TABLE X.—BLAST-FURNACE CHARGE CALCULATIONS—TOTAL CHARGE, ALL FURNACES._Modern Copper Smelting being lectures delivered at Birmingham University, greatly extended and adapted and with and introduction on the history, uses and properties of copper.

─────────────────────┬───────┬────────────┬────────────┬────────────

│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.

[Illustration: Fig. 47.—=V=-Shaped Charging Car, indicating Mechanism

for Release and Tilting.]

=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.

[Illustration: Fig. 48.—End View of Blast Furnace, showing Tilting of

Charge Car, Anaconda.]

[Illustration: Fig. 49.—Hodge’s Charging Car.]

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.

[Illustration: Fig. 50.—Freeland Charging Machine (D. S. C. & I. Co.).]

[Illustration: 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 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.

[Illustration: Fig. 52.—Slag Spout, showing Method of Trapping Blast,

also Replaceable Nose-piece of Spout (A).]

_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.

[Illustration: Fig. 53.—Details of Slag Spout, Cananea.]

[Illustration: Fig. 54.—Slag Spout, showing Method of Support.]

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.

[Illustration: Fig. 55.—General View of Settler (T. E. Co.).]

[Illustration: Fig. 56.—Method of Lining Settler, Cananea.]

[Illustration: 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 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.

[Illustration: Fig. 58.—Tap-hole Casting and Detail for Settlers.]

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.

[Illustration: Fig. 59.—Anaconda Blast Furnace (51 feet long), showing

Settlers.]

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.

[Illustration: 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

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.

Deficiency of silica also results in the production of _over-fire_,

owing to the fact that the air blast, being unable to bessemerise any

more iron sulphide at the tuyere zone, passes to the higher portions

of the furnace and gradually roasts the ore there, thus consuming the

sulphide fuel of the furnace which might otherwise be most effectively

used for bessemerising in the tuyere zone. This over-fire, resulting

from the heat of roasting which is given out in the upper part of

the furnace, is very disadvantageous in true pyritic smelting, and

successful control of the process depends on using up the whole of

an adequate air supply at the bessemerising zone, and on supplying

sufficient siliceous flux to combine at once with the whole of the iron

oxide produced. For fluxing purposes it is only the free silica in the

charge which is effective, since any silica existing as silicate is

already in a state of combination and thus is not free to act as flux.

The combined silica, except for its adding to the fusibility of the

charge by admixture, is very disadvantageous, consuming heat and space,

diluting the reaction intensities by presenting an inert substance

among the active constituents, and increasing the quantity of slag

which requires to be melted.

The three requirements—iron sulphide, oxygen supply, and fluxing

silica—thus bear an intimate relationship to one another in true

pyritic smelting, and alteration of any one factor requires

simultaneous adjustment of the others for the production of the same

grade of matte and slag. The speed and degree of oxidation primarily

depend on the air supply. The more iron burnt up, the greater is the

heat production and the higher the temperature at the tuyere zone,

and since the more basic slags are known to have the higher formation

temperatures, the basicity of the slags increases with the speed of

oxidation and consequent concentration.

Ores suitable for true pyritic smelting are not commonly met with

in practice, and the presence of earthy bases other than iron is

not desirable. Whilst the advantages of polybasic slags from the

point of view of reduced formation temperature, increased fusibility

and liquidity are very marked in ordinary smelting practice, their

presence is not so advantageous in true pyritic smelting, since they

consume silica which is required for the iron oxide at the instant

of formation, and thus tend to decrease the speed of oxidation and

concentration. Polybasic slags have a lower formation temperature, and

in consequence the production of the highly ferruginous slags of high

formation temperature which it is desired to make by the oxidation

of as much iron as possible is retarded. In addition, the presence

of other earthy bases in the charge dilutes its fuel value; they may

even consume valuable heat by requiring decomposition, as in the case

of carbonates. These considerations are not so important in partial

pyritic smelting, where the required heat balance can be adjusted by

coke.

_The Advantages of Pyritic Smelting._

(1) The possibility of direct and immediate treatment of highly pyritic

raw ore in the blast furnace, thus saving all the costs of preliminary

treatment and handling.

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

In former smelting practice, high sulphide contents in a copper ore

were particularly disadvantageous, since the higher the sulphur

contents of the charge the lower was the grade of the resulting matte,

when smelted directly in the blast furnace. In consequence, the higher

sulphur content necessitated a more complete roasting of the ore in

order to ensure a high-grade matte on smelting.

