Chapter C: THE MACDOUGAL TYPE.—The most important form of modern roaster (5)

Deficiency of silica also results in the production of _over-fire_, owing to the fact that the air blast, being unable to bessemerise any more iron sulphide at the tuyere zone, passes to the higher portions of the furnace and gradually roasts the ore there, thus consuming the sulphide fuel of the furnace which might otherwise be most effectively used for bessemerising in the tuyere zone. This over-fire, resulting from the heat of roasting which is given out in the upper part of the furnace, is very disadvantageous in true pyritic smelting, and successful control of the process depends on using up the whole of an adequate air supply at the bessemerising zone, and on supplying sufficient siliceous flux to combine at once with the whole of the iron oxide produced. For fluxing purposes it is only the free silica in the charge which is effective, since any silica existing as silicate is already in a state of combination and thus is not free to act as flux. The combined silica, except for its adding to the fusibility of the charge by admixture, is very disadvantageous, consuming heat and space, diluting the reaction intensities by presenting an inert substance among the active constituents, and increasing the quantity of slag which requires to be melted.

The three requirements—iron sulphide, oxygen supply, and fluxing silica—thus bear an intimate relationship to one another in true pyritic smelting, and alteration of any one factor requires simultaneous adjustment of the others for the production of the same grade of matte and slag. The speed and degree of oxidation primarily depend on the air supply. The more iron burnt up, the greater is the heat production and the higher the temperature at the tuyere zone, and since the more basic slags are known to have the higher formation temperatures, the basicity of the slags increases with the speed of oxidation and consequent concentration.

Ores suitable for true pyritic smelting are not commonly met with in practice, and the presence of earthy bases other than iron is not desirable. Whilst the advantages of polybasic slags from the point of view of reduced formation temperature, increased fusibility and liquidity are very marked in ordinary smelting practice, their presence is not so advantageous in true pyritic smelting, since they consume silica which is required for the iron oxide at the instant of formation, and thus tend to decrease the speed of oxidation and concentration. Polybasic slags have a lower formation temperature, and in consequence the production of the highly ferruginous slags of high formation temperature which it is desired to make by the oxidation of as much iron as possible is retarded. In addition, the presence of other earthy bases in the charge dilutes its fuel value; they may even consume valuable heat by requiring decomposition, as in the case of carbonates. These considerations are not so important in partial pyritic smelting, where the required heat balance can be adjusted by coke.

_The Advantages of Pyritic Smelting._

(1) The possibility of direct and immediate treatment of highly pyritic raw ore in the blast furnace, thus saving all the costs of preliminary treatment and handling.

(2) The saving of the costs of roasting heavy sulphides.

In former smelting practice, high sulphide contents in a copper ore were particularly disadvantageous, since the higher the sulphur contents of the charge the lower was the grade of the resulting matte, when smelted directly in the blast furnace. In consequence, the higher sulphur content necessitated a more complete roasting of the ore in order to ensure a high-grade matte on smelting.

With pyritic smelting the conditions are completely reversed, and the charge becomes more suitable for direct furnace treatment as its sulphide contents increase, so that the most suitable ores for pyritic smelting are those in which the greatest saving is effected by their not requiring a preliminary roasting operation.

As has been already indicated, this saving includes labour, plant, handling, time, and interest on capital tied up in the roast yards, as well as the avoiding of all the mechanical and other losses connected with such preliminary treatment. Thus at Ducktown, Tennessee, the material economies effected by the substitution of pyritic smelting for the processes involving preliminary roasting amounted to no less than 3 to 4 cents per pound of copper produced, in addition to the later advantages derived from the recovery of values from the gases, and from the improved conditions of life in the district.

(3) The cost of coke is saved.

Fuel is one of the main items of expense in blast-furnace smelting, and by the substitution of the cost-free natural-sulphide fuel for coke, the proportion of the latter required on the charge is reduced from the 9 to 10 per cent. formerly employed with roasted materials to about 3 to 5 per cent., and in certain special cases to very much smaller amounts.

_Difficulties of the Process._—That the technical difficulties in applying the process on a practical scale are considerable, under present conditions of working, will be understood from the nature of the operations.

(1) The pyritic process works on a narrow margin of heat, and allows of but little flexibility in the conditions of working, since there are few factors which can be altered should difficulties in operating arise, as compared with the circumstances when a free use of supplementary carbonaceous fuel may be employed. The only source of heat energy at the smelting zone is in the sulphide charge itself, and small variations in the working conditions may readily disturb the delicate equilibrium upon which successful working depends. Irregularities, stoppages, and variations in grade of matte may therefore arise, unless the operations are regulated with exceeding watchfulness. In true pyritic smelting the employment of coke for restoring the balance or for producing heat required at the tuyere zone is not permissible or practicable, since, as will be indicated later, such coke addition would altogether destroy the equilibrium in the process; the grade of matte and the composition of slag would be altered, the reactions disturbed, and a restoration to normal pyritic smelting conditions rendered almost impossible.

Difficulties in operation have therefore to be overcome along the lines of pyritic action—that is, in the further adjustment and manipulation of blast, sulphide or silica supply, or in charging methods, etc.—and in practice such careful “doctoring” is resorted to when the furnace shows signs of working unsatisfactorily.

