Chapter C: THE MACDOUGAL TYPE.—The most important form of modern roaster (4)

In modern smelting, particularly where partial pyritic work is conducted, and where fairly siliceous charges are worked, a slag running about 40 per cent. SiO_{2} 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 siliceous, whilst on account of the lower atomic weight of calcium, lime will, weight for weight, flux off a greater quantity of silica than will ferrous oxide. In forming a slag of similar oxygen ratio, thus—

Mono-silicate of lime, 2CaO . SiO_{2},
Lime : silica :: 112 to 60, or 1 part to 0·54 part.

Mono-silicate of iron, 2FeO . SiO_{2},
Iron oxide : silica :: 144 to 60, or 1 part to 0·42 part;

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:—

Cu. SiO_{2}. Fe(O). S.
% % % %
First-class smelting ore, 8·6 54·0 13·6 14·0
Concentrates, 10·9 26·0 32·0 32·0
Briquettes, 5·0 50·0 13·0 13·0
Lime-rock (flux), .. .. .. ..
Old converter slags and residues, .. .. .. ..

TABLE X.—BLAST-FURNACE CHARGE CALCULATIONS—TOTAL CHARGE, ALL FURNACES.

─────────────────────┬───────┬────────────┬────────────┬────────────
│Tons of│ SiO_{2}. │ FeO. │ CaO.
│Charge.├─────┬──────┼─────┬──────┼─────┬──────
│ │ % │ Tons.│ % │ Tons.│ % │ Tons.
─────────────────────┼───────┼─────┼──────┼─────┼──────┼─────┼──────
First-class ore, │ 28,646│52·80│15,125│14·90│ 4,268│ 0·50│ 143
Second-class ore, │ 1,913│53·50│ 1,023│15·79│ 300│ 0·60│ 11
Lining ore, │ 52│83·71│ 44│ 4·16│ 2│ 0·67│ 1
B. and B. slag, │ 6,667│35·98│ 2,399│47·27│ 3,152│ 1·11│ 74
B. and M. slag, │ 481│42·92│ 206│42·14│ 203│ 0·12│ 1
Precipitates, │ 333│ 8·00│ 27│12·40│ 42│ .. │ ..
Precipitates from │ │ │ │ │ │ │
old works, │ 41│ 2·70│ 1│15·40│ 6│ .. │ ..
Slimes from │ │ │ │ │ │ │
old works, │ 19│56·60│ 11│65·0 │ 1│ 0·80│ ..
Coarse concentrates, │ 14,083│25·27│ 3,558│32·96│ 4,642│ 0·45│ 63
Calcine bearings, │ 232│ 9·50│ 22│57·00│ 132│ 0·80│ 2
Briquettes, │ 27,560│48·77│13,441│15·16│ 4,177│ 0·65│ 179
Reverberatory matte, │ 146│ 4·30│ 6│37·50│ 55│ 0·80│ 1
Reverberatory slag, │ 687│43·10│ 296│39·60│ 272│ 4·00│ 27
Converter cold matte,│ 552│13·60│ 75│29·50│ 163│ 4·90│ 27
Converter slag, │ 9,999│31·30│ 3,129│55·90│ 5,589│ 0·70│ 70
Converter cleanings, │ 7,891│30·53│ 2,437│36·55│ 2,917│ 0·79│ 64
Lime-rock, │ 61,794│ 6·90│ 4,264│ 0·50│ 309│48·80│30,155
Coke, 18,766·235 │ │ │ │ │ │ │
tons, at 14.21 │ │ │ │ │ │ │
per cent. ash, │ 2,667│45·28│ 1,208│12·21│ 326│ 6·31│ 168
─────────────────────┼───────┼─────┼──────┼─────┼──────┼─────┼──────
Total charge, │163,853│28·85│47,272│16·21│26,556│18·91│30,986
─────────────────────┼───────┼─────┼──────┼─────┼──────┼─────┼──────
Total production,│ 18,447│ 6·38│ 1,191│29·36│ 5,486│ 1·57│ 293
─────────────────────┼───────┼─────┼──────┼─────┼──────┼─────┼──────
To slag, │ .. │ .. │46,081│ .. │21,071│ .. │30,693
─────────────────────┴───────┴─────┴──────┴─────┴──────┴─────┴──────

