  Charge Calculations—Charging—Working—Disposal of Products—Pyritic Smelting—Sulphuric Acid Manufacture from Smelter Gases. Charge Calculations.—Modern practice aims at the production of a matte of converter grade, containing usually from 40 to 50 per cent. of copper, and preferably in a single smelting operation; except in true pyritic work. Full analysis of the whole supply of material available at the smelter is essential, as well as a report on the quantities of each separate constituent. The first step in the charge-calculation is the computation of the total weights of copper, iron, and sulphur available for the smelting campaign; from these quantities the losses of copper and sulphur to be allowed for during the operation itself, as based on previous experience, are deducted. The balance indicates the quantities of these elements from which the matte and slag can be produced. The copper is transformed into matte, in which product it may be regarded as existing in the form of copper sulphide, Cu2S, and the sulphur required for this combination with the copper is calculated from the relation—
Thus every unit of copper combines with one-quarter of its own weight of sulphur. A matte of converter grade containing, say, 44 per cent. of copper is constituted as follows:—Copper, 44 per cent. + sulphur, 11 per cent., or copper sulphide, 55 per cent., the remaining portion of the matte being iron sulphide, which amounts to 100 - 55, or 45 per cent. Assuming as a first approximation that this iron sulphide has the formula FeS,
hence 7/11 of the remaining 45 per cent. of the matte is iron and 4/11 is sulphur—that is, the matte contains in addition, iron 28 parts, sulphur 17 parts. Hence the composition of the converter matte is approximately—Copper 44 parts, iron 28 parts, and sulphur 11+17=28 parts. The amount of copper for the matte is fixed by the available ore supply; the quantity of sulphur is controlled by the furnace operation and charges, as judged from previous experience—the oxidation being so regulated that the proper grade of matte is produced. The iron required for the matte is next considered. Every 44 parts of copper require 28 parts of iron for the production of a matte of the correct grade. If the quantity of iron in the materials available at the stock-bins be not sufficient to furnish the amount required, as just calculated, ferruginous material must be added as flux, if, on the other hand, there is a superabundance of iron available in the charges for this purpose, the excess must be fluxed off. In this manner the amounts of the constituents for the matte production are determined, and the composition and making up of the slag-forming constituents are next considered. In this connection the local conditions with respect to proximity and cost of suitable flux, as well as experience with the previous working of the furnace and ore charges are important factors in determining the type and composition of the slag to be made, whilst in true pyritic practice the special conditions of working fix certain limits to the composition of the slag, as will be indicated later—the pyritic furnace “tending to make its own slag.” In partial pyritic smelting, the coke allowance and the furnace conditions allow of fairly wide latitude in making up the charges for the production of suitable slags with which the furnace can deal efficiently, since the heat production is not dependent on the formation of any particular slag. It is always possible to add extra coke for the purpose of melting the slag desired. The scientific principle governing the calculations for slag composition is the proper proportioning of acid and basic constituents. This is based upon the oxygen ratio—i.e., the proportion of oxygen The requirements for a satisfactory slag are that it shall be—
It is well known that within certain broad limits of silica content, slags will fulfil these conditions to a greater or less extent, whilst the most suitable and economic slag under any particular circumstances is decided, as stated above, by the composition of the charge, the quantity and character of the available fluxes, and the previous experience with the furnace. The limits of the silica content for suitable slags as just indicated are fixed by several well-known general properties of the silicates. Speaking broadly, and from the point of view of the more or less ferruginous silicates constituting copper-smelting slags, the more basic silicates—such as the subsilicate class (oxygen in acid:oxygen in base < 1:1)—are generally characterised by high formation-temperature, and by being very fluid, thin and fiery, dense and corrosive. On the other hand, the more acid silicates, such as those of the multi-silicate class (oxygen in acid: oxygen in base > 2 :1) are characterised by lower formation-temperature and low density, and by being thick and viscous. As the silica content within this range of silicates increases, the melting point is lowered and the specific gravity is reduced, features which are very advantageous from the point of view of the production of clean slags. Their fluidity, however, decreases, and a very high temperature is thus required in order to render them sufficiently limpid to run freely from the furnace. On this account the highest proportions of silica usually considered feasible in a slag, correspond to the bisilicates of the representative composition, MO. SiO2. With high temperature conditions in the furnace and rapid working, such slags can be dealt with successfully, and if the charges are necessarily highly siliceous, it may be advantageous from the economic point of view to work with this class of slag. In proportion as the silica content gradually decreases and as