It is very often possible by such careful attention to gradually bring a furnace back to smooth running. It occasionally happens, however, that the conditions gradually become worse, and the furnace commences to show signs of “gobbing.” This is indicated at the top of the charge by the formation of crusts round the side and end walls, whilst from the slag spout below, there issues a much reduced quantity of thick siliceous slag, together with an abundant stream of thin low-grade matte. The furnace gradually ceases running, and it becomes necessary to stop its working, to take down the furnace jackets, bar out the debris, and restart operations. This is usually not so objectionable a procedure as it might appear, and indeed, within certain definite limits, such a course may economically be sound policy. In the modern operation of pyritic practice it often pays better to risk the occasional gobbing up of a furnace and clear out the debris, than to work with so large a quantity of coke as would avoid such a necessity. Not only is the modern furnace so designed and constructed as to entail but comparatively little trouble in cleaning out in this manner, but such practice, even if temporarily a necessary evil, may, in places where coke is expensive, and where conditions for pyritic smelting are otherwise favourable, be, within certain definite limits, actually the most profitable. It is by the taking of these risks, combined with further experiment and working experience in manipulation, such as in charging methods, blast conditions, and the height and distribution of charges, etc., that the ultimate continuous and successful working at still lower costs may be attained and the true pyritic process be worked as continuously as ordinary smelting practice. Short campaigns are not, therefore, unusual under the present conditions of true pyritic smelting, and the cleaning out of the generally fairly loose debris is accomplished with moderate ease, from 24 to 36 hours being the usual time required to take down, clean out, and restart a furnace, whilst the cost of such an operation (chiefly in labour) is not, under the circumstances, excessive. At Tennessee, hard driving and short campaigns result in lower costs and greater tonnage.

(2) The composition of the slag often prevents high concentration.

It has been indicated that the thermal conditions in the bessemerising zone of the pyritic furnace tend to the production of highly basic slags, which, though hot and limpid, are characterised by high density. Such slags are not conducive to good settling and separation of mattes, and they tend to occasion high copper losses, because—

(_a_) The difference in density of slag and matte is
not sufficiently great.
(_b_) The solubility of sulphides in the slag increases
with its basicity.

The greater the concentration effected by the smelting operation, the higher is the grade of the matte produced; at the same time, the actual weight of matte is smaller. On the other hand, since more iron is oxidised from the charge and slagged off, the quantity of slag produced increases proportionately. Contrasting then, the likely losses of copper which would result from the association of a small quantity of high-grade matte with much slag, compared with those resulting from the association of a considerable quantity of low-grade matte in the presence of but little slag, the former condition is obviously the more productive of heavy loss, for not only will many more shots of matte be held in suspension, but each shot of high-grade matte represents a larger quantity of copper.

It is found in practice that it is most economical to make a fairly low-grade matte on the first or “green-ore” smelting, and to re-concentrate this matte pyritically up to converter grade by a second smelting operation. The extra cost of casting the low-grade matte, of breaking up, rehandling it, and resmelting, with all the extra charges on capital, etc., involved, is less than the losses which would be incurred if higher-grade converter matte were made at the first smelting, although there is no difficulty at all in producing such mattes so far as the actual furnace operations are concerned. It is entirely a question of the slag losses involved.

Under ordinary smelting conditions (not truly pyritic), when using some coke for fuel, it would be readily possible to alter the density of the slag by adding suitable constituents, such as limestone or additional silica, but in pyritic smelting this is not practicable. The furnace chooses to make at the tuyere zone its own slag, and that a highly basic one. High concentration and a slag low in iron content cannot be obtained together in true pyritic smelting, since high concentration means rapid oxidation of iron sulphide, and this necessitates high temperature and produces a highly ferruginous slag in consequence. Additional silica added to the charge could not alter the slag composition markedly and still yield the same grade of matte. The silica content of the slag depends on the temperature at the tuyere zone, and this is governed by the rate of oxidation of the iron sulphide. If the slag is to be more siliceous it must be produced at a lower temperature, which would be obtained by oxidising the iron less rapidly. This would lead to the production of low-grade matte, and probably would so reduce the furnace activity that there would not be sufficient heat to keep the slag molten.

If extra silica be added to the charge, it would probably be unattacked unless more iron were oxidised in order to flux it off. In such a case the blast would have to be increased in order to produce iron oxide more rapidly, the temperature would in consequence be raised, a still more basic slag would be produced in larger quantity, whilst the matte would be increased in grade and reduced proportionally in weight.

The addition of sufficient lime to the charge, in order to produce a sufficiently low-gravity slag, is also impracticable in true pyritic work, because—

(_a_) The extra lime consumes silica, and interferes with the desired reactions at the bessemerising zone, tending to lower the concentration. It also absorbs heat.

Lime has a very powerful affinity for silica, more strongly marked than that of iron oxide, its replacing value is higher, its more siliceous silicates are readily formed and they have a lower formation temperature, all of which factors tend to an undue consumption of silica which is urgently required by the iron if the rate of oxidation is to be maintained. The marked tendency for lime in the charge to consume the silica tends to retard the oxidation of the iron sulphide, which proceeds most satisfactorily when free silica is available for the nascent iron-oxide, and in consequence concentration is decreased and the heating effect in the furnace reduced. In addition, the larger bulk of calcareous slag carries considerable heat from the smelting zone of the furnace. Lime silicates and the polybasic lime slags have a markedly lower formation temperature than the normal ferruginous slags of true pyritic smelting, they are hence formed readily without requiring so much oxidation activity at the tuyere zone. In consequence less iron is oxidised, and the resulting concentration in the matte is proportionately reduced.

(_b_) The lime is introduced in the form of limestone, and the carbon dioxide liberated from this material in the furnace is found to have a deleterious effect on the furnace gases if the manufacture of sulphuric acid from them is intended—this being a consideration of great economic importance in connection with many modern pyritic smelters.

Hence, in practice, pyritic smelting is at present generally conducted in two stages for the production of a matte of 30, 40, or 50 per cent. converter grade. The “green ore-matte,” or first matte, runs usually from 8 to 13 or 14 per cent. of copper, depending upon the copper ore available, which is usually very low grade—2 to 3 per cent. copper contents; the second or concentrated matte assays 28 to 40 per cent. copper. Special care is taken to ensure good settling of matte from the basic and irony slags, and by these means the copper losses in the slags are reduced to the comparatively moderate proportions associated with normal practice.