With pyritic smelting the conditions are completely reversed, and

the charge becomes more suitable for direct furnace treatment as its

sulphide contents increase, so that the most suitable ores for pyritic

smelting are those in which the greatest saving is effected by their

not requiring a preliminary roasting operation.

As has been already indicated, this saving includes labour, plant,

handling, time, and interest on capital tied up in the roast yards, as

well as the avoiding of all the mechanical and other losses connected

with such preliminary treatment. Thus at Ducktown, Tennessee, the

material economies effected by the substitution of pyritic smelting for

the processes involving preliminary roasting amounted to no less than

3 to 4 cents per pound of copper produced, in addition to the later

advantages derived from the recovery of values from the gases, and from

the improved conditions of life in the district.

(3) The cost of coke is saved.

Fuel is one of the main items of expense in blast-furnace smelting, and

by the substitution of the cost-free natural-sulphide fuel for coke,

the proportion of the latter required on the charge is reduced from the

9 to 10 per cent. formerly employed with roasted materials to about

3 to 5 per cent., and in certain special cases to very much smaller

amounts.

_Difficulties of the Process._—That the technical difficulties in

applying the process on a practical scale are considerable, under

present conditions of working, will be understood from the nature of

the operations.

(1) The pyritic process works on a narrow margin of heat, and allows

of but little flexibility in the conditions of working, since

there are few factors which can be altered should difficulties in

operating arise, as compared with the circumstances when a free use

of supplementary carbonaceous fuel may be employed. The only source

of heat energy at the smelting zone is in the sulphide charge itself,

and small variations in the working conditions may readily disturb

the delicate equilibrium upon which successful working depends.

Irregularities, stoppages, and variations in grade of matte may

therefore arise, unless the operations are regulated with exceeding

watchfulness. In true pyritic smelting the employment of coke for

restoring the balance or for producing heat required at the tuyere

zone is not permissible or practicable, since, as will be indicated

later, such coke addition would altogether destroy the equilibrium in

the process; the grade of matte and the composition of slag would be

altered, the reactions disturbed, and a restoration to normal pyritic

smelting conditions rendered almost impossible.

Difficulties in operation have therefore to be overcome along the lines

of pyritic action—that is, in the further adjustment and manipulation

of blast, sulphide or silica supply, or in charging methods, etc.—and

in practice such careful “doctoring” is resorted to when the furnace

shows signs of working unsatisfactorily.

It is very often possible by such careful attention to gradually bring

a furnace back to smooth running. It occasionally happens, however,

that the conditions gradually become worse, and the furnace commences

to show signs of “gobbing.” This is indicated at the top of the charge

by the formation of crusts round the side and end walls, whilst from

the slag spout below, there issues a much reduced quantity of thick

siliceous slag, together with an abundant stream of thin low-grade

matte. The furnace gradually ceases running, and it becomes necessary

to stop its working, to take down the furnace jackets, bar out the

debris, and restart operations. This is usually not so objectionable

a procedure as it might appear, and indeed, within certain definite

limits, such a course may economically be sound policy. In the modern

operation of pyritic practice it often pays better to risk the

occasional gobbing up of a furnace and clear out the debris, than to

work with so large a quantity of coke as would avoid such a necessity.

Not only is the modern furnace so designed and constructed as to entail

but comparatively little trouble in cleaning out in this manner, but

such practice, even if temporarily a necessary evil, may, in places

where coke is expensive, and where conditions for pyritic smelting are

otherwise favourable, be, within certain definite limits, actually the

most profitable. It is by the taking of these risks, combined with

further experiment and working experience in manipulation, such as in

charging methods, blast conditions, and the height and distribution

of charges, etc., that the ultimate continuous and successful working

at still lower costs may be attained and the true pyritic process be

worked as continuously as ordinary smelting practice. Short campaigns

are not, therefore, unusual under the present conditions of true

pyritic smelting, and the cleaning out of the generally fairly loose

debris is accomplished with moderate ease, from 24 to 36 hours being

the usual time required to take down, clean out, and restart a furnace,

whilst the cost of such an operation (chiefly in labour) is not, under

the circumstances, excessive. At Tennessee, hard driving and short

campaigns result in lower costs and greater tonnage.

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

It has been indicated that the thermal conditions in the bessemerising

zone of the pyritic furnace tend to the production of highly basic

slags, which, though hot and limpid, are characterised by high density.

