The Construction of the Blast Furnace.

Dimensions.—The modern blast furnace is a long, narrow, water-cooled shell, rectangular in plan. The dimensions, particularly the length, vary greatly, being regulated according to the anticipated output of the furnace-unit. The size is generally expressed in terms of the internal dimensions at the tuyere level, which represents the smelting area. The width of the modern furnace varies usually from 44 to 56 inches, according to the blast pressure, method and speed of working, concentration to be effected, and so forth. The length in many cases is between 15 and 25 feet, when the furnace may be conveniently worked in connection with one large settler. The capacity of such a unit naturally depends on the conditions of working; it may be taken roughly as from 4 to 6 tons of material per square foot of hearth area per twenty-four hours.

Foundations.—The furnaces are built upon a foundation which is necessarily very strong, being usually either of solid rock or of concrete.

Bottom Plate.—The bottom plate of the furnace usually carries part of the weight of the lower tier of water-jackets as well as the furnace burden, and is supported, some distance above the ground, on screw-jacks leaving an air-space below the furnace, which allows of convenient access for repairs or adjustment. The height of the construction is thus raised to a convenient distance for adjustment to the discharge to the settlers. The bottom plate should consist of sectionised water-cooled cast-iron plates bolted together, with a thin layer of brickwork placed above, to protect them from the corrosive influences to which they are subject. There is a slight slope towards the slag-notch. The actual working bed of the furnace is however, a chilled crust of material which sets on this bottom owing to radiation below, and which, when suitable precautions have been taken, usually adjusts itself naturally whilst the furnace is in operation, by what may be termed automatic radiation. Thus, apart from the water-cooling devices, if the working bottom wears down towards the metal plates, the loss of heat by radiation through the thin layer of material causes a chilling effect which leads to a thickening of the crust. Should the crust thicken unduly and so threaten to interfere with the discharge, the radiation is decreased owing to the thickness; and the high temperature which prevails upon this layer causes a partial melting so that it gradually becomes thinner again—thus regulating itself for the most part automatically.

Fig. 39.—Tapping Breast of Blast Furnace, Cananea ([see p. 139]).

Fig. 40.—Riveted Steel Water-Jacket, showing Tuyere Holes
and Water Inlets, etc. (P. & M. M. Co.).

Fig. 41.—Transverse Section through Modern Blast Furnace,
showing Arrangements of Boshed Lower Jackets, Upper Jackets
and Plates, Stays and Supports, etc.

Fig. 42.—Interior of Anaconda Blast Furnace,
showing Jacketing, Tuyere Holes, and Bridge.

Water-Jackets.—The usual height of the modern furnace, as reckoned from tap-hole to charge floor, is roughly from 14 or 16 feet up to 20 feet, water-jacketed all the way. The sides and ends of the furnace are constructed of sectionised water-jackets arranged horizontally in tiers and vertically in panels. There are usually two, occasionally three, tiers, suitably stayed and supported. The practice as regards the shape and arrangement of the jackets varies greatly. It was formerly not uncommon to work with three tiers of jackets for the sides; of these the lower tier extended only from the sole-plate to the level of the slag-notch, forming practically the crucible jackets, the height varying from 2 feet 6 inches to 4 feet. These were most used when the discharge to the settler was situated at the side wall of the furnace. Above these jackets was situated the second tier through which the tuyeres passed; these build up the boshes of the furnace, and are termed the “bosh” or “tuyere” jackets. In most modern furnaces these two tiers of lower jackets are replaced by one set of panels of from 7 to 10 feet in height, the jackets being given a slight slope towards each other at the bottom, so as to form a very small bosh angle; the contraction is about 8 inches. This improvement does away with a good deal of the jointing otherwise necessary near the hottest parts of the furnace, and thus lessens the danger of leakage at these points. The water-cooled breast-plate containing the opening for the escape of the products is now put in position as a separate piece, well secured to the rest of the jacketing (Fig. 39). Above the lower tier of jackets is placed the upper series, often from 7 to 9 feet in height, which carries the walls of the furnace up to within a few feet of the charging platform. These jackets are parallel, and no bosh is given ([see Fig. 41]).

The end jackets are usually built in two tiers only, the upper, 7 feet to 7 feet 6 inches, as a rule, and the lower, 8 feet to 9 feet 6 inches, according to circumstances; in the smaller furnaces the end wall may sometimes consist of a single jacket only. They are vertical, no end bosh being allowed. The end jackets are each single panels, whilst the side walls are built up in panel sections, the width of which vary, but are often 7 feet to 7 feet 6 inches wide, the panels being bolted or clamped together and strongly stayed.

The water-jackets are constructed of flanged steel plate, the inner sides of which are 5⁄16 to ⅜ inch thick, the outside ¼ to 5⁄16 inch. The seams are flanged outwards, so as to prevent joints, etc., being exposed to the inside of the furnace. The water space between the two plates of the jacket is from 3 to 4 inches.

It is usual to support the weight of these jackets on I-beams carried by the upright columns; very strong bracing and tieing is also necessary in order to prevent the side walls from bulging by the great pressure to which they are subjected. In order to protect the jackets themselves from buckling by the forces acting upon them, they are strengthened inside the water space by a series of ┻ bands, which run vertically downwards between the plates, and are rivetted to the outer side—this device is found not to interfere unduly with the proper circulation of the water. Leakage between the joints of the separate jackets is prevented by asbestos packing. In spite of the strong binding and bracing of the walls in this manner, the connections are so devised as to allow of their being unfastened very easily, so that jackets may be readily disconnected and taken down when it becomes necessary to do so.

Arrangements for the water supply to the jackets vary considerably. In localities where a plentiful supply is available, each jacket has its independent outlet and inlet pipes; in other cases it is common to arrange an independent feed to each set of panels, water being supplied first to the jackets of the lower tier, and being discharged from them to the jackets situated above. The supply pipes for the various jackets branch from water main pipes running at the sides of the furnace.

The tuyere or bosh jackets are pierced horizontally at intervals of about 1 foot, with a line of 5-to 7-inch holes for the fitting in of the tuyere pieces. These are formed of steel thimbles, of ⅜-inch metal, which have a slight taper, fitting secured against the inner plate and rivetted to the outer one, thus allowing of ready replacement when necessary ([see also Fig. 40]). Above the side jackets of the furnace there is usually a heavy mantel-plate, 2 feet to 2 feet 6 inches high, with a sloping front, and surmounting this are apron plates, 1 foot 6 inches to 2 feet high, inclined at 45°, constituting a hopper which directs the charge towards the centre of the furnace in such a way as to keep the fines nearer to the middle line, and thus leave the sides of the charge more open, in order to ensure more regular working.

Superstructure.—The jacketing, together with the apron and mantel plates carry the structure up to the charging floor. Above this is the superstructure with the arrangements for taking off the furnace gases, and for the feeding of material for the charge. In many cases the general practice still prevails of constructing the walls of this portion of brickwork, often about 14 feet high, surmounting this with a hood of metal from the top or sides of which large off-takes carry the furnace gases to the dust chambers, and thence to the flue system and stack. Modifications in the design of the blast-furnace superstructure have been, however, in course of progress at many works, particularly in connection with the employment of automatic or mechanical charging appliances and the taking-off of the gases below the feed-floor level. This is specially the case at plants operating the pyritic process and where the gases are to be utilised for acid manufacture, as well as in connection with the treatment of smelter fume. Several furnaces are also at work using either metallic water-cooled or air-cooled tops, from which the removal of accretions is stated to be very readily effected.