It does not appear improbable that with the developments of basic converter practice, involving eventually the continuous converting of low-grade mattes, the necessity for this second pyritic smelting and re-concentration may be avoided. The removal of this feature from pyritic smelting practice would add enormously to the potential economies arising from the method.

In spite of the difficulties connected with the process, as detailed above, the method has proved itself an exceedingly profitable one on a large scale, and the experience of the companies financially interested, as well as the opinions of managers of the plants in practical operation, leave no doubt as to the economic success of this application of scientific principles to a practical problem on a very extended scale.

SPECIAL FEATURES OF PYRITIC SMELTING.—Several points of particular interest have given rise to much discussion in connection with pyritic smelting practice. These include the question of the coke proportion required on the charge, and the advisability or otherwise of employing heated blast for the furnace.

_Coke Proportion._—Whilst ideal pyritic practice involves the entire absence of supplementary carbonaceous fuel, it has not been found practicable, up to the present, to ensure satisfactory working over any reasonable period of time, unless a minimum of about 1·25 per cent. of coke is incorporated with the charge. The function of this coke has been a matter of much speculation, but the investigations of Sticht already referred to, now permit the tracing, with some considerable accuracy, of its function and of its action in the furnace.

It is found that in true pyritic smelting the coke does not reach the bessemerising zone at all, but that it is completely consumed in the regions above this point. It is, moreover, not burned by the oxygen of the air, none of which exists above the tuyere zone, since all this oxygen is consumed by the combustion of the sulphide. It appears that the coke is oxidised by the SO_{2} which results from this sulphide combustion. The examination and analysis of samples of the gases withdrawn from different parts of the furnace have confirmed this view, and have elucidated the probable reason for the apparent necessity of a certain small proportion of coke in the process, under the present conditions of working. The heat generated from the oxidation of the coke by the SO_{2} is of much value in preheating the materials of the charge for the removal of excess sulphur and the liquation of the sulphides. The amount of heat which is available for this operation is small, being practically all derived from that carried upwards by the hot gases leaving the smelting zone, and none is obtainable by the usual processes of coke or sulphide oxidation in the upper regions of the furnace, since no available oxygen is believed to get past the bessemerising zone and reach these upper areas. It is indeed necessary for the success of pyritic smelting that such oxidation or roasting of sulphides in the upper part of the furnace should be prevented, since every available particle of iron sulphide is required for heat production at the smelting zone, by its combustion there, and any oxidation elsewhere not only deprives this zone of fuel, but spreads the heat over too wide an area for sufficiently intense combustion.

Thus, by supplying an additional amount of heat to the upper parts of the furnace, where heat is needed to assist in the preparation and liquation of the sulphides, the extra coke, in being oxidised by the SO_{2} without robbing the tuyere zone of fuel or air, just fulfils its useful purpose at the required place, in such a way as to keep the smelting operation running smoothly.

The presence of more coke than is absolutely necessary for the fulfilment of this purpose is, in addition to its extra cost, of no advantage, and in true pyritic smelting none should reach the tuyere zone, since it introduces a reducing influence where the most marked oxidising effect is required. By consuming oxygen for its combustion, it deprives the iron sulphide of this material, less iron is, therefore, oxidised, and the matte is consequently increased in quantity and lowered in grade, whilst the amount of iron carried into the slag is decreased.

1·25 per cent. of coke is about the minimum quantity with which it is found practicable to maintain satisfactory working of the furnace under present conditions, 0·5 per cent. has been worked with occasionally, and none at all over certain short periods of time. The average quantity employed is from 2 to 3 per cent., and when about 5 per cent. is used, coke reaches the tuyere zone and the process ceases to be truly pyritic—the reactions and smelting conditions become entirely changed.

It does not seem unlikely that, as knowledge of these conditions increases and as the mechanism of the process becomes more generally understood, modifications in furnace design and blast conditions may lead to the successful operating of the pyritic process entirely independent of the use of coke fuel.

_Heating of the Blast._—For true pyritic smelting it has been shown in practice that the use of heated blast possesses no advantages; many smelters operating the process have tested the effects, and have usually given the method up, whilst the work of Sticht and Peters affords valuable evidence and close argument as to the reasons for its unsuitability. Success in true pyritic working depends upon the intensity of oxidation of the sulphides, and upon the localisation of the resulting heat at the narrow bessemerising zone situated just above the tuyeres. The greater the quantity of iron which is there oxidised per minute, the better is the concentration, the greater is the smelting and fluxing intensity and the higher is the resulting temperature. Since the character and composition of the slag vary in accordance with these conditions, depending largely upon the temperature in the tuyere zone, the furnace works most rapidly and satisfactorily when slags of high formation temperature are being produced. These can only be formed if much iron is being oxidised, because iron is the chief fuel in the process. The addition of extra heat by warming the blast appears to allow of the formation of silicate slags possessing a lower formation temperature, such slags are less basic, and consequently less iron need be oxidised and slagged off per minute in order to produce them. Less iron sulphide fuel is, therefore, burned, and the reaction intensity at the tuyere zone is reduced, so that the necessary heat margin for satisfactory smelting may not be attained. The extra heat carried in by the warmed blast may not be sufficient to compensate for that which is lost owing to this decrease in oxidation intensity; the furnace consequently tends to work cold, whilst the excess air supply leads to the production of over-fire, by the oxidising of sulphides higher up in the charge.

These features are specially interesting, as they afford one of the most marked distinctions between true and partial pyritic smelting. In the latter process, the fuel value in the adjustable supply of coke at the tuyeres allows of the ready production of any extra heat which might be required. The slag composition is, in consequence, more independent of the furnace conditions, since the heat required for the smelting operation does not depend so much on the formation of slag of any particular composition. Sufficient heat is always obtainable by coke additions when smelting for any special slag which may be desired. Neither is localisation of the heat at the narrow tuyere zone so essential in partial pyritic smelting. Warm blast produces a greater combustion intensity when employed in oxidising carbon, so that it may present advantages, both economic and operative, in partial pyritic work, whereas it is distinctly disadvantageous in the true pyritic method.