Such slags are not conducive to good settling and separation of mattes,

and they tend to occasion high copper losses, because—

(_a_) The difference in density of slag and matte is

not sufficiently great.

(_b_) The solubility of sulphides in the slag increases

with its basicity.

The greater the concentration effected by the smelting operation,

the higher is the grade of the matte produced; at the same time, the

actual weight of matte is smaller. On the other hand, since more iron

is oxidised from the charge and slagged off, the quantity of slag

produced increases proportionately. Contrasting then, the likely losses

of copper which would result from the association of a small quantity

of high-grade matte with much slag, compared with those resulting from

the association of a considerable quantity of low-grade matte in the

presence of but little slag, the former condition is obviously the more

productive of heavy loss, for not only will many more shots of matte

be held in suspension, but each shot of high-grade matte represents a

larger quantity of copper.

It is found in practice that it is most economical to make a fairly

low-grade matte on the first or “green-ore” smelting, and to

re-concentrate this matte pyritically up to converter grade by a second

smelting operation. The extra cost of casting the low-grade matte,

of breaking up, rehandling it, and resmelting, with all the extra

charges on capital, etc., involved, is less than the losses which would

be incurred if higher-grade converter matte were made at the first

smelting, although there is no difficulty at all in producing such

mattes so far as the actual furnace operations are concerned. It is

entirely a question of the slag losses involved.

Under ordinary smelting conditions (not truly pyritic), when using

some coke for fuel, it would be readily possible to alter the density

of the slag by adding suitable constituents, such as limestone or

additional silica, but in pyritic smelting this is not practicable.

The furnace chooses to make at the tuyere zone its own slag, and

that a highly basic one. High concentration and a slag low in iron

content cannot be obtained together in true pyritic smelting, since

high concentration means rapid oxidation of iron sulphide, and this

necessitates high temperature and produces a highly ferruginous slag

in consequence. Additional silica added to the charge could not alter

the slag composition markedly and still yield the same grade of matte.

The silica content of the slag depends on the temperature at the

tuyere zone, and this is governed by the rate of oxidation of the iron

sulphide. If the slag is to be more siliceous it must be produced at a

lower temperature, which would be obtained by oxidising the iron less

rapidly. This would lead to the production of low-grade matte, and

probably would so reduce the furnace activity that there would not be

sufficient heat to keep the slag molten.

If extra silica be added to the charge, it would probably be unattacked

unless more iron were oxidised in order to flux it off. In such a case

the blast would have to be increased in order to produce iron oxide

more rapidly, the temperature would in consequence be raised, a still

more basic slag would be produced in larger quantity, whilst the matte

would be increased in grade and reduced proportionally in weight.

The addition of sufficient lime to the charge, in order to produce a

sufficiently low-gravity slag, is also impracticable in true pyritic

work, because—

(_a_) The extra lime consumes silica, and interferes with the

desired reactions at the bessemerising zone, tending to lower the

concentration. It also absorbs heat.

Lime has a very powerful affinity for silica, more strongly marked

than that of iron oxide, its replacing value is higher, its more

siliceous silicates are readily formed and they have a lower formation

temperature, all of which factors tend to an undue consumption of

silica which is urgently required by the iron if the rate of oxidation

is to be maintained. The marked tendency for lime in the charge to

consume the silica tends to retard the oxidation of the iron sulphide,

which proceeds most satisfactorily when free silica is available for

the nascent iron-oxide, and in consequence concentration is decreased

and the heating effect in the furnace reduced. In addition, the larger

bulk of calcareous slag carries considerable heat from the smelting

zone of the furnace. Lime silicates and the polybasic lime slags have

a markedly lower formation temperature than the normal ferruginous

slags of true pyritic smelting, they are hence formed readily without

requiring so much oxidation activity at the tuyere zone. In consequence

less iron is oxidised, and the resulting concentration in the matte is

proportionately reduced.

(_b_) The lime is introduced in the form of limestone, and the carbon

dioxide liberated from this material in the furnace is found to have a

deleterious effect on the furnace gases if the manufacture of sulphuric

acid from them is intended—this being a consideration of great economic

importance in connection with many modern pyritic smelters.