they become more basic, the silicates are more and more corrosive and fiery, and especially in the case of the iron silicates, they gradually attain such a high specific gravity that efficient settling of the matte is not As is well known, mixed silicates—i.e., silicates of two or more bases—are generally characterised by the properties of increased fusibility, and often of increased fluidity, and their employment is usual and generally advantageous in smelting practice. The relative proportion between the various bases in such mixed silicates is largely a matter depending upon the prevailing conditions at the smelter. In modern smelting, particularly where partial pyritic work is conducted, and where fairly siliceous charges are worked, a slag running about 40 per cent. SiO2 is aimed for, iron and earth oxides constituting the remaining 60 per cent. or so. In cases where this quantity of iron is present in the charge, the slag may be constituted chiefly of iron silicate, but even in such instances the advantages of lime additions are marked. When iron is not available in sufficient quantity, the extra fuel costs and working difficulties of running with more siliceous slags would render their production undesirable, and the purchase of limestone or similar earthy flux is particularly advantageous. The purely iron silicates are usually dense, and thus tend to hold up copper values both in mechanical suspension as well as in solution; the addition of lime, which has a marked effect in reducing the specific gravity, permits of more basic slags being worked with, where necessary, without such heavy losses in the slag. The presence of lime silicate with the iron silicates has a marked influence on the fluidity of the slags, even when they are more highly
hence for the production of a slag of the same oxygen ratio, less weight of lime would be required to flux off the same weight of silica; in other words, the replacing values of the two oxides are as 112 to 144, or 7 to 9. Of the other bases which are occasionally present in slags, the proportions of the oxides of magnesium and zinc are sometimes considerable, the calculations being analogous to the previous cases. The case of alumina is anomalous, and its behaviour in slag production is not definitely understood. Many experienced workers hold the view that it tends to act either as acid or base, according to the proportions of silica. Thus, in a very siliceous slag, alumina in moderate quantity behaves as a basic oxide, forming aluminium silicates, and in very basic or low silica slags the alumina appears either to neutralise some of the excess base, acting as an acidic oxide, or to dissolve as such in the slag, whilst in intermediate cases it possibly behaves partly as an acid and partly as base. This view has recently been questioned, and it has been suggested by Shelby that alumina always acts as an acid in the formation of slags. The matter is thus one which requires further considerable investigation. Usually neither alumina nor zinc oxide behave very satisfactorily in the furnace when present in large quantities, tending to thicken the slags and to promote viscosity. Anaconda Practice in Charge Calculations.—An example of some of the practical considerations which enter into the calculation and making up of charges is well illustrated in certain particulars of the practice as conducted at Anaconda. Details of the materials charged over a period of one month are indicated in Table X. The important charge constituents available in large quantity include:—
The briquettes are next worked in to as great an extent at possible, since by this means the large stocks of settling-pond slime and of screened fines are reduced and their 5 per cent. of copper is extracted. The whole stock of old slags and residues is used up on the charge, these materials introducing considerable amounts of copper, whilst being irony, they further help to reduce the acidity of the slag, thus saving the employment of the lime-rock otherwise required for fluxing. The total quantity of copper, iron, and sulphur available being then calculated, and the allowances for sulphur elimination and for the copper loss on smelting (2 to 7 per cent.), as based upon previous experience, being deducted, the amount of iron required to constitute the 45 per cent. copper matte is estimated. From this figure the FeO remaining for slag production is determined. The silica introduced by the above materials is also known, and the amount of lime-rock required to produce an easily running slag is next calculated. The slag which is found by experience to give the most satisfactory running has a composition of about—
Variations from this composition, especially as regards higher silica contents, immediately introduce difficulties, increasing the expense of furnace running, by requiring more fuel and care in working, reducing tonnage, and producing a slag which runs far less freely. So that although the large quantity of siliceous material at hand might tempt the management to work with a more siliceous slag, and so save the procuring of such large amounts of barren lime-rock, the cost of this material is much more than compensated for by the advantages which result from the working with a slag which contains only about 40 per cent. of silica. The quantities of the charge constituents thus calculated, divided by the likely number of charges to be worked, are entered up on the charge sheet, which is handed over to the charge foreman. Fig. 47.—V-Shaped Charging Car, indicating The Charging of the Blast Furnace.