=Pyritic Smelting Practice in Tennessee.=—The pyritic process is operated in Tennessee at two smelters; that at Copperhill under the Tennessee Copper Company, and at Isabella by the Ducktown Sulphur, Copper and Iron Co. The ore averages from 2 to about 2½ per cent. copper, 31 to 37 per cent. iron, 20 to 30 per cent. sulphur, 10 to 25 per cent. silica, the remainder being earths, including lime about 6 per cent., magnesia 2 per cent., zinc 2 per cent., and alumina—_i.e._, a heavy sulphide ore with but little excess of free silica available for the fluxing of iron.

_Copperhill._—The process is conducted very much according to the principles just considered. The Copperhill plant operates seven furnaces of the ordinary rectangular water-jacketed type—the general features of furnace design being little different at present, whether true or partial pyritic practice be conducted. Several important devices in detail have been introduced with successful results, and the management is distinguished for its pioneer work and experimental enterprise in connection with the process. The furnaces were formerly all 56 inches wide; three of them are 180 inches long, the other four being 270 inches. The height of charge is from 10 to 12 feet, the capacity of the smaller furnaces 375 to 400 tons of charge daily, and a blast of 19,000 cubic feet of air per minute at 50 ozs. pressure is supplied to each. The larger furnaces have a capacity of 500 to 600 tons daily. Many trials have been made to determine the best shape for the water-jacketed sections, both broad and narrow panels having been employed. In one of the furnaces, curved end-jackets were tried, with the object of lessening the production of crusts which tend to form at the corners, owing to coldness and reduced furnace activity at these points. The advantages expected have not been realised, the tendency to crusting has not been lessened, and although barring has been rendered easier, the disadvantages of rounded corner-jackets and their greatly increased cost of construction outweigh their advantages, and their use has now been given up.

An important modification in the form of the tuyeres has been introduced with the object of furnishing more effectively the necessary large volume of air at suitable pressure, and of increasing the efficiency at the tuyere zone. Instead of supplying the air to the furnace at a number of separated points, it was felt that the closer these could be brought together the better. A narrow slot all round the furnace for air admission has been held to be the most perfect method, but hitherto it has been thought impracticable, though a recent form of furnace (not at this plant) has been devised on this system. The improvement here has been the use of slotted tuyeres, 12 inches long by 4 inches wide, each of which replaces two of the older tuyeres of 3¼ inches diameter. These have proved very successful, the furnace thus equipped handling a much larger tonnage, and it has been decided to adopt the new form on all the furnaces.

Charging is by side-dumping V-shaped cars, and great care is taken in the handling and distribution of the charges. The furnaces are fitted with tops of special design, and with elaborate dust-catching devices which have been the subject of long and numerous experiments; the special purpose being to allow the taking off of the gases below the feed-floor, and to reduce the height of the superstructure to the smallest possible proportions, so as to prevent excessive dilution (by air) of the furnace gases, which are used for sulphuric acid manufacture. The furnace tops were originally of the standard form—brick walls supported by steel frame-work. It was, however, necessary to damper down the flues in order to obtain sufficient pressure to force the gases through the Glover towers, and the heat has caused the steel work to warp badly. A low top was tried, using a brick-lined flue at the end for taking off the gases below the feed-floor. This was found to be good for charge-dumping and general convenience, but it allowed the escape of too much smoke and flames, which greatly interfered with the furnace manipulation. In consequence the tubular top was used, gradually raised until a suitable height was reached. This form has been described on p. 140.

The present practice at Copperhill is to smelt the ore pyritically for a 9 to 10 per cent. matte, passing the products through the 16-foot settlers which are now lined with siliceous copper ore, then tapping the matte into ladles which empty it into beds of flue-dust. Alternate layers of matte and dust are thus incorporated, and yield a porous material convenient for the concentrating pyritic smelt which follows. This re-concentration is now conducted in a furnace narrowed to 44 inches, which has been found specially well suited for the work; the furnace runs fast, smelting sometimes over 800 tons of charge per day. The system of working is that of hard driving so long as the furnace smelts rapidly. As soon as it slows down, the furnace is tapped out and started afresh. The re-concentrating charge contains some limestone in order to reduce the copper losses in the slag, the saving effected by this feature being equivalent to 2 lbs. of copper per ton of ore smelted. The resulting matte is bessemerised.

The furnace gases are utilised for sulphuric acid manufacture, the acid plant being the largest in the world, with an ultimate capacity of 400 tons per day.

_Ducktown._—It was at the Ducktown Company’s smelter that the first work on pyritic smelting in the district was carried out, and the successful development of the process generally, owes much to Freeland’s early pioneer work, the remarkable results of which led Parke-Channing to adopt the process at the Copperhill plant.

TABLE XI.—TYPICAL CHARGING TABLES AT PYRITIC SMELTER.