Some of the most recent developments in the design of blast-furnace superstructure have been described by Emmons in reviewing the experiments at the Copperhill Smelter, Tennessee. The gases here are used for acid-making, and are sent to Glover towers under some pressure. The furnace top consists of cast-iron corner-posts and dividers, the walls and ends laid up with brickwork, surmounted by a tubular top of the Shelby type from which the gas off-takes lead. The horizontally pivoted doors open inwards and fit tightly. These arrangements are stated to be very satisfactory.

Fig. 43.—Showing Upper Jackets, Apron and Mantel Plates,
and Superstructure of Blast Furnace, Anaconda.

Fig. 44.—Charging Blast Furnaces at Anaconda.

The charging platform, suitably supported on vertical columns, runs at the upper level, being provided, on either side of the furnace, with tracks of rails for the charge cars. The charging doors usually correspond in position to the panels of water-jackets, and are situated along the whole length of each side furnace-wall, the bottom of the charging opening being flush with the floor. They are generally moved up and down in the grooved guides of the upright columns between them, and are of sheet steel suitably strengthened, from 6 to 7 feet wide and 4 feet 6 inches to about 5 feet high, supported by wire-rope and chains, and operated by compressed air cylinders.

The Air Supply to the Blast Furnace.—The quantity of air required by the blast furnace varies very widely with the class of work, rapidity of output, character of charge, and general smelting conditions. It may be stated roughly as being from 300 to 500 cubic feet of air per minute per square foot of hearth area, at a pressure of about 40 to 50 ozs. per square inch.

The rotary blower of the Roots or Connersville type is very well suited for the supply of these enormous quantities of air at moderate pressures, but for blast at higher pressures the air leakage becomes excessive, and piston-driven blowing engines become almost a necessity. Such improvements have, however, been made in rotary-blowing appliances within recent years that most blast-furnace plants are equipped with blowers of the rotary type, which are found highly satisfactory. The air is brought along blast mains of considerable size—about 30 inches diameter—to the furnace building, thence to the bustle pipes of 24 inches diameter, which surround the furnace, from which branch off the pipe connections (5 or 6 inches diameter) for the tuyeres. The practice of equipping each furnace with its own blowing unit is fairly general, making the necessary reserve connections in case of temporary breakdown; many smelters, however, adopt the system of delivering the air from all the engines into one large common air main, making the necessary connections from this to each separate furnace. The importance of avoiding leakages is recognised, and the requisite valves for regulating and controlling the air supply are arranged for.

From the bustle pipe the air passes down the pipe connections which are attached by flanged joints, thence to the tuyere pipes, which are of cast-iron, the blast being regulated by valves. The actual form of tuyere employed varies considerably, each smelter usually having its own special devices for the convenience of repair, renewal, and fixing, as well as for valve regulation and punching. The tuyere is held against the face of the jackets by bolts, leakages being prevented by asbestos packing.

Fig. 45.—Blast-Furnace Shell, with Air Connections (P. & M. M. Co.).

Fig. 46.—Details of Tuyere, Cananea Blast Furnace.

The tuyeres are usually 4½ to 5 inches in diameter, and are generally placed about 12 inches apart. Air is supplied only through the side jackets, and not at the ends of the furnace.

Heating the Air Blast.—The advisability of heating the air-supply for copper blast-furnace smelting has been the subject of very considerable discussion, the question requiring consideration both with respect to its influence on the rationale of the smelting operation as well as from the economic standpoint. The matter is dealt with more fully in connection with pyritic practice, from which point of view Peters has reviewed the subject exhaustively. It may be here stated that there appears to be no advantage in preheating the air when the true pyritic process is operated, and actual trial has resulted in the rejection of the method at the smelters practising this work.

Where, however, coke fuel to any considerable extent is employed on the charge, a supply of heated air through the tuyeres may result in an increased rapidity of smelting, as well as in the production of hotter and more fluid slags. Especially in partial pyritic smelting and more particularly when working charges which contain but little sulphide and where the employment of much coke is not advantageous, the use of preheated blast may be economically very useful. In such cases, the heat production in the furnace is not so fundamentally bound up with the thermo-chemical reactions of slag formation as it is in true pyritic smelting, and therefore the enhanced intensity of combustion of coke-fuel at the tuyere-zone by the use of hot air may exert an important influence in improving the furnace operation and in decreasing the amount of coke-fuel required. In many such instances indeed it has been chiefly the economic factor with reference to the cost of installing and operating suitable devices for warming the air-supply which has determined the question of adopting this system. As is well known, the use of a supply of heated air causes a largely increased calorific intensity from the combustion of coke, resulting in higher temperature at the tuyere-zone, under which circumstances the charge materials are smelted more rapidly, and the resulting products are more fluid, whilst slags of higher silica content (sometimes economically advisable) can be conveniently worked with.

The devices employed for the preheating of the blast vary considerably—cheapness, capacity, simplicity in design and operation being the main essentials.

The utilisation of the waste heat from the smelting furnaces or products would suggest itself as an economical method for accomplishing the warming of the blast, but in practice several difficulties are encountered in efficiently making use of this heat. Heat is available from two sources, either from the furnace gases or from the hot slag. The very successful operation in cast-iron smelting, of hot-blast stoves worked by the “waste gases,” cannot, however, be applied to copper blast-furnace smelting, since the gases in this case do not possess similar calorific value owing to the small proportions of carbon monoxide present. Further, the temperature of these gases is not sufficiently high to allow of the effective application of the regenerative principle using brickwork chambers. In consequence, the use of metal pipe-stoves offers the only method of utilising the heating values of the furnace gases, but their comparatively low temperature does not afford sufficient heat for the warming of the large quantities of air which are required at the tuyeres.

The much higher temperature of the reverberatory furnace gases offers, however, much greater scope for their utilisation in this respect, if both classes of furnace happen to be in operation at the plant and if they are conveniently situated for the purpose.

At several smelters, blast furnaces have been equipped with hot-blast “tops” for the purpose of preheating the air supply, the air-heating pipes being exposed to the gases in the upper portions of the furnace. The Giroux blast-heating device has been installed on furnaces at smelters in Mexico and Arizona, whilst at others in the same localities, the Mitchell system of baffle passages has been successfully used. The Kiddie system of running the blast pipes through the dust chambers has been tried at Tyee, B.C. The advantages of thus utilising the heat of waste gases have generally, however, been found to be more than balanced by the extra costs involved.

Efforts have been made to use the heat contained in molten slag for warming the air, but owing to the low conducting power of these materials, and the difficulty of bringing extended surfaces in close contact, the method has not proved itself very efficient. Blast is occasionally warmed by passing the air through tunnels in which bogies of molten slag are allowed to remain for some time.

When methods of utilising waste heat from the furnace products fail, the fuel-heated iron pipe-stove is generally employed. Since the temperatures required are comparatively low, and the margin of profit involved by the use of hot blast is usually small, the use of the cheapest class of fuel available is imperative; but many classes of fuel unsuitable for other purposes may find useful application for this work.

The stove is of the usual

cast-iron pipe form, designed to give the maximum exposing surface, suitably strengthened and protected from direct action of the fire. Much valuable information on the advantages, disadvantages, and appliances for blast heating was afforded by the smeltermen who contributed to the symposium on “Pyrite Smelting,” which Rickard edited for the Engineering and Mining Journal.

References.