—The method of “hand charging,” as employed in the older processes of working, when using small furnaces of small output, possessed several theoretical advantages, but it is essential in modern practice, where at least 300 tons of charge, and often much larger quantities, are fed into the blast furnace daily, to employ mechanical means for charging. At many smelters, however, the coke is added separately, from barrows. Care in the charging is now recognised as being of special importance for successful blast-furnace operation, especially for the purpose of procuring the correct distribution of coarse and fine material. The principle of keeping the sides more open by distributing the coarser materials against the jackets and keeping the fine parts nearer to the The practice still commonly employed is to feed the materials from side-dumping cars (of very varied design) brought along in a train drawn by locomotives and travelling along tracks running at each side of the furnace. A form of car frequently used has a V-section, and it is secured in a vertical position whilst in transit by some form of catch-pin device, which is readily released when it is required to tilt the car for charging. Another form, employed at Anaconda, has a shaped section, the sides of which are pivoted and admit of being very readily secured or unfastened as desired. The car bottom itself is tilted by connecting it with a compressed air lift by means of a hook situated at the side of the car remote from the furnace. The material is thus discharged along the inclined chute so produced. An interesting method is employed at the Granby Smelter, where the Hodge car and the end-feeding method are in use. The cars, which have a double-hopper discharge, are divided into four compartments by vertical plates. These cars enter at the ends of the furnace through suitable openings at the level of the feed-floor, and run by small wheels on tracks which are built inside the furnace along the side of each vertical wall. In this manner a straight vertical fall for the charge is arranged, and this affords the best control of proper distribution. The furnace holds three cars at a time, and there are patent openers and closers for manipulating the end doors of the furnace, as well as for releasing the hopper-bottoms of the cars. A particularly ingenious and successful device is in use at the Ducktown Smelter of the D.S.C.I. Co., Fig. 48.—End View of Blast Furnace, Fig. 49.—Hodge’s Charging Car. At the Cananea smelter is operated an ore-bedding system, the store-bins feeding the charge down hoppers through which it falls directly into the furnace. A similar feeding system is in use at Garfield, Utah. The lay-out of the plant to allow of the most efficient charging is so arranged as to locate the stock-bins at a high level, so that ore is fed directly from the discharge chutes into the cars of the charge trains which run on tracks underneath, and these tracks are situated at Fig. 50.—Freeland Charging Machine (D. S. C. & I. Co.). Fig. 51.—Freeland Charger—Details. The charge foreman receives from the blast-furnace department his charge sheets which inform him of the amounts of the various materials to be loaded on to each car—calculated in the manner already indicated. Proceeding to the stock-bins, the gates and chutes of which are automatically controlled, he sets the scale of the weigh-bridge which is situated under each bin to the desired weight. At the same time an electric-light indicator is switched on in front of the particular bins from which material is to be withdrawn, thus assisting in spotting the cars and checking the weighing-out. The charge train is brought along the tracks running underneath the bins, and into each car The Coke Allowance.—As has been already indicated, the coke allowance depends largely upon the nature of the charges and the individual experience at the smelter. The main principle involved is to reduce the coke consumption as much as possible by applying the pyritic principle to the fullest possible extent, working as much sulphide material into the charge as is economically practicable. In partial pyritic smelting, where the coke may constitute from 5 to 10 or 12 per cent. of the total charge, it is usual not to feed it in with the rest of the materials from the cars, but to charge it into the furnace separately. The charge foreman puts it in just when and how he considers it necessary, and he is encouraged to use as little as possible, consistent with proper running of the products at the slag spout. In pyritic smelting proper, the small amount of coke is fed on to the top of the charge-material in the charge-cars. Working of the Blast Furnace.—The top of the charge, which is usually some 3 to 5 feet below the level of the feed-floor, appears fairly uneven, there being a tendency for it to sink along the middle. It is moderately hot, showing practically a black heat except where red-hot patches near the side appear in positions corresponding to where the tuyeres are situated below. There is not very much fume at the feed-floor level if the chimney draft be good, nor excessive agitation at the top, unless much fine material is being worked. Sulphide fines tend to the formation of accretions near the top of the charge and occasionally lower down, also to a considerable extent against the walls of the brick superstructure—this is said to be lessened considerably by the use of water-jacketing at these parts, which also greatly assists the barring down of the masses. A considerable amount of barring is sometimes necessary when much fine concentrate is worked, otherwise a well-managed furnace runs smoothly and satisfactorily under favourable conditions. Trouble may arise occasionally by leakages occurring in the jackets or spouts, but by the modern methods of sectional construction and by the devices for time-saving in making the necessary connections, working is usually not seriously interfered with for a very long period. Even for the removal or replacement of a slag spout, the slag-hole is plugged, and the repair is completed within an hour and a-half, by which time slag is again running freely over the replaced slag spout. The tuyeres are punched regularly two or three times per shift, and a steady stream of material issues from the slag notch and over the spout to the settlers. Disposal of the Furnace Products.—Under ordinary circumstances, the products resulting from the blast-furnace operations include—