────────────────────────────┬──────────────────────────────────────
│ B.F. NO. 3—NIGHT SHIFT.
│ Typical Green Ore Charges.
├───────┬───────┬───────┬───────┬──────
│ I. │ II. │ III. │ IV. │
────────────────────────────┼───────┼───────┼───────┼───────┼──────
│ Lbs. │ Lbs. │ Lbs. │ Lbs. │
Coke, │ 180 │ 240 │ 240 │ 400 │ ..
Ore A., │ 5,000 │ .. │ .. │ .. │ ..
Ore B., │ .. │ 5,000 │ .. │ .. │ ..
Ore C., │ .. │ .. │ 5,000 │ .. │ ..
Slag, │ .. │ .. │ .. │ 4,000 │ ..
Lime rock, │ .. │ .. │ .. │ .. │ ..
Green ore (low grade) matte,│ .. │ .. │ .. │ .. │ ..
Flue-dust, │ .. │ .. │ .. │ .. │ ..
Quartz (for flux), │ 950 │ .. │ .. │ .. │ ..
────────────────────────────┼───────┼───────┼───────┼───────┼──────
Total weight of charge, │ 6,130 │ 5,240 │ 5,240 │ 4,400 │ ..
────────────────────────────┼───────┼───────┼───────┼───────┼──────
│ No. of│ No. of│ No. of│ No. of│ Total
_Hours of Charging_— │Charges│Charges│Charges│Charges│per Hr
6–7, │ 2 │ 2 │ .. │ .. │ 4
7–8, │ 2 │ .. │ 2 │ 2 │ 6
8–9, │ 2 │ 2 │ .. │ .. │ 4
9–10, │ 2, 2 │ .. │ 2, 2 │ .. │ 8
10–11, │ .. │ 2 │ .. │ 2 │ 4
11–12, │ 2 │ .. │ 2 │ .. │ 4
12–1, │ 2, 2 │ 2 │ .. │ .. │ 6
1–2, │ 2 │ .. │ 2 │ 2 │ 6
2–3, │ 2 │ 2 │ .. │ .. │ 4
3–4, │ 2 │ .. │ 2 │ .. │ 4
4–5, │ 2 │ .. │ .. │ 2 │ 4
5–6, │ 2, 2 │ 2 │ 2 │ .. │ 8
────────────────────────────┼───────┼───────┼───────┼───────┼──────
Total No. of charges daily,│ 28 │ 12 │ 14 │ 8 │ 62
────────────────────────────┴───────┴───────┴───────┴───────┴──────

─────────────────────┬───────────────────────────────────────────────
│ B.F. NO. 5—DAY SHIFT.
│ Typical Concentrating Charges.
├───────────────────┬──────────┬─────────┬──────
│ I. │ II. │ III. │
─────────────────────┼───────────────────┼──────────┼─────────┼──────
│ Lbs. │ Lbs. │ Lbs. │
Coke, │ 150 │ 400 │extra 700│ ..
Ore A., │ .. │ .. │ .. │ ..
Ore B., │ .. │ .. │ .. │ ..
Ore C., │ .. │ .. │ .. │ ..
Slag, │ .. │ 4,000 │ .. │ ..
Lime rock, │ 700 │ .. │ .. │ ..
Green ore (low grade)│ │ │ │
matte, │ 3,500 │ .. │ .. │ ..
Flue-dust, │ .. │ .. │ .. │ ..
Quartz (for flux), │ 1,050 │ .. │ .. │ ..
─────────────────────┼───────────────────┼──────────┼─────────┼──────
Total weight of │ │ │ │
charge, │ 5,400 │ 4,400 │ 700 │ ..
─────────────────────┼───────────────────┼──────────┼─────────┼──────
│ No. of │ No. of │ │Total
_Hours of Charging_— │ Charges │ Charges │ │per Hr
6–7, │2, 2, 2, 2, 2, 2, 2│ .. │ .. │ 14
7–8, │ 2, 2, 2, 2 │ 2 │ .. │ 10
8–9, │ 2, 2, 2, 2, 2 │ .. │ .. │ 10
9–10, │ 2, 2, 2, 2, 2 │ .. │ .. │ 10
10–11, │ 2, 2 │2, 2, 2, 2│ .. │ 12
11–12, │ 2, 2, 2, 2, 2 │ .. │ .. │ 10
12–1, │ 2, 2, 2, 2, 2 │ .. │ .. │ 10
1–2, │ 2, 2, 2, 2, 2 │ .. │ .. │ 10
2–3, │ 2, 2, 2 │ 2,2 │ .. │ 10
3–4, │ 2, 2, 2, 2, 2, 2 │ .. │ .. │ 12
4–5, │ 2, 2, 2, 2, 2 │ .. │ .. │ 10
5–6, │ 2, 2, 2 │2, 2, 2, 2│ .. │ 14
─────────────────────┼───────────────────┼──────────┼─────────┼──────
Total No. of charges│ │ │ │
daily, │ 120 │ 12 │ .. │132
─────────────────────┴───────────────────┴──────────┴─────────┴──────

It will be observed that the concentrating furnace works twice as quickly as the green ore matting furnace, and hence one furnace only is required for the concentration of the matte product from two of the matting furnaces.

The Isabella smelter comprises two furnaces of moderate size, 17 feet by 3 feet 4 inches at the tuyeres, having a joint capacity of 500 to 600 tons daily. The furnaces are about 9 feet high, and are water-cooled. Air at only 20 to 30 ozs. pressure is supplied through 3-inch tuyeres. The smelting scheme is somewhat analogous to that adopted at Copperhill, the first smelting producing a 20 per cent. copper matte from the 2 per cent. ore, whilst the re-concentration results in a converter-grade matte assaying 50 per cent. The coke proportions are somewhat similar to those used at Copperhill, being 5·0 per cent. for the first smelting, and 3·5 per cent. for the second. The furnace management at this small plant is exceedingly efficient, and the campaigns are long, it being claimed that the furnace operations have never had to be completely stopped on account of crusting or gobbing. This is held to be due to the results of special care in feeding and charge distribution, the ingenious Freeland charger already described being used. The charge is kept low (6 to 8 feet above the tuyeres), and is evenly red hot all through. The slags assay 35 to 36 per cent. silica, 38·8 per cent. iron, and 8·0 per cent. lime—with moderate copper losses. The annual output is equivalent to about 3,000 tons of metallic copper. An acid-making plant is also attached to these works.