─────────────────────┬───────┬────────────┬─────────────────
│Tons of│ Sulphur. │ Copper.
│Charge.├─────┬──────┼──────┬──────────
│ │ % │ Tons.│ % │ Pounds.
─────────────────────┼───────┼─────┼──────┼──────┼──────────
First-class ore, │ 28,646│15·50│ 4,440│ 6·641│ 3,804,555
Second-class ore, │ 1,913│14·60│ 279│ 5·476│ 209,447
Lining ore, │ 52│ 1·45│ 1│ 3·834│ 3,988
B. and B. slag, │ 6,667│ .. │ .. │ 0·797│ 106,325
B. and M. slag, │ 481│ .. │ .. │ 1·919│ 18,450
Precipitates, │ 333│ .. │ .. │58·853│ 392,352
Precipitates from │ │ │ │ │
old works, │ 41│ .. │ .. │68·344│ 56,607
Slimes from │ │ │ │ │
old works, │ 19│ 7·50│ 1│ 4·203│ 1,637
Coarse concentrates, │ 14,083│32·10│ 4,521│10·782│ 3,036,802
Calcine bearings, │ 232│ 4·50│ 10│ 9·321│ 43,230
Briquettes, │ 27,560│15·32│ 4,223│ 4·928│ 2,716,299
Reverberatory matte, │ 146│23·30│ 34│35·752│ 104,651
Reverberatory slag, │ 687│ 1·10│ 8│ 1·566│ 21,501
Converter cold matte,│ 552│18·20│ 100│42·675│ 470,839
Converter slag, │ 9,999│ 1·10│ 110│ 3·018│ 603,429
Converter cleanings, │ 7,891│ 6·60│ 528│16·840│ 2,688,024
Lime-rock, │ 61,794│ .. │ .. │ .. │ ..
Coke, 18,766·235 │ │ │ │ │
tons, at 14.21 │ │ │ │ │
per cent. ash, │ 2,667│ .. │ .. │ .. │ ..
─────────────────────┼───────┼─────┼──────┼──────┼──────────
Total charge, │163,853│ 8·70│14,255│ 4·357│14,278,136
─────────────────────┼───────┼─────┼──────┼──────┼──────────
Total production,│ 18,447│21·13│ 3,898│39·483│14,567,376
─────────────────────┼───────┼─────┼──────┼──────┼──────────
To slag, │ .. │ .. │ .. │ .. │ ..
─────────────────────┴───────┴─────┴──────┴──────┴──────────

=SiO_{2}= in slag, 46,081 ÷ 110,810 tons slag
=FeO= in slag, 21,071 ÷ 110,810 "
=CaO= in slag, 30,693 ÷ 110,810 "
──────
Total, 97,845 tons, at 88·30 per cent.
= 110,810 tons slag.
_Coke consumption_, 10·63 per cent. wet weight
= 10·96 per cent. dry weight.

Analysis.
┌──────────────────┴───────────────────┐
│ │
=SiO_{2}= in slag, = Calc. 41·59 % Actual 41·30 %
=FeO= in slag, = 19·01 " 19·00 "
=CaO= in slag, = 27·70 " 28·00 "
───── ─────
88·30 " 88·30 "

The other constituents used in the charge comprise varying quantities of materials which accumulate round the works, and which, being rich in copper values, it becomes useful and essential to clean up. For the calculating of the furnace charges, the amounts of cupriferous material available at the stock-bins are reported to the blast-furnace department. The quantities decided upon are divided among the number of charges which are considered likely to be worked off during the day, this number averaging about 1,100. The result of this calculation indicates the amount of each kind of material to be weighed for the separate charges; the analysis of each constituent being naturally known. The materials available for smelting are highly siliceous in character, the first-class smelting ore, of which large quantities are treated, giving a strongly acid composition to the charge; copper-bearing basic materials suitable for fluxing are not available in large quantity, and this necessitates the purchase of barren lime-rock, this item being the largest of the blast-furnace charge. In making up the charge sheet, as large a quantity of concentrate as possible is included, since this constituent is not only high in copper values, but owing to a high iron and sulphur proportion, it increases the fuel value of the charge, the influence on the coke consumption being very marked. The concentrate further forms a base for the matte, and introduces iron, of which there is a shortage, into the slag, thus reducing its too-siliceous character and lessening the quantity of lime which it would otherwise be necessary to procure for the purpose.

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—

SiO_{2}, 41 per cent.
FeO, 19 "
CaO, 29 "

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.