Fig. 52.—Slag Spout, showing Method of Trapping Blast, The Matte and Slag.—In modern practice, as already indicated, the fluid products of the blast furnace are run out of the furnace as rapidly as possible, and flow continuously, as they are formed, through a trapped slag notch. So important has this principle of rapid removal of the fluid products become, that the hearth or crucible portion is being made smaller and smaller. The slag notch, is, in addition, placed so low that only so much molten material remains in the furnace bottom as is necessary for the regulation of the temperature for maintaining perfect fluidity of the materials during their discharge, and for avoiding crust formation on the hearth. The depth of material remaining in the bottom—that is, the distance from the hearth bottom to the slag notch—is from about 8 to 12 inches, depending on the conditions just indicated. The discharge of the furnace products takes place through the trapped slag notch of the furnace, an opening constructed in the tapping-breast The slag spouts are often of sheet steel, sometimes of copper or of bronze, and are from 3 feet 6 inches to 5 feet in length, being separately water-cooled units. The discharge at the end is from 12 to 18 inches higher than the centre of the slag notch in the tap-jacket through which the molten material issues from the furnace. The spout is secured to the tap-jacket, being arranged so as to admit of ready replacement where necessary. Usually it is bolted to the jacket and is securely wedged up against it, being supported at the discharge end by the wall of the settler, and the joints are made perfectly tight by very careful asbestos packing and claying. The spout lasts for several months, the greatest wear being at the end over which the molten stream issues, but the life has been considerably lengthened, with greatly increased convenience of furnace working, by providing the spouts with separate easily replaceable water-cooled nose-pieces of cast-iron which are bolted to the ends, thus taking up most of the wear and tear, and allowing of a very ready removal and replacement without disturbing the slag-spout connections to the furnace itself. These are indicated in Figures 52 (A) and 59. The slag spout is protected along its entire length by a hood of clay, by which means the stream of matte and slag running down it is maintained hot and fluid. Fig. 53.—Details of Slag Spout, Cananea. Fig. 54.—Slag Spout, showing Method of Support. Settlers.—The modern type of settler is often circular in section, about 16 to 18 feet in diameter and 5 feet in height, storing about 40 tons of matte. Other forms, rectangular or oval, are, however, also employed. The outer shell is of ½-inch steel plate bound together by band-bolts, the lining is often 9 to 15 inches in thickness, with an inside layer of looser stuff. The lining material employed varies greatly, according to the grade of matte, character of slag, and working conditions. The wearing out of the lining depends very largely on the class of material passing through the settler, the most rapid wear being occasioned by the fiery and corrosive low-grade mattes and basic slags, whilst high-grade mattes and more siliceous slags give little trouble in this connection. The more corrosive the products, the more refractory and hard-wearing must be the lining, and consequently the materials employed for the purpose range from chromite, silica brick and firebrick down to loam, according to the requirements; the chief duty is that of being non-corrodible and of protecting the outer shell. It is not an uncommon practice to thicken the walls close to the tap-holes, where they are subjected to most wear, and often chromite is used at these points owing to its power of withstanding the forces of erosion. On the other hand, at the Copperhill Smelter of the Tennessee Copper Company the settlers have been found to give as satisfactory service on fairly low-tenor matte, when lined throughout with good firebrick as with the more expensive materials formerly used, whilst still more recently, siliceous copper ores have been successfully employed as lining material instead of bricks. Fig. 55.—General View of Settler (T. E. Co.). Fig. 56.—Method of Lining Settler, Cananea. Fig. 57.—Arrangement for Matte and Slag Discharge from Settlers (T. C. C.) There is usually a spray of water from a circular pipe which surrounds the settler near the top—this playing against the steel sides keeps the outside cool and protects the lining. The settler is roofed over with slag, except at the back where the stream of matte and slag enters, and also at those points where the slag overflows. The slag escapes over short launders attached to the top of the steel casing. The position of these discharges depends largely on the arrangement of tracks, size of furnace, temperature of working, and quality of products. Under modern conditions of high temperature and rapid working, they are situated as far away from the entrance as possible, thus giving fuller opportunities for very quiet settling in a large pool and affording gentle overflow of slag with little abrasive action on the linings. These outlets may be situated opposite to the entrance or at the sides. The discharge spouts for slag may be one or two in number, usually of cast-iron coated with thin clay, and often roughly hooded over with clay. They have replaceable cast-iron nose-pieces to facilitate repair after wearing down. The continual gentle stream of Fig. 58.