=The Manufacture of Sulphuric Acid from Pyritic Furnace Gases.=—Modern legislative requirements make severe demands upon the managements of smelter-works where sulphury ores are dealt with, by reason of the disastrous effects of the sulphurous gases upon the conditions of life generally in the vicinity. In other cases, litigation by neighbouring farmers and others impose restrictions on the amount and character of the gases which the smelters are allowed to emit from their furnace stacks. So serious has the problem become that several smelters have had to cease operations altogether, others have been mulcted in enormous costs by law suits, by claims for compensation, or by the installation of plant and processes which they have been compelled to adopt for dealing with the gases. These matters have become subjects of historical importance in the development of smelter practice.

As has been the case in analogous circumstances elsewhere, when interference with the uncontrolled dispersion of then-considered waste products has often proved of ultimate benefit and a source of much profit to their producers, the enforced treatment of highly sulphurous furnace gases has in several instances resulted in considerable gain to the copper smelters.

Among the methods which are at present economically practicable for dealing with the smelter gases, those of dilution, and of utilisation for acid manufacture are the most important.

The considerations which decide the best course of treatment depend on the numerous economic and local factors which are always of such prime importance in connection with industrial undertakings demanding large capital outlay. The installation of a plant for making sulphuric acid from the gases largely depends on—

(_a_) The technical factor as to whether the composition
of the gases is suitable for the making of acid.
(_b_) The economic factor as to whether such acid can
be put upon the market on a satisfactory basis.

(_a_) For the successful operation of acid-making plant, as at present developed, it is necessary that the proportions of sulphur dioxide in the gases shall not fall below a certain minimum, and further, that the gases shall not contain more than certain limiting proportions of other interfering constituents, such as, for instance, CO_{2}. It is for this reason that the blast furnace operating the true pyritic process furnishes gases of the type most suitable for acid manufacture, since by this process the sulphur-dioxide is obtained in the gases in the most concentrated and the least contaminated form possible under smelting conditions. Even under these circumstances the gases are not in the least of an ideal composition for treatment, owing to their dilution with nitrogen, etc., and the development of the acid-making plants and processes adopted for the successful utilisation of copper blast-furnace gases furnishes a record covering many years of very slow and costly experiment, marked by many preliminary failures and disappointments. These difficulties have now been overcome, as the working of the successful plants attached to both of the Tennessee copper smelters affords conclusive proof, and the sulphur which formerly cost money to dissipate by roasting, now not only acts as fuel, but furnishes a very profitable bye-product.

The requirements for the gases are chiefly the presence of sufficient SO_{2} and oxygen, and of as little CO_{2} as possible—factors which depend largely on the proportions of sulphide in the charge. The gas for the acid plant must be supplied in regular and continuous amount, at a specific temperature, and this calls for special care in the smelting operation, furnace manipulation and blast supply, supplementary air admission, etc.

About 3·5 to 4 per cent. of SO_{2} in the gases delivered at the chambers is the minimum proportion for satisfactory working; CO_{2} should not exceed about 5 per cent., and about 6·0 per cent. or more of oxygen is also necessary.

(_b_) In addition to the capital charges involved in the acid-making installation and the costs of adapting the furnace plant and operations to the process, the problem of putting the acid upon the market on a satisfactory economic basis is important, particularly in view of the competition from other sources. The districts which offer a consuming area for the large and regular supply of acid from the smelters are not unlimited in number, and are probably readily accessible to other sources. In view of the costs of production, the distance of the smelter from the market is a serious consideration, since freight charges on sulphuric acid are high, involving special regulations with respect to the form of car and conditions of traffic, and they may readily exceed all possible profits resulting from the sale of the product.

In Tennessee the companies were forced to install acid plants. That at Copperhill is the largest in the world; commenced in 1906, acid manufacture began about two years later, after much experimenting, and further units have gradually been added. The plant now includes two Glover towers, 30 feet across and 50 feet high, 64 cooling chambers about 11 feet × 11 feet × 70 feet high, eight cooling chambers 11 feet × 24 feet × 70 feet high, twelve old chambers 50 feet × 50 feet × 70 feet, six new chambers 50 feet × 50 feet × 75 feet, eight new chambers 23 feet × 50 feet × 80 feet, eight Gay-Lussac towers, with complementary tanks, etc.—producing at the rate of 168,000 tons of 60° B. acid per annum.

The Ducktown Company’s plant was installed in record time, and, like the Copperhill plant, comprises elaborate dust chambers and flues, with Glover and Gay-Lussac towers of special design and construction, and enormous acid-making chambers with complex valves and fittings. The plant is designed to produce about 160 tons of 60° B. acid daily. The analysis of the gases supplied to the towers varied during the early working of the plant; under fairly normal conditions the average analysis of the gases delivered is SO_{2} 3·5 per cent., CO_{2} 3·5 per cent., SO_{3} trace; the oxygen in the mixture being about 8·0 per cent. The temperature is also apt to vary. Full details on these points are not yet available for general service.

The management of both companies have been successful in obtaining particularly satisfactory contracts for the purchase of their acid by fertiliser corporations.

References.

Peters, E. D., “Principles” and “Practice of Copper Smelting.”