Hence, in practice, pyritic smelting is at present generally conducted

in two stages for the production of a matte of 30, 40, or 50 per cent.

converter grade. The “green ore-matte,” or first matte, runs usually

from 8 to 13 or 14 per cent. of copper, depending upon the copper ore

available, which is usually very low grade—2 to 3 per cent. copper

contents; the second or concentrated matte assays 28 to 40 per cent.

copper. Special care is taken to ensure good settling of matte from

the basic and irony slags, and by these means the copper losses in the

slags are reduced to the comparatively moderate proportions associated

with normal practice.

It does not appear improbable that with the developments of basic

converter practice, involving eventually the continuous converting

of low-grade mattes, the necessity for this second pyritic smelting

and re-concentration may be avoided. The removal of this feature

from pyritic smelting practice would add enormously to the potential

economies arising from the method.

In spite of the difficulties connected with the process, as detailed

above, the method has proved itself an exceedingly profitable one

on a large scale, and the experience of the companies financially

interested, as well as the opinions of managers of the plants in

practical operation, leave no doubt as to the economic success of this

application of scientific principles to a practical problem on a very

extended scale.

SPECIAL FEATURES OF PYRITIC SMELTING.—Several points of particular

interest have given rise to much discussion in connection with pyritic

smelting practice. These include the question of the coke proportion

required on the charge, and the advisability or otherwise of employing

heated blast for the furnace.

_Coke Proportion._—Whilst ideal pyritic practice involves the entire

absence of supplementary carbonaceous fuel, it has not been found

practicable, up to the present, to ensure satisfactory working over

any reasonable period of time, unless a minimum of about 1·25 per

cent. of coke is incorporated with the charge. The function of this

coke has been a matter of much speculation, but the investigations

of Sticht already referred to, now permit the tracing, with some

considerable accuracy, of its function and of its action in the furnace.

It is found that in true pyritic smelting the coke does not reach the

bessemerising zone at all, but that it is completely consumed in the

regions above this point. It is, moreover, not burned by the oxygen of

the air, none of which exists above the tuyere zone, since all this

oxygen is consumed by the combustion of the sulphide. It appears that

the coke is oxidised by the SO_{2} which results from this sulphide

combustion. The examination and analysis of samples of the gases

withdrawn from different parts of the furnace have confirmed this view,

and have elucidated the probable reason for the apparent necessity of

a certain small proportion of coke in the process, under the present

conditions of working. The heat generated from the oxidation of the

coke by the SO_{2} is of much value in preheating the materials of

the charge for the removal of excess sulphur and the liquation of the

sulphides. The amount of heat which is available for this operation

is small, being practically all derived from that carried upwards by

the hot gases leaving the smelting zone, and none is obtainable by the

usual processes of coke or sulphide oxidation in the upper regions of

the furnace, since no available oxygen is believed to get past the

bessemerising zone and reach these upper areas. It is indeed necessary

for the success of pyritic smelting that such oxidation or roasting

of sulphides in the upper part of the furnace should be prevented,

since every available particle of iron sulphide is required for heat

production at the smelting zone, by its combustion there, and any

oxidation elsewhere not only deprives this zone of fuel, but spreads

the heat over too wide an area for sufficiently intense combustion.

Thus, by supplying an additional amount of heat to the upper parts of

the furnace, where heat is needed to assist in the preparation and

liquation of the sulphides, the extra coke, in being oxidised by the

SO_{2} without robbing the tuyere zone of fuel or air, just fulfils

its useful purpose at the required place, in such a way as to keep the

smelting operation running smoothly.

The presence of more coke than is absolutely necessary for the

fulfilment of this purpose is, in addition to its extra cost, of

no advantage, and in true pyritic smelting none should reach the

tuyere zone, since it introduces a reducing influence where the most

marked oxidising effect is required. By consuming oxygen for its

combustion, it deprives the iron sulphide of this material, less iron

is, therefore, oxidised, and the matte is consequently increased in

quantity and lowered in grade, whilst the amount of iron carried into

the slag is decreased.

1·25 per cent. of coke is about the minimum quantity with which it is

found practicable to maintain satisfactory working of the furnace under

present conditions, 0·5 per cent. has been worked with occasionally,

and none at all over certain short periods of time. The average

quantity employed is from 2 to 3 per cent., and when about 5 per cent.

is used, coke reaches the tuyere zone and the process ceases to be

truly pyritic—the reactions and smelting conditions become entirely

changed.

It does not seem unlikely that, as knowledge of these conditions

increases and as the mechanism of the process becomes more generally

understood, modifications in furnace design and blast conditions may

lead to the successful operating of the pyritic process entirely

independent of the use of coke fuel.