—Tap-hole Casting and Detail for Settlers. The hole through the brick wall for tapping is about 1½ inches in diameter, and the matte is discharged through a tapping piece of cast-iron, 6 inches in diameter and 3 inches thick, perforated by a 1-inch hole. This iron disc has, cast around it, a copper tapping-plate about 1 foot in width and 2 feet high, which is recessed into the steel sheet of the settler. In the iron tapping-piece is a conical recess, into which the conical clay plug is rammed when closing the tapping-hole. These iron tapping-pieces withstand the action of converter grade matte fairly well, and are conveniently replaced when necessary—about once a month. They are illustrated in Fig. 58. The tapping-plate is fixed into position in a special section of the shell, known as the launder casting, to which the matte launder is secured, whilst a newer form of settler has the tap sections also removable, so that these can be taken out and the brick renewed during the campaign of a furnace, being as readily removable as a furnace jacket. The matte launder is of cast-iron or of steel, thickly coated with clay or suitable material (slime-pond product, etc.) to protect it from corrosion. In modern work the steel tapping bar is always rammed through the conical plug and tapping-hole until it just reaches the matte, so that its withdrawal by ring and wedge is readily performed when the matte is to be tapped whilst by this means the tap-hole is securely closed. The workers are protected from shots of matte, etc., during tapping or closing, by means of a slotted sheet-iron hood which can be swung back when not required, a convenient and useful as well as necessary precautionary device. Matte is tapped from the settlers into ladles as required by the converters; such ladles are constructed of thick steel plate, washed with clay, and often lined with a hull of chilled material. It is sampled at the runner with each tapping. The tap-hole is closed by a clay plug on the end of a dolly which is rammed home, and a warm pointed steel bar is then driven through until it reaches the matte, being knocked in occasionally as the end is very slowly eaten away. Several of the features named in the previous sections are well indicated in the photograph (Fig. 59) of the tapping platform at the Anaconda Smelter. The “Gaseous” Products of the Furnace.—Great variation is to be found in the arrangement at different works for the disposal of the gaseous products of the furnace. Reference will be made later to the methods employed in connection with pyritic work, and where the gases are to be utilised for the production of sulphuric acid. Formerly the general method, even at the large modern plants, was to lead the gases from the top of the superstructure to the off-takes and large dust-catcher flues, thence to the stack. Fig. 59.—Anaconda Blast Furnace (51 feet long), showing Settlers. Fig. 60.—Hoppers of Flue-dust Chambers and Tracks for Cars underneath. The off-take flues of the modern furnace are of steel, 4 to 6 feet in diameter—lined or unlined according to circumstances—and leading to very large dust chambers of varying design, sometimes rectangular, often of large circular section, or of balloon-shaped section, etc. In all cases these flues are provided with hopper discharge openings at suitable intervals, under which cars run on tracks, for the collection and conveying of the dust. Arrangements for the further settling and collection of the flue-dust are essential in connection with modern blast-furnace plants, where blast pressures of from 40 to 50 ozs. per inch are employed and where it is often found economical to work with The gaseous products of the furnace carry solid matter in two forms. As a rule, under the usual conditions of copper-smelting charges, the larger portion of the solid matter thus carried is in the form of very fine particles of charge material itself, mechanically suspended and carried over in the current of the escaping gases. This is the flue-dust. In addition, values in the form of volatilised metallic products are also conveyed by the gases, particularly when lead, zinc, arsenic, etc., are present in the furnace charge, and these are carried forward in the form of fume. They tend to solidify as the temperature of the gases becomes lower, although their settling is very greatly impeded owing to the exceeding minuteness of their particles and also to their dilution; the problem of separating and collecting them is in consequence attended with great difficulty. Chambers of enormous capacity are required in order to give the fine solid particles an opportunity of settling by decreasing the velocity of the gases and by cooling them down, whilst for the settling of fume, capacious flues in which are suspended wires or similar devices for assisting the process must be adopted. Where large quantities of lead, etc., are present some bag-house system of fume filtration is necessary, especially if silver be present, since this metal tends to be carried over in the leady fume. At the majority of copper smelters such extreme refinements are rarely necessary, although modern legislative requirements make severe demands on the managements for the freedom of the gases from injurious constituents. Dry settling methods and filtration are in general use where such separation is required and the