_Blast-furnace Manipulation._
Shelby, Geo. F., “Alumina in Blast-Furnace Slags.”
_Eng. and Min. Journ._, 1908.
Offerhaus, C., “Copper Blast-Furnace Smelting at Anaconda.”
_Eng. and Min. Journ._, 1908, Aug. 7, pp. 243–250.
Sackett, B. L., “The Granby Smelter Equipment.”
_Mines and Minerals_, 1910, April, p. 524.
“Operations of the Tennessee Copper Company.”
_Official Annual Reports of the General Manage.r_
Walker, A. L., “The Metallurgy of Copper in 1910.”
_Eng. and Min. Journ._, 1911, Jan. 7, p. 39.
Austin, L. S., “Review of Metallurgy in 1910.”
_Met. and Chem. Ind._, 1911, Jan. 11, p. 40.
Rice, Claude T., “Handling Copper Smelting Gases.”
_Eng. and Min. Journ._, 1911, Mar. 25, p. 614.
“Cottrell’s Fume Smelter.” _Min. and Scient. Press_,
Aug. 26, Sept. 2, 1911.
Herrick, R. L., “Boston and Montana Co.’s Smelter at Great Falls.”
_Mines and Minerals_, 1909, Dec., p. 257.
Harvard, F. T., “Condensation of Fume and Neutralisation of
Furnace Gases.” _Bull. Amer. Inst. Min. Eng._,
No. 44, 1910, Aug.
“Mineral Industry.” Annual.

_Pyritic Smelting._
Holway, John, “A new Application of Bessemer’s Method of
Rapid Oxidation, by which Sulphides are utilised for Fuel.”
_Journ. Society of Arts_, Feb. 1879.
Rickard, T. A., “Pyrite Smelting.”
Sticht, Robert, “Ueber das Wesens des Pyrites Verfahrens.”
_Metallurgie_, Nov. 22, Dec. 8, 1906.
Wintle and Alabaster, “Pyritic Smelting.”
_Trans. Inst. Min. and Met._, 1906, vol. xv., p. 269.
Nicholls, F. S., “Pyrite Smelting in Tilt Cove, Newfoundland.”
_Eng. and Min. Journ._, 1908, Sept. 5, p. 462.
Wright, L. T., 44 “Pyritic Smelting without Coke.”
_Min. and Scient. Press_, 1906, Sept. 29.

_Sulphuric Acid Manufacture._
Falding, F. J., and Channing, J. P.,
“Pyrite Smelting and Sulphuric Acid Manufacture.”
_Eng. and Min. Journ._, 1910, Sept. 17, p. 555.
Freeland, W. H., and Renwick, C. W.,
“Smeltery Smoke as a Source of Sulphuric Acid.”
_Eng. and Min. Journ._, 1910, May 28, p. 1116.

LECTURE VIII.

THE BESSEMERISING OF COPPER MATTES.

Development of the Process — The Converter — Converter
Linings — Grade of Matte — Operation of the
Process — Systems of Working.

In modern copper smelting practice, matte of “converter grade,” containing from 30 to 50 per cent. of copper, is bessemerised for the production of metallic copper. Successful practice depends upon a regular and continuous output of matte from the furnace plant being available, and upon a capitalisation and resources on a sufficiently large scale for continuous operation of the whole of the smelting plant.

=Development of the Process for Bessemerising Copper Mattes.=—The success of Bessemer’s process, which was applied in 1856 to the production of steel by blowing air through molten cast-iron, led to a suggestion for its application to copper mattes and to some experiments on the subject by Semenikow, a Russian engineer, ten years later. It was not until 1878 that any further work was conducted on a practical scale. In that year John Holway suggested and worked out the scheme already referred to, the principles of which as outlined by him, form the foundation of the pyritic and converter practice of the present time. Air was blown through heated Rio Tinto pyrites in an ordinary Bessemer steel converter and the experiments met with considerable success. The apparatus was, however, not deemed convenient, as the process worked very intermittently and large quantities of slag were produced which required to be poured off at intervals, whilst the position of the tuyeres in this form of converter was found to be unsatisfactory. There are many practical difficulties in employing the same kind of apparatus for the converting of copper mattes as for the bessemerising of cast-iron into steel. In the first instance, the final steel product differs but little in weight or bulk from the original charge, whilst the process produces but little slag, owing to the comparatively small proportions of silicon and manganese which require to be oxidised—whereas in copper converting, the quantity of slag produced is almost equal in weight to the amount of matte originally charged, whilst the resulting copper product amounts to less than one-half of this weight. Further, in bessemerising cast-iron, the blow is of very short duration; in copper matte converting, it occupies more than two hours, and the relative heat losses are, in consequence, markedly different. Finally, the lining of the steel converter chiefly serves to protect the shell; its function in the copper converter was to act also as flux for the iron oxides produced on blowing.

In Holway’s final form of apparatus for the pyritic smelting of copper ore to metal, the introduction of siliceous material as a flux for the iron oxide and the use of basic lining were arranged for, with the object of overcoming the difficulties caused by the corrosion of the siliceous lining which acted as flux.

Though several years elapsed before the pyritic treatment of ore was successfully conducted, the process of bessemerising the fluid matte to metal was successfully applied on a commercial scale by Manhès in 1880, although it was not until the following year that David’s device of placing the tuyeres horizontally and at such a height above the bottom as not to interfere with the metal which is obtained, solved the final difficulties of operation on a practical scale. In 1883–4 the Manhès converter was introduced into the United States, and at about the same time the barrel form was designed by Manhès and David, and was also readily adopted. Both forms developed in size, increasing in capacity from 1 ton to that of 7 to 10 tons.

Until comparatively recent years, the chief modifications in practice were concerned with operating and constructional details rather than with radical changes in the principles of work. Experiments and research have meanwhile been in constant progress with the object of overcoming several of the grave defects connected with the apparent necessity for the destruction of the siliceous converter-lining by using it as flux, which was due to the difficulties of causing the iron oxide to flux with silica when introduced in any other way.

The most vital improvement introduced into converting practice, and that with which the future developments are most closely bound, is the successful adaptation of basic material for the purpose of lining the converter. This achievement, together with recent success in the introducing of siliceous flux, promises to solve many of the difficulties connected with the bessemerising of low-grade matte by a continuous process.