_Heating of the Blast._—For true pyritic smelting it has been shown

in practice that the use of heated blast possesses no advantages;

many smelters operating the process have tested the effects, and have

usually given the method up, whilst the work of Sticht and Peters

affords valuable evidence and close argument as to the reasons for

its unsuitability. Success in true pyritic working depends upon the

intensity of oxidation of the sulphides, and upon the localisation

of the resulting heat at the narrow bessemerising zone situated just

above the tuyeres. The greater the quantity of iron which is there

oxidised per minute, the better is the concentration, the greater is

the smelting and fluxing intensity and the higher is the resulting

temperature. Since the character and composition of the slag vary

in accordance with these conditions, depending largely upon the

temperature in the tuyere zone, the furnace works most rapidly and

satisfactorily when slags of high formation temperature are being

produced. These can only be formed if much iron is being oxidised,

because iron is the chief fuel in the process. The addition of extra

heat by warming the blast appears to allow of the formation of silicate

slags possessing a lower formation temperature, such slags are less

basic, and consequently less iron need be oxidised and slagged off per

minute in order to produce them. Less iron sulphide fuel is, therefore,

burned, and the reaction intensity at the tuyere zone is reduced, so

that the necessary heat margin for satisfactory smelting may not be

attained. The extra heat carried in by the warmed blast may not be

sufficient to compensate for that which is lost owing to this decrease

in oxidation intensity; the furnace consequently tends to work cold,

whilst the excess air supply leads to the production of over-fire, by

the oxidising of sulphides higher up in the charge.

These features are specially interesting, as they afford one of the

most marked distinctions between true and partial pyritic smelting.

In the latter process, the fuel value in the adjustable supply of

coke at the tuyeres allows of the ready production of any extra heat

which might be required. The slag composition is, in consequence, more

independent of the furnace conditions, since the heat required for the

smelting operation does not depend so much on the formation of slag

of any particular composition. Sufficient heat is always obtainable

by coke additions when smelting for any special slag which may be

desired. Neither is localisation of the heat at the narrow tuyere zone

so essential in partial pyritic smelting. Warm blast produces a greater

combustion intensity when employed in oxidising carbon, so that it may

present advantages, both economic and operative, in partial pyritic

work, whereas it is distinctly disadvantageous in the true pyritic

method.

=Pyritic Smelting Practice in Tennessee.=—The pyritic process is

operated in Tennessee at two smelters; that at Copperhill under the

Tennessee Copper Company, and at Isabella by the Ducktown Sulphur,

Copper and Iron Co. The ore averages from 2 to about 2½ per cent.

copper, 31 to 37 per cent. iron, 20 to 30 per cent. sulphur, 10 to 25

per cent. silica, the remainder being earths, including lime about 6

per cent., magnesia 2 per cent., zinc 2 per cent., and alumina—_i.e._,

a heavy sulphide ore with but little excess of free silica available

for the fluxing of iron.

_Copperhill._—The process is conducted very much according to the

principles just considered. The Copperhill plant operates seven

furnaces of the ordinary rectangular water-jacketed type—the general

features of furnace design being little different at present, whether

true or partial pyritic practice be conducted. Several important

devices in detail have been introduced with successful results, and

the management is distinguished for its pioneer work and experimental

enterprise in connection with the process. The furnaces were formerly

all 56 inches wide; three of them are 180 inches long, the other four

being 270 inches. The height of charge is from 10 to 12 feet, the

capacity of the smaller furnaces 375 to 400 tons of charge daily, and

a blast of 19,000 cubic feet of air per minute at 50 ozs. pressure is

supplied to each. The larger furnaces have a capacity of 500 to 600

tons daily. Many trials have been made to determine the best shape for

the water-jacketed sections, both broad and narrow panels having been

employed. In one of the furnaces, curved end-jackets were tried, with

the object of lessening the production of crusts which tend to form at

the corners, owing to coldness and reduced furnace activity at these

points. The advantages expected have not been realised, the tendency to

crusting has not been lessened, and although barring has been rendered

easier, the disadvantages of rounded corner-jackets and their greatly

increased cost of construction outweigh their advantages, and their use

has now been given up.

[Illustration: 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.

Modern Copper Smelting being lectures delivered at Birmingham University, greatly extended and adapted and with and introduction on the history, uses and properties of copper. · 第 49 / 71 题