=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 centre is often favoured, since this device reduces the tendency to crusting by the finer sulphide particles against the walls. It is partly with this object in view that the mantel and apron plates are arranged in the hopper form, whilst at the same time the distance between the top of the charge and the feed-floor level is maintained at such a height that this desired distribution of the fresh charges is obtained.

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 [shape: flat base with sides sloping inwards] 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.,[14] Tennessee, where the pyritic process is operated. Careful charging is here held to be one of the great essentials for successful working of the process, especially in the narrow furnaces in use, where the dangers of crusting are greatly increased. The principle of working is, that by dropping the charges vertically downwards, having previously arranged the materials in the desired order across the furnace, they will fall into the position, and be distributed just as desired. The Freeland charger is a kind of conveyor belt made of overlapping steel plates, which is exactly the length and width of the furnace, so that when the machine is brought over it, the furnace opening is entirely covered. The conveyor is carried on a frame mounted on wheels, and this is moved forward and backward by a motor in the front, near which is seated the chargeman who is also the motorman. An independent switch and gearing causes the belt to move round and thus deposit its charge over the end. In front of the frame is a strong catch, fitting into a recess on the cover of the furnace, which is water-cooled and mounted on wheels, so that as the conveyor is brought into position the cover is moved back. All these run along a track which extends below the stock feed bins in the same straight line. The furnace gases are drawn off below the feed-floor.

The method of working is to bring the charger under the bins and to drop the various materials for the charge—weighing 2 tons—on to the belt. By deflectors on the ore chutes, the charge can be directed to any desired position across the belt, and material is thus deposited near the outer or inner side as desired—in falling into the furnace it is found to take the same position that it had on the plates. The charger moves forward and reaches the furnace top, the catch is fastened, and as the charger now advances the cover is pushed back, the conveyor thus taking its place until in its turn it covers the top of the furnace. The motion is now reversed, the conveyor gradually recedes, bringing the cover along with it; meantime the chargeman has set the belt-conveyor gearing working independently, and the belt thus travelling round and over the end pulleys, discharges its burden into the furnace. The disposition of the charge along the length of the furnace can be altered at will by increasing or reducing the speed of the frame. When the conveyor has at last traversed the furnace, the cover is in its place—the charger is now disconnected, and goes back for a fresh load. The furnaces are charged eight times per hour with 2 tons of material. The operations are fascinating to observe, and the control over the disposal of the charge is quite complete, whilst the conditions for the operator are not exceptionally arduous. Many other suitable devices are in use at different works.

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 such a level that the trains are readily and conveniently hauled to the charging platforms of the blast furnaces.

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 is dumped the correct amount of charge, usually to within 50 lbs., with rapidity and ease. The train then passes to the furnace building, where the charges are dumped or otherwise emptied into the furnace.

_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—

(_a_) The liquid matte and slag mixture which is given
opportunities to settle and separate into valuable
matte and waste clean slag.
(_b_) The “gaseous” products carrying considerable
quantities of fume and dust which are settled and
separated in dust catchers and flues, where the
solid matter is collected.

_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 or tap-jacket, which is usually a small special jacket-portion constructed and kept in position separately on account of the great local wear at this point (see Fig. 39). The trapping device is an important and essential feature in connection with the modern practice of rapid and continuous running, the principle being to arrange a sufficient height of molten material at the outer side of the slag opening to overcome the inside blast pressure, and thus prevent the escape of blast with its attendant inconveniences and danger. The flow of liquid material can thus proceed quietly and uninterruptedly. The blast is trapped by the construction of a dam in the form of a slag spout around the slag opening, of such a shape and secured to the tap-jacket in such a manner and position, that the molten material before overflowing at the end, fills the spout and thus covers the discharge outlet of the furnace, trapping the blast so that as fast as the molten products form, a constant stream overflows into the settlers (see Fig. 52).

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.

The position of the outlets from the furnace, connecting to the settlers, is largely affected by the available floor space and the general lay-out and arrangements of the plant. Under suitable conditions, and especially with long furnaces, the arrangement of the settler in front of the furnace works very advantageously, leaving the alignment of the blast furnaces free, and allowing plenty room for working around the settlers. The settlers are then arranged in the middle line of the crucible portion of the furnace, so that working is conducted evenly from both ends of the furnace towards the discharge in the centre, and the smelting is thus regular and allows of good control. At many smelters the discharge of products takes place from spouts at the ends of the furnaces, the settlers thus being in alignment with them. This plan, under suitable conditions, has several advantages, permitting of ready access to the sides of the furnace, even working of the furnace by discharge at both ends, and ready co-operation between adjoining furnaces and settlers.