Suggested by Holway, basic linings were tried at the Parrott Smelter, Butte, in 1890, by Keller and others, but under the conditions of working at that time they were found to be unsuccessful when operated on an industrial scale. Valuable pioneer work was undertaken by Baggaley in Montana, and after many trials, his method was successfully operated for some months at the Pittsmont Smelter under Heywood’s direction in 1906. Visits of inspection to this smelter in 1908 proved disappointing, it being found that most of the plant which had promised the solution of such difficult problems had been dismantled, largely owing to economic difficulties connected with its operation, and the works were in process of re-organisation for the older system of working. Meanwhile, since 1903, Knudsen, at Sulijtelma, Norway, has successfully employed a small basic-lined converting furnace for the combined pyritic smelting and converting of heavy sulphide ores. The process consists usually of pyritic liquation of the sulphides, followed by a further concentration of the matte up to ordinary converter grade by bessemerising, the higher grade matte being then transferred to a silica-lined vessel and blown to metal in the usual way.

The successful operating of the basic-lined converter on the large scale and under the conditions of working at great modern plants was first established by Smith and Pierce at the Baltimore Copper Company’s Smelter, and the method has since been installed and worked with success at Garfield, Utah (five converters in operation, one in reserve); at Perth Amboy, N.J.; at the Washoe Smelter at Anaconda—where the whole plant is being adapted for basic-converting—and at several other works.

A recent and promising development has been the reported successful blowing of fine siliceous concentrates through the tuyeres of converters at the Garfield Smelter, a method by which it might be possible to effect the rapid and efficient extraction of values from fine material otherwise difficult to deal with, affording at the same time a means of conveniently supplying siliceous flux in a manner possessing many advantages.

=Principles of the Bessemerising Process.=—The principles underlying the converter process are those which form the basis of pyritic smelting practice—of which bessemerising is but a phase. The reactions involve the very rapid oxidation of iron and sulphur under practically ideal conditions, and the fluxing by silica of the iron oxide so produced. The heat of oxidation keeps the materials in a thoroughly molten state, and maintains the temperature well above that required for slag formation and perfect fluidity. The heat derived by the combination of oxygen with the iron and sulphur and that of the iron oxide with silica is developed so rapidly and in such quantity, owing to the large masses now worked with, as to cause a reaction-activity sufficient to make the process independent of heat from external sources.

It will be noted how markedly the more recent developments of copper smelting have taken advantage of the factors of the _time element_ and _mass influence_ in obtaining enormous heat intensities and consequent high temperatures, by conducting oxidation of sulphides as rapidly and in as large mass as possible. The same absolute quantities of heat per unit weight of charge were liberated in the older smelting methods involving roasting, but the more leisurely manner of operating allowed the dissipation and dispersion of much of this heat, thus necessitating the employment of supplementary carbonaceous fuel.

=The Converter.=—The converter is a lined steel vessel in which the molten matte is contained, and which allows of air being blown through the material by means of tuyeres which pass through the walls.

The early form of converter was bottom-blown, and similar to that invented by Bessemer, but it was not successful in operation on the small quantities of copper matte worked with, owing to the chilling effect of the cold air on the copper, which, when produced, sank to the bottom and set above the tuyeres, stopping the air blast, and causing much loss of metal in the slag.

The later form of converter was barrel-shaped, with a horizontal row of tuyeres situated at some distance above the bottom so as to allow the copper to settle, protected from the action of the blast, and also to allow of the punching of the tuyeres as required.

The modern forms of converter comprise both the vertical and the barrel types, modified largely as regards size and constructional details, and although the vertical form is still in use and is even preferred at several smelters, it has been largely superseded at most plants by the barrel-shaped variety, whilst the possibilities of greatly enlarged vessels using basic linings are likely to favour this replacement still further.

1. _The Upright Bessemer Vessel_ is used, and found satisfactory at Great Falls and at Mt. Lyell. The general size has been 8 feet diameter and 16 feet height, with a capacity ranging from 5 to 12 tons, according to the condition of the lining, though at Great Falls converters of 12 feet diameter with corresponding capacity are now in use. The advantages of the vertical form are, that, owing to the greater depth of matte through which the air passes, the oxidation is more rapidly conducted, the lining is more efficiently supported, and the wear by abrasion upon the lining is found to be considerably less in amount and to be more uniformly distributed.

On the other hand, the greater depth of matte necessitates a greater blowing pressure in order to force the air through the material, whilst control over the operations becomes a matter of greater difficulty.

2. _The Barrel Form of Converter_ is the type in common use. Among the advantages claimed for this form are those which accrue from being able to operate the same weight of matte in more shallow layers, as compared with the upright form—thus requiring lower blast pressures. Another advantage is the greater ease of regulating the depth of material blown through, by tilting the converter and thus altering the relative position of the tuyeres.

Owing to the successful adoption of the basic lining, the barrel type of converter has now to be divided into two classes, since the basic converter differs from the silica-lined type in constructional details, and is usually of much larger dimensions. Its operation is also conducted on somewhat different lines.

(_a_) _The silica-lined barrel converter_ varies somewhat in size, the Anaconda converters were, however, representative of the most convenient dimensions.

The shell consists of ¾-inch boiler plate, 8 feet in diameter, and 12 feet 6 inches long. The converter is constructed in two portions, the body and the hood, in order to facilitate removal, relining, and general repairs. The ends are lined with 9 inches of firebrick, and the body with 4 inches; it is then rammed with lining material to a thickness of about 18 inches in all parts. There are 16 1-inch tuyeres placed horizontally, and in the latest forms of converter, the air is supplied by individual tuyeres which are connected to the blast box, and which are provided with ball-valves to prevent leakages and back-running during the necessary punching. The cavity is about 8 feet × 4 feet by 6 feet deep when first made, and the converter then holds conveniently about 7 tons of matte. The weight of lining is about 16 tons, and it lasts six to nine blows. The blast-pressure used is 16 lbs. per square inch.