Mathewson, E. P., “The Development of the Modern Blast Furnace.” Eng. and Min. Journ., May 27, 1911, p. 1057.

Wright, Lewis T., “Metal Losses in Copper Slags.” Bulletin Amer. Inst. Min. Eng., 1909, Sept., No. 33, p. 817.

Shelby, Geo. F., “Cananea Blast Furnaces.” Engineering and Mining Journal, April 25th, 1908.

Emmons, N. H., “Copper Blast-Furnace Tops.” Bulletin Amer. Inst. Min. Eng., Feb., 1911, p. 119.

“Heating Blast.” Engineering and Mining Journal, June 16 and Sept. 15 and 29, 1906.

“Pyrite Smelting,” T. A. Rickard.

Also the Authors already referred to, Austin (p. 80), Gowland (p. 17), Peters (p. 80).

LECTURE VII.
Modern Blast-Furnace Practice (Continued).

Charge Calculations—Charging—Working—Disposal of Products—Pyritic Smelting—Sulphuric Acid Manufacture from Smelter Gases.

Charge Calculations.—Modern practice aims at the production of a matte of converter grade, containing usually from 40 to 50 per cent. of copper, and preferably in a single smelting operation; except in true pyritic work.[12]

Full analysis of the whole supply of material available at the smelter is essential, as well as a report on the quantities of each separate constituent.

The first step in the charge-calculation is the computation of the total weights of copper, iron, and sulphur available for the smelting campaign; from these quantities the losses of copper and sulphur to be allowed for during the operation itself, as based on previous experience, are deducted. The balance indicates the quantities of these elements from which the matte and slag can be produced. The copper is transformed into matte, in which product it may be regarded as existing in the form of copper sulphide, Cu2S, and the sulphur required for this combination with the copper is calculated from the relation—

Cu2S = Cu2 : S :: 2 × 63·5 : 32
::  127 : 32
::   4 : 1 approximately.

Thus every unit of copper combines with one-quarter of its own weight of sulphur.

A matte of converter grade containing, say, 44 per cent. of copper is constituted as follows:—Copper, 44 per cent. + sulphur, 11 per cent., or copper sulphide, 55 per cent., the remaining portion of the matte being iron sulphide, which amounts to 100 − 55, or 45 per cent.

Assuming as a first approximation that this iron sulphide has the formula FeS,[13] the proportions of iron to sulphur in this material are

Fe : S :: 56 : 32
:: 7 : 4

hence 7⁄11 of the remaining 45 per cent. of the matte is iron and 4⁄11 is sulphur—that is, the matte contains in addition, iron 28 parts, sulphur 17 parts. Hence the composition of the converter matte is approximately—Copper 44 parts, iron 28 parts, and sulphur 11 + 17 = 28 parts.

The amount of copper for the matte is fixed by the available ore supply; the quantity of sulphur is controlled by the furnace operation and charges, as judged from previous experience—the oxidation being so regulated that the proper grade of matte is produced. The iron required for the matte is next considered. Every 44 parts of copper require 28 parts of iron for the production of a matte of the correct grade. If the quantity of iron in the materials available at the stock-bins be not sufficient to furnish the amount required, as just calculated, ferruginous material must be added as flux, if, on the other hand, there is a superabundance of iron available in the charges for this purpose, the excess must be fluxed off.

In this manner the amounts of the constituents for the matte production are determined, and the composition and making up of the slag-forming constituents are next considered. In this connection the local conditions with respect to proximity and cost of suitable flux, as well as experience with the previous working of the furnace and ore charges are important factors in determining the type and composition of the slag to be made, whilst in true pyritic practice the special conditions of working fix certain limits to the composition of the slag, as will be indicated later—the pyritic furnace “tending to make its own slag.”

In partial pyritic smelting, the coke allowance and the furnace conditions allow of fairly wide latitude in making up the charges for the production of suitable slags with which the furnace can deal efficiently, since the heat production is not dependent on the formation of any particular slag. It is always possible to add extra coke for the purpose of melting the slag desired.

The scientific principle governing the calculations for slag composition is the proper proportioning of acid and basic constituents. This is based upon the oxygen ratio—i.e., the proportion of oxygen in the acid constituents compared with that in the bases. With the doubtful exception of alumina in certain cases, silica constitutes the entire acid portion of most copper-smelting slags.

The requirements for a satisfactory slag are that it shall be—

It is well known that within certain broad limits of silica content, slags will fulfil these conditions to a greater or less extent, whilst the most suitable and economic slag under any particular circumstances is decided, as stated above, by the composition of the charge, the quantity and character of the available fluxes, and the previous experience with the furnace. The limits of the silica content for suitable slags as just indicated are fixed by several well-known general properties of the silicates.

Speaking broadly, and from the point of view of the more or less ferruginous silicates constituting copper-smelting slags, the more basic silicates—such as the subsilicate class (oxygen in acid : oxygen in base < 1 : 1)—are generally characterised by high formation-temperature, and by being very fluid, thin and fiery, dense and corrosive. On the other hand, the more acid silicates, such as those of the multi-silicate class (oxygen in acid: oxygen in base > 2 : 1) are characterised by lower formation-temperature and low density, and by being thick and viscous.

As the silica content within this range of silicates increases, the melting point is lowered and the specific gravity is reduced, features which are very advantageous from the point of view of the production of clean slags. Their fluidity, however, decreases, and a very high temperature is thus required in order to render them sufficiently limpid to run freely from the furnace. On this account the highest proportions of silica usually considered feasible in a slag, correspond to the bisilicates of the representative composition, MO. SiO2. With high temperature conditions in the furnace and rapid working, such slags can be dealt with successfully, and if the charges are necessarily highly siliceous, it may be advantageous from the economic point of view to work with this class of slag.

In proportion as the silica content gradually decreases and as they become more basic, the silicates are more and more corrosive and fiery, and especially in the case of the iron silicates, they gradually attain such a high specific gravity that efficient settling of the matte is not possible. In addition, the more basic the silicate the greater is its dissolving power for sulphides, hence high copper losses in the slags result from these combined causes. Such basic silicates possess, however, the advantage of marked liquidity, and of flowing from the furnace in a thin limpid stream. The high density and the solvent power of basic slags thus fix a limit to the composition which is considered economically suitable, and the lowest proportions of silica usually worked with correspond to the mono-silicates represented by the formula 2MO. SiO2. Slags containing a greater proportion of base (usually iron) possess too high a density to permit of clean settling. In practice, therefore, the majority of slags are mixed silicates of a composition ranging between the limpid but somewhat dense mono-silicate and the lighter but more viscous bisilicate, corresponding to silica contents of from 30 to 48 per cent. of silica, and within the limits of 35 to 45 per cent. of silica most copper blast-furnace slags will be found. The composition roughly corresponds in a large number of cases to that of the sesqui-silicates of the general formula 4MO. 3SiO2 (oxygen in base : oxygen in acid :: 4 : 6 :: 1 : 1½).

As is well known, mixed silicates—i.e., silicates of two or more bases—are generally characterised by the properties of increased fusibility, and often of increased fluidity, and their employment is usual and generally advantageous in smelting practice. The relative proportion between the various bases in such mixed silicates is largely a matter depending upon the prevailing conditions at the smelter.