use of high-tension electricity has been successfully tried at Californian smelters. Wet methods have so far not proved economically successful. The flue-dust from the flues is dealt with in a number of ways, according to the conditions at the smelter. It may be smelted with the “roaster-calcines charges” in the reverberatory furnaces, although excessive quantities have proved difficult to deal with in certain Still more recently, the East Butte Copper Mining Company has installed and successfully operated a sintering plant on the Dwight-Lloyd principle for the treatment of the flue-dust preparatory to blast-furnace smelting. The capacity of the plant is 100 tons per day. The material is rendered more or less cohesive by the effects of heat alone, but the operation is not yet perfect. (See Mining Journal, Jan. 6th, 1912, p. 21.) The freed gases finally pass along series of long and capacious brick main-flues connecting with all the branch flues, furnished with discharge hoppers at intervals, gradually rising and discharging into a wide stack of such a height that damage to vegetation in the district is entirely prevented. Pyritic Smelting.—Modern blast-furnace practice, as has been stated, is conducted according to two main systems of working:—
The term Pyritic Smelting (or pyrite smelting) is thus applied to that class of practice in which the whole of the heat required in the smelting zone is obtained by the combustion of the ore or matte charge itself; it implies the application of the pyritic principle to the extreme limit, the use of carbonaceous fuel being reduced to a minimum. Ideal working is to feed unroasted ore or matte, together with the requisite fluxes, into the blast furnace, and by the action of an adequate air blast, to burn out part of the sulphur and iron, the former escaping with the furnace gases, the latter being slagged off, whilst the copper in the charge is concentrated in the matte product of the operation. This type of smelting is conducted at a number of large modern works, and though up to the present time the use of coke on the charge has not been entirely eliminated, research and practical experience have History.—The idea originated with John Holway, of London, who sought to extend to the smelting of copper the principles so brilliantly applied by Bessemer to steel manufacture, and who, in a work which was published in 1879, suggested and demonstrated the process of utilising the heat of oxidation of the iron and sulphur constituents of copper-bearing materials for the smelting and extraction of the copper. That work is to-day recognised as one of the most masterly expositions of the principles underlying pyritic smelting and converting, and many of the most important and recent developments in these branches of work are proceeding on lines forecasted by him. Holway’s experiments, conducted on a considerable scale, proved the feasibility of the principles underlying the process, which was to prepare metallic copper from sulphide ores in one combined series of operations in a single furnace unit. Owing, however, to mechanical troubles and difficulties of operation, as well as to the ultimate withdrawal of financial support, he was unable to carry the process to a commercial success, and the single-stage process is at present regarded as being beset by almost insuperable difficulties, although the latest phases in modern practice are tending towards a realisation of Holway’s scheme of working. His paper and published results deserve the closest study. Inspired by the pamphlet, an English Company in 1887–8 attempted the practice at a smelter at Toston, Montana, and showed the possibilities of the method, although the plant available did not lend itself to completely successful operation. L. S. Austin, who took a leading part in this work, patented the process in the United States, and developed the practice, and in 1891 Dr. Peters conducted a very full enquiry into the conditions of working, which placed the system on a definite practical basis. From that time the method has developed coincidently with the more empirical practice at many works of replacing coke fuel by sulphides to as great an extent as possible. T. A. Rickard focussed scientific and practical opinion on the subject in the symposium on “Pyrite Smelting,” which he called forth and edited, and many celebrated smeltermen have contributed to the progress of pyritic smelting practice. At the Copperhill Smelter of the Tennessee Copper Company and at the Ducktown Sulphur, Copper and Iron Co.’s Smelter at Isabella, Tennessee, remarkably good pioneer work was done by Parke Pyritic smelting is at the present time being very successfully practised at Mt. Lyell, Tasmania; at Tennessee; Tilt Cove, Newfoundland; and other districts, whilst the smoke problem alone has prevented for a time a number of other smelters from successfully operating the process. The Mechanism of the Process.