BY F. O. DOELTZ

(January 27, 1906)

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

PbS + CaSO₄ = PbSO₄ + CaS.

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

PbS + CaSO₄ = PbSO₄ + CaS

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

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

PbO + SO₂ + O = PbSO₄.

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

THE HUNTINGTON-HEBERLEIN PROCESS

BY DONALD CLARK

(October 20, 1904)

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

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

CaO + PbS = CaS + PbO.

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

CaSO₄ + PbO = CaPbO₃ + SO₂.

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

CaSO₄ + PbS + O = CaPbO₃ + SO₂.

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

ZnS + 3ZnSO₄ = 4ZnO + 4SO₂.

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

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

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

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

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

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

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

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

THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHÜTTE[29]

BY A. BIERNBAUM

(September 2, 1905)

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

BY A. BIERNBAUM

(October 14, 1905)

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

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

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

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

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

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

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

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

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

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

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

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

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