In modern smelting, particularly where partial pyritic work is conducted, and where fairly siliceous charges are worked, a slag running about 40 per cent. SiO2 is aimed for, iron and earth oxides constituting the remaining 60 per cent. or so. In cases where this quantity of iron is present in the charge, the slag may be constituted chiefly of iron silicate, but even in such instances the advantages of lime additions are marked. When iron is not available in sufficient quantity, the extra fuel costs and working difficulties of running with more siliceous slags would render their production undesirable, and the purchase of limestone or similar earthy flux is particularly advantageous. The purely iron silicates are usually dense, and thus tend to hold up copper values both in mechanical suspension as well as in solution; the addition of lime, which has a marked effect in reducing the specific gravity, permits of more basic slags being worked with, where necessary, without such heavy losses in the slag.

The presence of lime silicate with the iron silicates has a marked influence on the fluidity of the slags, even when they are more highly 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—

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. SiO2. Fe(O). S.
%%%%
First-class smelting ore,8·654·013·614·0
Concentrates, 10·9 26·032·032·0
Briquettes,5·050·013·013·0
Lime-rock (flux),........
Old converter slags and residues, ........

TABLE X.—Blast-Furnace Charge Calculations—
Total Charge, all Furnaces.

Tons of
Charge. 		SiO2.FeO.CaO.
%Tons.%Tons.%Tons.
First-class ore,28,646 52·80 15,125 14·90 4,2680·50143
Second-class ore,1,91353·501,02315·793000·6011
Lining ore,5283·7144 4·1620·671
B. and B. slag,6,66735·982,39947·273,1521·1174
B. and M. slag,48142·9220642·142030·121
Precipitates,333 8·002712·4042....
Precipitates from old works,41 2·70115·406....
Slimes from old works,1956·601165·0 10·80..
Coarse concentrates,14,08325·273,55832·964,6420·4563
Calcine bearings,232 9·502257·001320·802
Briquettes,27,56048·7713,44115·164,1770·65179
Reverberatory matte,146 4·30637·50550·801
Reverberatory slag,68743·1029639·602724·0027
Converter cold matte,55213·607529·501634·9027
Converter slag,9,99931·303,12955·905,5890·7070
Converter cleanings,7,89130·532,43736·552,9170·7964
Lime-rock,61,794 6·904,264 0·50309 48·80 30,155
Coke, 18,766·235 tons, at 14·21 per cent. ash, 2,66745·281,20812·213266·31168
 Total charge,163,85328·85 47,27216·2126,55618·9130,986
 Total production,18,447 6·381,19129·365,4861·57293
 To slag, ....46,081.. 21,071.. 30,693
Tons of
Charge. 		Sulphur.Copper.
%Tons.%Pounds.
First-class ore,28,64615·504,4406·6413,804,555
Second-class ore,1,91314·602795·476209,447
Lining ore,521·4513·8343,988
B. and B. slag,6,667....0·797106,325
B. and M. slag,481....1·91918,450
Precipitates,333....58·853392,352
Precipitates from old works,41....68·34456,607
Slimes from old works,197·5014·2031,637
Coarse concentrates,14,08332·104,52110·7823,036,802
Calcine bearings,2324·50109·32143,230
Briquettes,27,56015·324,2234·9282,716,299
Reverberatory matte,14623·303435·752104,651
Reverberatory slag,6871·1081·56621,501
Converter cold matte,55218·2010042·675470,839
Converter slag,9,9991·101103·018603,429
Converter cleanings,7,8916·6052816·8402,688,024
Lime-rock,61,794........
Coke, 18,766·235 tons, at 14·21 per cent. ash, 2,667........
 Total charge,163,8538·70 14,2554·35714,278,136
 Total production,18,447 21·13 3,898 39·483 14,567,376
 To slag,..........
Analysis.
SiO2 in slag,46,081 ÷ 110,810 tons slag =Calc. 41·59 % Actual 41·30 %
FeO in slag,21,071 ÷ 110,810 "=19·01 " 19·00 "
CaO in slag,30,693 ÷ 110,810 "=27·70 " 28·00 "
 Total,97,845 tons, at 88·30 per cent. = 110,810 tons slag.88·30 " 88·30 "
 Coke consumption, 10·63 per cent.wet weight = 10·96 per cent. dry weight.

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—

SiO2 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.

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

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.

Fig. 48.—End View of Blast Furnace,
showing Tilting of Charge Car,
Anaconda.

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.

Fig. 50.—Freeland Charging Machine (D. S. C. & I. Co.).

Fig. 51.—Freeland Charger—Details.

The charge foreman receives from the blast-furnace department his charge sheets which inform him of the amounts of the various materials to be loaded on to each car—calculated in the manner already indicated. Proceeding to the stock-bins, the gates and chutes of which are automatically controlled, he sets the scale of the weigh-bridge which is situated under each bin to the desired weight. At the same time an electric-light indicator is switched on in front of the particular bins from which material is to be withdrawn, thus assisting in spotting the cars and checking the weighing-out. The charge train is brought along the tracks running underneath the bins, and into each car 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—

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.

Fig. 53.—Details of Slag Spout, Cananea.

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.

Fig. 55.—General View of Settler (T. E. Co.).

Fig. 56.—Method of Lining Settler, Cananea.

Fig. 57.—Arrangement for Matte and Slag Discharge from Settlers (T. C. C.)

There is usually a spray of water from a circular pipe which surrounds the settler near the top—this playing against the steel sides keeps the outside cool and protects the lining. The settler is roofed over with slag, except at the back where the stream of matte and slag enters, and also at those points where the slag overflows. The slag escapes over short launders attached to the top of the steel casing. The position of these discharges depends largely on the arrangement of tracks, size of furnace, temperature of working, and quality of products. Under modern conditions of high temperature and rapid working, they are situated as far away from the entrance as possible, thus giving fuller opportunities for very quiet settling in a large pool and affording gentle overflow of slag with little abrasive action on the linings. These outlets may be situated opposite to the entrance or at the sides. The discharge spouts for slag may be one or two in number, usually of cast-iron coated with thin clay, and often roughly hooded over with clay. They have replaceable cast-iron nose-pieces to facilitate repair after wearing down. The continual gentle stream of 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.

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.

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.

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:—

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, FeS2). When heated in an atmosphere free from oxygen, this pyrites loses some of its sulphur and approaches pyrrhotite in composition. On further heating in a neutral atmosphere more sulphur is evolved and the material approaches FeS in composition, whilst at very high temperatures and under favourable circumstances, a still further quantity of sulphur is liberated, resulting in the production of the well-known fusible iron sulphide, which is the eutectic of the iron : iron-sulphide series of alloys, melting at 970° C., and containing about 85 per cent. of FeS. Thus in the pyritic furnace, free sulphur is liberated as such at the upper levels, and passes up the furnace unchanged until it meets free air above the surface of the charge, when it there burns to SO2. 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 SO2, iron is converted to the oxide which instantly combines with the white-hot silica skeleton that is present and forms an iron-silicate slag, evolving still more heat. This slag, with the enriched matte, melt thoroughly at the prevailing temperatures, and issue from the slag spout of the furnace.

The work of Sticht and Peters thus allow of the mechanism of the processes being followed during the passage of the materials through the furnace.