—The mechanism of the changes involved in the pyritic process is now fairly well understood in general outline. One of the most important steps in elucidating the matter was made by Sticht’s discovery that the oxidation area of the furnace in pyritic smelting was confined to a narrow zone situated just a little higher than the tuyere level; by actual experiment it was found that scarcely any free oxygen existed above this narrow tuyere zone. The usual and typical ore charged into the furnace in pyritic smelting is impure chalcopyrite (essentially a copper-bearing pyrites, FeS2). When heated in an atmosphere free from oxygen, this pyrites loses some of its sulphur and approaches pyrrhotite in composition. On further heating in a neutral atmosphere more sulphur is evolved and the material approaches FeS in composition, whilst at very high temperatures and under favourable circumstances, a still further quantity of sulphur is liberated, resulting in the production of the well-known fusible iron sulphide, which is the eutectic of the iron: iron-sulphide series of alloys, melting at 970° C., and containing about 85 per cent. of FeS. Thus in the pyritic furnace, free sulphur is liberated as such at the upper levels, and passes up the furnace unchanged until it meets free air above the surface of the charge, when When, however, they reach the blast of air which enters the furnace at the tuyeres, an intense action proceeds as the sulphides become bessemerised. The heat of oxidation of iron sulphide has long been known to be very great, and Holway pointed out that this heat corresponds to the large quantity of heat which is developed by the free roasting of heavy sulphides, compressed into the space of a few moments, and thus results in an exceedingly great intensity with consequent high temperature. Sulphur is burnt out to SO2, iron is converted to the oxide which instantly combines with the white-hot silica skeleton that is present and forms an iron-silicate slag, evolving still more heat. This slag, with the enriched matte, melt thoroughly at the prevailing temperatures, and issue from the slag spout of the furnace. The work of Sticht and Peters thus allow of the mechanism of the processes being followed during the passage of the materials through the furnace. At the Mount Lyell Smelter, where Sticht operated, the charge extends about 12 feet above the tuyeres. In the upper 6 or 7 feet, elemental sulphur is driven off from the pyritic materials by the effects of heat alone, and the furnace gases in this zone consist chiefly of nitrogen, SO2 (from the bessemerising), sulphur vapour, a little CO2, but practically no free oxygen. About half-way down, the temperature is sufficiently high to melt out the fusible sulphides from the charge; these liquate and trickle unchanged through the still solid masses of gangue and silica-flux, until they meet with free oxygen of the air blast, when they are oxidised and burnt up with great rapidity and with the evolution of intense heat. This bessemerising zone extends from a short distance above the tuyeres to a point where all the oxygen is used up by the iron and sulphur. The distance is variable, but is probably some 2 feet or so. At this level the ferrous oxide produced is instantaneously seized by the white-hot particles of free silica with the production of a silicate slag, the composition of which corresponds to the silicate whose formation temperature is equal to that prevailing in this bessemerising zone. Control of the Operations.—It has thus been established that the oxygen of the air blast entering the furnace through the tuyeres is practically all expended in this bessemerising of the liquated sulphides in the narrow bessemerising zone, and that it does not operate at all by any roasting reactions in the upper part of the furnace, as had been formerly supposed. From this knowledge it therefore becomes possible to indicate the essential factors which control the successful operation of true pyritic smelting. The degree of bessemerising depends upon the amount of air supplied for the oxidation of the sulphides, and upon the quantity of siliceous flux present to slag off the iron oxide produced. The actual smelting takes place at the focus where the liquated sulphides are instantaneously bessemerised, and the more rapid this oxidation, the more intense are the reactions and the higher the temperatures which result. For successful pyritic smelting it is, therefore, essential that there shall be present—
(a) The supply of heat required for the smelting of the charge and the thorough fusion of the products depends entirely on the intense combustion of the iron and sulphur constituents, and the greater the proportion of these materials oxidised per minute, the higher is the temperature. As has been already noted, such heat intensity increases at a rate greater than the mere arithmetical increase in the fuel proportion, by reason of well-known thermo-chemical laws regarding mass effects. Indirectly, too, the higher the proportions of sulphides present, the smaller is the quantity of inert or useless matter which requires to be heated and slagged off in the furnace—apart from the question of the necessary flux material. Hence the higher the iron and sulphur contents of the ore, the more successfully may true pyritic smelting be applied to it. True pyritic smelting may be said to cease when carbonaceous fuel requires to be burnt at the tuyere zone in order to supplement the heat derived from the sulphides, and broadly speaking, from about 28 per cent. of iron and about 30 per cent. of sulphur are necessary in the charge for good pyritic work under present (b) Being supplied with enough sulphide fuel, the requisite quantity of air for the rapid and sufficient combustion of this iron and sulphur is essential. The oxygen is used up entirely in the bessemerising of the sulphides at the tuyere zone of the furnace, and in consequence, not only the heat supply, but also the concentration depends upon the amount of oxygen furnished at this point, since the greater the quantity of oxygen which is used up, the greater is the amount of sulphur eliminated and the amount of iron oxidised and slagged off, and in consequence, the higher is the proportion of copper in the resulting matte. In other words, the oxygen supply largely controls the concentration effected in the smelting process, and consequently an adequate quantity is of the utmost importance. The amount of air theoretically required per minute is readily calculated