_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.

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 slag runs along launders, where it is either discharged into slag bogies and dumped, or much better, is met by a strong stream of water which immediately granulates it, and washes it along flumes to the dumps.

The matte tap-holes are generally two in number, situated close to the bottom of the settler, and usually at an angle of 120° from each other and from the entrance spout.

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.

With the introduction of automatic and mechanical charging methods, now being inaugurated to a considerable extent in place of dumping from cars alongside the furnace, the method of withdrawing the gaseous products just below the level of the feed-floor is being adopted.

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 as much fine material as possible, either as such or in an agglomerated form; where too, the dropping of charges from some height and the agitation caused by the blast are practically unavoidable. Rarely less than 2 per cent. of flue-dust is made in any modern blast furnace, whilst 5 per cent. is by no means uncommon, and even larger quantities are often produced. Such dust is, moreover, often somewhat higher in copper contents than the original charge, owing to the brittleness of copper sulphide minerals, which, being more readily broken up, are carried over in the form of fine particles. Hence the economic aspect of the recovery of values, in addition to legislative requirements, call for efficient collection of these products.

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 instances, it may be included in the charges for sintering or briquetting processes, and it has been very successfully incorporated with the matte in beds when it has been necessary to cast low-grade matte into cakes previous to re-concentration in the blast furnace, at a smelter employing the pyritic process.

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:—

(_a_) That in which the heat required in the smelting
zone is provided by the oxidation of the sulphide
materials of the charge—_Pyritic Smelting_.
(_b_) That in which coke or other carbonaceous fuel
is necessary for supplying some of the heat required
in the smelting zone of the furnace, even when the
pyritic effect of the charge is utilised to the fullest
extent—_Partial Pyritic Smelting_.

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 demonstrated that the small quantity which is employed is not utilised as fuel by combustion in the air blast at the tuyeres, but that it is, in fact, oxidised in another manner at some considerable height in the furnace.

_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 Channing, Freeland, and others in developing the process. Enormous service has been rendered within recent years by the masterly researches and brilliant exposition of Robert Sticht, in which latter work Peters has worthily seconded him.

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.[15] It thus became evident that the first series of changes near the top of the charge were those mainly caused by the effects of heat alone, and that only by a second series of changes lower down at the tuyere zone were the reactions of rapid and intense combustion and oxidation of the sulphides being effected. Finally, at the bottom of the furnace, the molten matte and slag collected and ran out. Thus the furnace operations proceed in two main stages; preparation (liquation of the sulphides from the charge) in the upper portion, and oxidation and fluxing (bessemerising of the liquated sulphides) in the oxidising tuyere zone or _focus_.

The usual and typical ore charged into the furnace in pyritic smelting is impure chalcopyrite (essentially a copper-bearing pyrites, FeS_{2}). 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 it there burns to SO_{2}. The residual fusible sulphide melts, trickles down, and becomes the true pyritic fuel of the furnace. The copper sulphide constituents of the charge are practically unaffected in composition by heat alone, and they pass down the furnace with the rest of the charge unchanged until the hotter zones of the furnace are reached, when these sulphides also liquate out, become dissolved in the melting iron sulphides, and are thus carried down to the oxidising zone. Until the sulphides meet free oxygen, no further reactions proceed, since they are without action on silica at even the highest furnace temperatures.

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 SO_{2}, 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, SO_{2} (from the bessemerising), sulphur vapour, a little CO_{2}, 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_) Sufficient sulphides in the charge to give out
the heat necessary for the smelting and for the
thorough fusion of the products.
(_b_) Sufficient oxygen (air) for the rapid and
necessary oxidation of this sulphur and iron.
(_c_) Sufficient free siliceous flux for the satisfactory
slagging of the iron oxides produced.

(_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 conditions. At Tennessee, with about these proportions, the coke consumption on the charge is reduced to about 3 to 4 per cent.; at Mt. Lyell, where the ore runs from 40 per cent. of iron with a corresponding quantity of sulphur, the coke consumption amounts to only about 1·25 per cent. None of this coke probably reaches the bessemerising zone at all.

(_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. An addition of silica to the furnace charge under such circumstances would thus raise the grade of the matte by encouraging the slagging of more iron, and would produce slag of approximately the same composition as before, though in larger quantity.