At the Mount Lyell Smelter, where Sticht operated, the charge extends about 12 feet above the tuyeres. In the upper 6 or 7 feet, elemental sulphur is driven off from the pyritic materials by the effects of heat alone, and the furnace gases in this zone consist chiefly of nitrogen, SO2 (from the bessemerising), sulphur vapour, a little CO2, but practically no free oxygen. About half-way down, the temperature is sufficiently high to melt out the fusible sulphides from the charge; these liquate and trickle unchanged through the still solid masses of gangue and silica-flux, until they meet with free oxygen of the air blast, when they are oxidised and burnt up with great rapidity and with the evolution of intense heat. This bessemerising zone extends from a short distance above the tuyeres to a point where all the oxygen is used up by the iron and sulphur. The distance is variable, but is probably some 2 feet or so. At this level the ferrous oxide produced is instantaneously seized by the white-hot particles of free silica with the production of a silicate slag, the composition of which corresponds to the silicate whose formation temperature is equal to that prevailing in this bessemerising zone.

Control of the Operations.—It has thus been established that the oxygen of the air blast entering the furnace through the tuyeres is practically all expended in this bessemerising of the liquated sulphides in the narrow bessemerising zone, and that it does not operate at all by any roasting reactions in the upper part of the furnace, as had been formerly supposed.

From this knowledge it therefore becomes possible to indicate the essential factors which control the successful operation of true pyritic smelting. The degree of bessemerising depends upon the amount of air supplied for the oxidation of the sulphides, and upon the quantity of siliceous flux present to slag off the iron oxide produced.

The actual smelting takes place at the focus where the liquated sulphides are instantaneously bessemerised, and the more rapid this oxidation, the more intense are the reactions and the higher the temperatures which result.

For successful pyritic smelting it is, therefore, essential that there shall be present—

(a) The supply of heat required for the smelting of the charge and the thorough fusion of the products depends entirely on the intense combustion of the iron and sulphur constituents, and the greater the proportion of these materials oxidised per minute, the higher is the temperature. As has been already noted, such heat intensity increases at a rate greater than the mere arithmetical increase in the fuel proportion, by reason of well-known thermo-chemical laws regarding mass effects. Indirectly, too, the higher the proportions of sulphides present, the smaller is the quantity of inert or useless matter which requires to be heated and slagged off in the furnace—apart from the question of the necessary flux material. Hence the higher the iron and sulphur contents of the ore, the more successfully may true pyritic smelting be applied to it. True pyritic smelting may be said to cease when carbonaceous fuel requires to be burnt at the tuyere zone in order to supplement the heat derived from the sulphides, and broadly speaking, from about 28 per cent. of iron and about 30 per cent. of sulphur are necessary in the charge for good pyritic work under present 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 SO2 which results from this sulphide combustion. The examination and analysis of samples of the gases withdrawn from different parts of the furnace have confirmed this view, and have elucidated the probable reason for the apparent necessity of a certain small proportion of coke in the process, under the present conditions of working. The heat generated from the oxidation of the coke by the SO2 is of much value in preheating the materials of the charge for the removal of excess sulphur and the liquation of the sulphides. The amount of heat which is available for this operation is small, being practically all derived from that carried upwards by the hot gases leaving the smelting zone, and none is obtainable by the usual processes of coke or sulphide oxidation in the upper regions of the furnace, since no available oxygen is believed to get past the bessemerising zone and reach these upper areas. It is indeed necessary for the success of pyritic smelting that such oxidation or roasting of sulphides in the upper part of the furnace should be prevented, since every available particle of iron sulphide is required for heat production at the smelting zone, by its combustion there, and any oxidation elsewhere not only deprives this zone of fuel, but spreads the heat over too wide an area for sufficiently intense combustion.

Thus, by supplying an additional amount of heat to the upper parts of the furnace, where heat is needed to assist in the preparation and liquation of the sulphides, the extra coke, in being oxidised by the SO2 without robbing the tuyere zone of fuel or air, just fulfils its useful purpose at the required place, in such a way as to keep the smelting operation running smoothly.

The presence of more coke than is absolutely necessary for the fulfilment of this purpose is, in addition to its extra cost, of no advantage, and in true pyritic smelting none should reach the tuyere zone, since it introduces a reducing influence where the most marked 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.

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.

TABLE XI.—Typical Charging Tables at Pyritic Smelter.

B.F. No. 3—Night Shift.
Typical Green Ore Charges.
I.II.III.IV.
Lbs.Lbs.Lbs.Lbs.
Coke, 180 240 240 400..
Ore A.,5,000........
Ore B.,..5,000......
Ore C.,....5,000....
Slag,......4,000..
Lime rock,..........
Green ore (low grade) matte, ..........
Flue-dust,..........
Quartz (for flux), 950........
 Total weight of charge,6,1305,2405,2404,400..
Hours of Charging—No. of
Charges 		No. of
Charges 		No. of
Charges 		No. of
Charges 		Total
per Hr
6–7,22....4
7–8,2..226
8–9,22....4
9–10,2, 2..2, 2..8
10–11,..2..24
11–12,2..2..4
12–1,2, 22....6
1–2,2..226
2–3,22....4
3–4,2..2..4
4–5,2....24
5–6,2, 222..8
 Total No. of charges daily,281214862
B.F. No. 5—Day Shift.
Typical Concentrating Charges.
I.II.III.
Lbs.Lbs.Lbs.
Coke, 150 400 extra 700 ..
Ore A.,........
Ore B.,........
Ore C.,........
Slag,..4,000....
Lime rock, 700......
Green ore (low grade) matte, 3,500......
Flue-dust,........
Quartz (for flux),1,050......
 Total weight of charge,5,4004,400700..
Hours of Charging—No. of
Charges 		No. of
Charges 		Total
per Hr
6–7, 2, 2, 2, 2, 2, 2, 2 ....14
7–8,2, 2, 2, 22..10
8–9,2, 2, 2, 2, 2....10
9–10,2, 2, 2, 2, 2....10
10–11,2, 22, 2, 2, 2..12
11–12,2, 2, 2, 2, 2....10
12–1,2, 2, 2, 2, 2....10
1–2,2, 2, 2, 2, 2....10
2–3,2, 2, 22, 2..10
3–4,2, 2, 2, 2, 2, 2....12
4–5,2, 2, 2, 2, 2....10
5–6,2, 2, 22, 2, 2, 2..14
 Total No. of charges daily,12012..132

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

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

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

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

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

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

(a) For the successful operation of acid-making plant, as at present developed, it is necessary that the proportions of sulphur dioxide in the gases shall not fall below a certain minimum, and further, that the gases shall not contain more than certain limiting proportions of other interfering constituents, such as, for instance, CO2. It is for this reason that the blast furnace operating the true pyritic process furnishes gases of the type most suitable for acid manufacture, since by this process the sulphur-dioxide is obtained in the gases in the most concentrated and the least contaminated form possible under smelting conditions. Even under these circumstances the gases are not in the least of an ideal composition for treatment, owing to their dilution with nitrogen, etc., and the development of the acid-making plants and processes adopted for the successful utilisation of copper blast-furnace gases furnishes a record covering many years of very slow and costly experiment, marked by many preliminary failures and disappointments. These difficulties have now been overcome, as the working of the successful plants attached to both of the Tennessee copper smelters affords conclusive proof, and the sulphur which formerly cost money to dissipate by roasting, now not only acts as fuel, but furnishes a very profitable bye-product.

The requirements for the gases are chiefly the presence of sufficient SO2 and oxygen, and of as little CO2 as possible—factors which depend largely on the proportions of sulphide in the charge. The gas for the acid plant must be supplied in regular and continuous amount, at a specific temperature, and this calls for special care in the smelting operation, furnace manipulation and blast supply, supplementary air admission, etc.

About 3·5 to 4 per cent. of SO2 in the gases delivered at the chambers is the minimum proportion for satisfactory working; CO2 should not exceed about 5 per cent., and about 6·0 per cent. or more of oxygen is also necessary.