from the estimated capacity of the furnace and from the charge analysis. Liberal allowances are required for losses, leakages, blower efficiency, etc.; and the volume necessary at the furnace amounts to something like 5,000 cubic feet per minute per 100 tons of sulphide. (c) Sufficient siliceous flux is required for the satisfactory slagging of the iron oxides produced. The presence of the requisite silica on the charge is exceedingly important. The iron of the sulphides, upon oxidation by the air blast, is converted into iron oxides, primarily FeO. This oxide is incapable of existing by itself, but possessing when nascent a powerful affinity for silica at high temperatures, it produces ferrous silicates, which are, in the main, fusible slag-like products. This action is particularly evident in the tuyere zone of the pyrite furnace, where the silica is present in a white-hot condition. If sufficient silica be not present to combine with the iron oxide produced, the ferrous oxide which is exceedingly unstable, finding itself without the necessary flux, is converted under the continued oxidising effect of the blast into higher oxides of iron such as ferric oxide or magnetic oxides, materials which are practically infusible, and this results in the production of an infusible sinter which leads to the choking of the furnace. On the other hand, if excess of silica be present in the charge, highly siliceous and unworkable products result, which will not run out of the furnace. Any further excess of silica simply remains unfused and unattacked, and causes the ultimate stoppage of the furnace operations. The silica for fluxing is consequently an important factor in controlling the running of the pyritic furnace, and the provision of the requisite quantity, as nearly as possible, is essential, since otherwise the presence of adequate sulphide and air blast is not in itself sufficient to ensure satisfactory working. The actual quantity of silica required is determined by the factor known as the formation temperature of the silicates. Every silicate has a definite formation temperature—i.e., a definite mixture of iron oxide and silica requires a definite temperature in order that complete combination may occur and a chemical compound silicate be formed. Conversely, at any definite temperature, only those silicates having a corresponding formation temperature to this degree of heat can be produced. In consequence, if the oxidation of the sulphides at the tuyere zone produces any particular temperature, that particular silicate whose formation temperature corresponds to this will tend to be formed, and the required quantity of free silica must be present to yield this definite silicate with the whole of the iron oxidised. Only a limited quantity of silica can thus be taken up for any definite rate of oxidation of iron sulphide, and the presence of either more or less silica does not greatly affect the composition of the slag. Thus the concentration (sulphide oxidation) is primarily dependent on the oxygen supply, which determines how much iron shall be burnt, but the success of the operation depends upon the presence of the correct amount of silica to flux off this iron oxide. This proportion is fixed by the temperature attained at the tuyere zone, which restricts the silicate produced to such a composition that its formation temperature coincides with this degree of heat. Hence the general law has been deduced and has been confirmed in practice, that “a pyritic furnace produces a slag corresponding in composition to the silicates whose formation temperature equals that prevailing at the tuyere zone,” accounting for the well-known observation “that the pyritic furnace tends to make its own slag.” If the smelting operation is to proceed satisfactorily, slag approaching this composition will be produced, and assuming the air supply to be adequate for the purpose, the absence of the requisite silica on the charge affects the quantity rather than the character of the slag. The amount of iron sulphide oxidised depends largely upon the presence of silica to combine with the iron oxide produced; so much will be oxidised as the silica can deal with, and in consequence, if the free silica supply is deficient, a smaller quantity of slag is formed, whilst the matte will be larger in amount but of lower grade. 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 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 (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 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 (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 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 SO2 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 SO2 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 SO2 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 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 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 Fig. 61.—Slotted Tuyeres, 12 inches by 4 inches (T. C. C.). 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.
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 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) 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, CO2. 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 SO2 and oxygen, and of as little CO2 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 SO2 in the gases delivered at the chambers is the minimum proportion for satisfactory working; CO2 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 SO2 3·5 per cent., CO2 3·5 per cent., SO3 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. 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. 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