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

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

The Ducktown Company’s plant was installed in record time, and, like the Copperhill plant, comprises elaborate dust chambers and flues, with Glover and Gay-Lussac towers of special design and construction, and enormous acid-making chambers with complex valves and fittings. The plant is designed to produce about 160 tons of 60° B. acid daily. The analysis of the gases supplied to the towers varied during the early working of the plant; under fairly normal conditions the average analysis of the gases delivered is SO2 3·5 per cent., CO2 3·5 per cent., SO3 trace; the oxygen in the mixture being about 8·0 per cent. The temperature is also apt to vary. Full details on these points are not yet available for general service.

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

References.

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

Blast-furnace Manipulation.

Shelby, Geo. F., “Alumina in Blast-Furnace Slags.” Eng. and Min. Journ., 1908.

Offerhaus, C., “Copper Blast-Furnace Smelting at Anaconda.” Eng. and Min. Journ., 1908, Aug. 7, pp. 243–250.

Sackett, B. L., “The Granby Smelter Equipment.” Mines and Minerals, 1910, April, p. 524.

“Operations of the Tennessee Copper Company.” Official Annual Reports of the General Manage.r

Walker, A. L., “The Metallurgy of Copper in 1910.” Eng. and Min. Journ., 1911, Jan. 7, p. 39.

Austin, L. S., “Review of Metallurgy in 1910.” Met. and Chem. Ind., 1911, Jan. 11, p. 40.

Rice, Claude T., “Handling Copper Smelting Gases.” Eng. and Min. Journ., 1911, Mar. 25, p. 614.

“Cottrell’s Fume Smelter.” Min. and Scient. Press, Aug. 26, Sept. 2, 1911.

Herrick, R. L., “Boston and Montana Co.’s Smelter at Great Falls.” Mines and Minerals, 1909, Dec., p. 257.

Harvard, F. T., “Condensation of Fume and Neutralisation of Furnace Gases.” Bull. Amer. Inst. Min. Eng., No. 44, 1910, Aug.

“Mineral Industry.” Annual.

Pyritic Smelting.

Holway, John, “A new Application of Bessemer’s Method of Rapid Oxidation, by which Sulphides are utilised for Fuel.” Journ. Society of Arts, Feb. 1879.

Rickard, T. A., “Pyrite Smelting.”

Sticht, Robert, “Ueber das Wesens des Pyrites Verfahrens.” Metallurgie, Nov. 22, Dec. 8, 1906.

Wintle and Alabaster, “Pyritic Smelting.” Trans. Inst. Min. and Met., 1906, vol. xv., p. 269.

Nicholls, F. S., “Pyrite Smelting in Tilt Cove, Newfoundland.” Eng. and Min. Journ., 1908, Sept. 5, p. 462.

Wright, L. T., 44 “Pyritic Smelting without Coke.” Min. and Scient. Press, 1906, Sept. 29.

Sulphuric Acid Manufacture.

Falding, F. J., and Channing, J. P., “Pyrite Smelting and Sulphuric Acid Manufacture.” Eng. and Min. Journ., 1910, Sept. 17, p. 555.

Freeland, W. H., and Renwick, C. W., “Smeltery Smoke as a Source of Sulphuric Acid.” Eng. and Min. Journ., 1910, May 28, p. 1116.

LECTURE VIII.
The Bessemerising of Copper Mattes.

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

Fig. 62.-Sectional Elevation and Plan of
Barrel-Shaped Silica-Lined Converter (Peters).

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

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

Fig. 63.—Latest Form of Silica-Lined Barrel Converter.

(a) The silica-lined barrel converter varies somewhat in size, the Anaconda converters were, however, representative of the most convenient dimensions.

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

The hood is bolted on to the body, and is furnished with conical safety-pieces to give notice of the wearing through of the lining. The converters tilt upon rails, which are strapped round the body, and which travel upon rollers. Motion is communicated to the converter either by connection with an electrical drive, or very often by hydraulic power connecting through a rack to a pinion attached to one of the trunnions. The air supply is usually from piston-driven blowing engines, communicating through a blast pipe to the hollow supporting trunnion of the converter, from which the air passes to the blast box.

Fig. 64.—Longitudinal Section of Basic-Lined Converter.

(b) The Basic-lined Converter.—The adoption of basic linings is of such recent date that although the present form appears to have given satisfaction, later developments in basic practice may cause further modifications in design. R. H. Vail gives the following details:—

As at present operated, the basic-lined converters are long barrel-shaped vessels consisting of a ¾-inch steel shell, 23 feet long and 10 feet in diameter, lined with magnesite materials so as to leave a cavity about 20 feet × 7 feet × 6 feet. Air is supplied from thirty-two 1¼-inch tuyeres, each separately connected with the blast box and controlled by a valve. Provision has to be made for the marked expansion of the basic lining-material by leaving the top of the steel shell open, joining-up the free ends by tie-rods (13, Fig. 65), whilst the tuyere-pipe connections are flexible. The main opening or throat, for the charging of matte and flux, is situated in the arch at one end of the converter; it is 40 inches in diameter, and surmounted by a short chimney-cap of iron, which is 30 inches high and lined inside with clay. The vessel is charged through this opening. Metal and slag are poured from the converter through an opening in the side opposite the tuyeres, which is kept closed by bricks during the operations. An oil-burner is provided at one end, for the purpose of supplying such extra heat as might be required, in consequence of undue cooling of the copper towards the end of the blow or for heating up the lining after repairs. The converter is supported as in acid practice, though a tilting device employing wire ropes attached to hydraulic plungers is now being introduced in place of the rack and pinion method.

Fig. 65.—Basic-Lined Converter, indicating Tuyeres, Lining, etc.

Converter Linings.—The question of the lining has been the most important consideration in copper matte converting-practice.

The functions proper of the lining material are—

The employment of the lining material as a provider of suitable siliceous flux for the iron oxide, though until recently of vital importance for the practical operation of the bessemerising process, has been a necessary evil in many cases, and although it might have been a source of considerable profit under certain conditions, this function is unlikely in the future to be the consideration of greatest moment.

The vital requirements in modern converter practice are permanence of the lining and efficient means of effecting the fluxing of the iron oxide produced in the converting operation. The necessity for the frequent relining of converters involves not only heavy direct expenses, but it occasions waste of heat in the old linings, waste of material, loss of time, interruption of the processes, liabilities to outbreaks from the converters, and necessitates much heavy machinery for the conveying of vessels for relining, as well as large capital outlay in relining shops, plant, and appliances. In consequence, the employment of siliceous lining material as flux is usually a most expensive method of supplying the requisite silica; and so much is this the case, that an arbitrary limit to the iron contents of the matte has been rendered necessary, in order to prevent too much of the lining material being used up at a single blow. It was found cheaper to use other means of concentrating low-grade matte to a suitable grade for bessemerising—i.e., to flux off the excess of iron by means of silica in the blast-or the reverberatory-furnace processes.

Siliceous Linings.—Until recently, the only method for fluxing the iron in bessemerising, found practicable on a commercial scale, has been by the destruction of the siliceous lining, minimising the dead losses as much as possible by employing for the purpose siliceous materials from which values in the form of gold, silver, or copper could be simultaneously extracted and collected in the products of the operation.

Numerous attempts were made to effect combination of the iron oxides with silica introduced by some other method, but none met with success. Manhès blew sand through the tuyeres, and obtained as result a spongy unfused mass in the converter—whilst silica introduced in the form of lumps rose to the surface unchanged. In each case what silica was required for flux, was taken up from the siliceous lining. Experiments of a similar nature, in which basic linings were worked with, resulted in the fluxing silica being unabsorbed as before, whilst the iron which was in process of oxidation, not finding a suitable flux, became super-oxidised, resulting in the production of very infusible masses of magnetic or ferric oxides which rendered the process unworkable. Baggaley and others in Montana devoted much attention to experiments on different methods for introducing silica which would flux successfully, methods such as superheating or introducing silica held in suspension in fused silicates being tried, but without marked success, and for many years siliceous linings were necessarily worked with.

Owing to the large quantities consumed, the siliceous material must be obtainable cheaply and in abundant quantities. It should be high in free silica contents, since this constituent alone is effective as flux; it should have the property of binding well with clay or other material, so as to yield a rigid and impervious lining; and most important of all from the economic standpoint, it should carry values, since by this means only, could its destruction become an actual source of profit. At first barren quartz and barren clay were largely used for linings, but practice gradually developed in the direction of employing more profitable materials, and especially those from which the extraction of the values might present difficulties, in treatment by ordinary smelting methods. The practice as followed until recently at Anaconda is typical of such progress. Until 1908 the lining was chiefly made from highly siliceous ore obtained from Snowstorm, Idaho, carrying 80 to 85 per cent. of SiO2, 4 per cent. copper, as well as gold and silver, and a little iron and sulphur. This ore was crushed in mills and mixed with sufficient slime from the slime ponds of the concentrating plant to make a binding mixture. The slime, which carries about 60 per cent. of silica and also 2·5 per cent. of copper has excellent binding properties, owing to its clayey consistency. The proportions employed were 3 of siliceous rock to 1 of slime—no water was used, the mixture being almost dry to the touch. Since May, 1909, instead of employing ore obtained from outside sources, siliceous second-class Butte ore, which was formerly concentrated, has been very largely incorporated in the mixture used as lining material, it contains 65 per cent. silica, about 3·5 per cent. copper, a little gold and silver, and also iron and sulphur. The lining mixture consisted of 2·9 parts of this material with 1 part of slime. It was thought at first that owing to the greater proportion of sulphides and the lower silica content of the Butte ore, this lining mixture might prove inefficient compared with the former material, but with somewhat greater care in lining, it was found that very little more ore was required, and that tested by comparative silica contents it was more effective. Thus, where the former linings lasted for an average of six 7½-ton charges, equal to 20½ tons of copper per lining, the new ones last 5¼ such charges, equivalent to 17¾ tons of copper per lining, showing that although the efficiency per lining was reduced to 90 per cent., yet, calculated on comparative silica content, the new lining proved to be the more efficient.

The operation of lining is conducted with much care; the old lining is knocked away where necessary, rods are placed through the tuyere holes, and lining mixture is dumped in; 6-inch layers of material at a time being stamped down hard by means of an Ingersoll-Sargent tamping machine, until the lining reaches within 6 inches of the tuyeres. The wooden mould for the cavity, made up of a number of jointed pieces, is then placed in position, and the ramming of layer after layer round the sides is continued as before. The hood, inverted, is lined in a similar manner, it is then placed in position on the converter body and bolted down, a joint being made of moistened lining material. The whole operation takes about 1½ hours. The converter is then slowly dried by a wood fire, coal being subsequently added and kept burning under the action of a low blast for five or six hours; it is conveyed to the stand when required, dropped into position on the trunnion bearings, and the connections and adjustments very readily made.

The manipulation of relining at the Tennessee smelter is conducted in a very similar manner.

Basic Linings.—The all-important feature of the basic lining is its permanence, which, rendering the frequent relining of the converter unnecessary, allows of many economies in connection with capital outlay on plant and in operating costs. Further, owing to the lessened need for lining repairs, the frequent hauling of converters to the repair-shops situated at the further end of the buildings is avoided. This allows the employment of much larger converter units, with obvious attendant advantages, whilst it increases the ultimate possibility of continuous operation. Thus, the size at present employed, though the process has been in operation but a short time, is 26 feet by 12 feet, with a capacity of 35 to 45 tons of matte, and a daily output of 33 tons of copper from 40 per cent. matte. Such a converter, lined with 9 inches of basic material, will operate for 2,000 to 3,000 tons of copper before requiring repairs.

Keller’s report on basic linings in 1890 stated that they could not be employed successfully, because (a) basic material, being a good conductor, caused the outside of the converter to become too hot and the inside too cold; (b) such material broke up easily and so was unsuitable for use in permanent linings; and (c) even when basic linings were employed, the silica which was added as flux, refused to combine with the iron oxides. These views were very generally accepted for some years, until Baggaley’s persistent efforts and finally those of Pierce and Smith showed that by perfecting the constructional methods and details, by preventing heat losses as much as possible, and by operating on very large masses of hot material, the above difficulties could all be overcome and the basic lining successfully employed. The lining is of magnesia brick, and is 9 inches in thickness, except at the tuyeres, where the bricks are 18 inches thick. In the bottom of the converter and extending to within 18 inches of the tuyere level is placed a filling of ordinary firebrick, which is 13½ inches thick in the middle and 4 inches thick at the sides. The magnesite bricks are laid in dry magnesite powder, except near the tuyeres, where a mixture of magnesia and linseed oil is used. Expansion cushions of wood are inserted at intervals along the side of the fresh linings which are then “seasoned” with molten copper.

The required quantity of siliceous flux, as calculated, is now successfully introduced by dumping it into the converter, and pouring the matte charge upon it.

The Grade of Matte for Converting.—The grade of matte which is economically the most profitable to treat in the converter is a factor of great importance, since, if limits be fixed, the preliminary smelting stages for matte production are made less flexible, whilst in order to obtain matte of the correct grade, the smelting operations may require to be conducted at greater cost, or else additional smeltings for further concentration of the first matte may be necessitated—as is the case, for instance, in pyritic smelting at present.

The grade of a matte is usually expressed in percentages of copper, but from the standpoint of the practical converter operations, the proportion of iron is the factor which decides the suitability or otherwise of the matte for treatment, and since mattes may be regarded as mixed sulphides of iron and copper, a matte rich in copper is correspondingly low in iron contents, whilst a low-grade matte is high in iron.

The importance of the iron contents of the matte from the viewpoint of converter practice is due to iron being the chief source of heat in the operations, and to the fact that the iron oxide produced from it is the constituent which requires a supply of flux in order that the reactions may proceed and the process be successfully operated. The economic limit to the grade of matte suitable for the converter process is reached when it becomes less costly and more profitable to supply the required siliceous flux for the iron in the ordinary smelting furnace rather than in the converter. So long as the destruction of the lining was practically the only medium by which silica could be efficiently supplied, the limit to the iron contents of the matte was fairly rigid.

The bessemerising of a low-grade matte (low in copper contents, high in iron) entails the great advantage that a high temperature is obtained, owing to the fuel-value of the iron. On the other hand, however, grave disadvantages attend such practice, especially when working with the comparatively small quantities of material usually operated, and when employing siliceous linings. These disadvantages include the factors that—

In bessemerising a high-grade matte, the heat production is much smaller, owing to the decrease in the quantity of iron, which is the chief fuel of the process, and the limiting grade is quickly reached above which the bessemerising operation upon the matte ceases to be self-supporting.

In consequence, up to a comparatively recent date, a compromise has necessarily been effected, and the grade of matte operated upon has been such as to cause as much heat production as possible, together with the smallest practicable amount of fluxing action.

On these grounds, a matte containing from 40 to 50 per cent. of copper (equivalent to 32 to 22 per cent. of iron) has been found generally the most suitable. At several smelters, lower-grade mattes of from 32 to 40 per cent. copper-contents are converted most profitably, owing to such special circumstances as the profits resulting from the destruction of lining material, or in consequence of the fact that greater operating costs would be involved in concentrating the matte to a higher grade by the ordinary furnace-smelting methods.

In this connection, the successful adaptation of the basic lining by permitting the supplying of flux by means other than from the linings, has very important application and possibilities.

Owing to the frequent relining of these converters being then no longer necessary, mechanical difficulties of conveying the converter bodies to the relining shops are lessened, and larger converter units can now be employed, treating, even at the present stage of development, between six and seven times as large an amount of matte as formerly. By operating on such big charges, pouring off slag as produced, and adding fresh matte and flux without fear of destroying the lining, the difficulties attending the converting of low grade mattes have been successfully overcome.

The limit to the grade of matte economically suitable for the process will depend, in the future, chiefly upon the comparative costs of effecting the required concentration up to any desired grade, in the blast-or reverberatory-furnace, or in the converter.

The modern smelting scheme appears, therefore, likely to develop into the preliminary smelting of the ores by the cheapest method available, for matte of a grade best suited economically to the running of the furnace, the grade being independent of any rigid limit for the subsequent converting operations—the matte being then bessemerised as usual.

The Converting Process—Acid Lining.—There are two main stages in the converting of copper mattes. The first is essentially elimination of iron sulphide; the second, elimination of the remaining sulphur.

The product of the first main stage is a white metal, practically pure copper sulphide, the iron of the matte having been slagged off in the form of silicate, and the corresponding sulphur eliminated as SO2. The reactions during this stage are well known: the oxygen of the air blown in, yields oxides of iron and of sulphur, as well as some copper oxide. The latter, immediately reacting with iron sulphide which still remains, re-forms copper sulphide, with the production of more iron oxide. The iron oxides are fluxed by the siliceous materials present, forming ferrous silicate slags. The iron oxidation is productive of the greater part of the heat in the operation, and high temperature usually marks this stage of the process, which may be termed “the slagging stage.”

The flame which issues from the converter during this period is usually characterised by a green colour, caused apparently by the formation of iron-silicate slag.

When this stage is completed and the slag poured off, the white metal is blown up to blister copper—this constituting the second main stage of the process. The chief reactions are those of sulphur elimination and the production of metallic copper, caused by the action of some of the copper oxide first produced, upon the copper sulphide still present.

The flame during this period is small, thin, and fairly non-luminous, usually of a red-purple to bronze purple colour.[16]

The progress of the blowing from copper matte to white metal and thence to blister copper is usually indicated and controlled at the smelter by the appearance of the flame which issues from the nose of the converter during the first periods, and by the character of emitted shots during the later stages. This is particularly the case with mattes of moderate purity worked in the silica-lined converter. The successive changes in these indications are gradual, but are easily followed by the experienced skimmer, who is thus able to judge readily as to the manner in which the blow is progressing, and also as to the temperature, composition, and nature of the metal in the converter.

TABLE XII.—Changes in Composition during Bessemerising.

Time.Composition.
Hrs Mins Cu.Au.Ag.
%Oz.Oz.
11.52 amCharged No. 1 blast matte, .... 46·08 0·15 31·50
11.54 amBlow commenced.
12.04 pmSample No. 1 blowing,..1046·020·1731·80
12.14 pmSample No. 2 blowing,..2051·460·1835·80
12.18 pmPunched 3 minutes.
12.24 pmSample No. 3 blowing,..3053·270·2037·80
12.25 pmPunched 7 minutes.
12.34 pmSample No. 4 blowing,..4056·290·2140·80
12.40 pmPunched 2 minutes.
12.44 pmSample No. 5 blowing,..5059·900·2243·70
12.45 pmPunched 2 minutes.
12.54 pmSample No. 6 blowing,10062·670·2344·90
12.55 pmPunched 6 minutes.
1.04 pmSample No. 7 blowing,11067·890·2549·20
1.07 pmPunched 3 minutes.
1.12 pmBlow stopped.
1.13 pmSkimmed.
1.14 pmBlow resumed.
1.14 pmSample No. 8 blowing,12073·970·2754·90
1.17 pmPunched 2 minutes.
1.21 pmBlow stopped.
1.22 pmSkimmed.
1.25 pmBlow resumed.
1.25 pmSample No. 9 blowing,13077·820·2857·30
1.34 pmSample No. 10 blowing,14074·160·2654·30
1.44 pmSample No. 11 blowing,15081·720·1557·60
1.54 pmSample No. 12 blowing,20098·500·78107·70
2.02 pmPunched 1 minute.
2.04 pmSample No. 13 blowing,21098.570·4081.60
2.05 pmPunched 2 minutes.
2.08 pmBlow stopped, test for Cu.
2.09 pmBlow resumed.
2.10 pmBlow finished.
Converted copper,21699.080·3883·80
Total time punching,..28......
Total time of blow,216......
Actual time of blow,209......
Time.Composition.
Hrs Mins Insoluble Fe.S.As.
%%%.%.
11.52 amCharged No. 1 blast matte, ....0·15 24·30 24·700·22
11.54 amBlow commenced.
12.04 pmSample No. 1 blowing,..101·3023·7022·950·07
12.14 pmSample No. 2 blowing,..200·3020·5023·100·06
12.18 pmPunched 3 minutes.
12.24 pmSample No. 3 blowing,..301·1018·7022·150·06
12.25 pmPunched 7 minutes.
12.34 pmSample No. 4 blowing,..401·3016·2021·850·06
12.40 pmPunched 2 minutes.
12.44 pmSample No. 5 blowing,..500·9013·7021·950·06
12.45 pmPunched 2 minutes.
12.54 pmSample No. 6 blowing,1001·3011·4021·350·06
12.55 pmPunched 6 minutes.
1.04 pmSample No. 7 blowing,1100·65 7·60
1.07 pmPunched 3 minutes.
1.12 pmBlow stopped.
1.13 pmSkimmed.
1.14 pmBlow resumed.
1.14 pmSample No. 8 blowing,1200·25 3·4020·100·05
1.17 pmPunched 2 minutes.
1.21 pmBlow stopped.
1.22 pmSkimmed.
1.25 pmBlow resumed.
1.25 pmSample No. 9 blowing,1300·15 0·9019·600·04
1.34 pmSample No. 10 blowing,1403·30 2·6016·600·04
1.44 pmSample No. 11 blowing,1500·25 0·2015·350·04
1.54 pmSample No. 12 blowing,200 0·017trace0·780·050
2.02 pmPunched 1 minute.
2.04 pmSample No. 13 blowing,210 0·052 0·010·780·033
2.05 pmPunched 2 minutes.
2.08 pmBlow stopped, test for Cu.
2.09 pmBlow resumed.
2.10 pmBlow finished.
Converted copper,216 0·017 trace 0·010·033
Total time punching,..28........
Total time of blow,216........
Actual time of blow,209........
Samples taken each 10 minutes from beginning of blow until finished.

In general character, this colour sequence, during the bessemerising of the ordinary class of copper mattes—i.e., those consisting largely of iron, copper, and sulphur, with but moderate quantities of impurity—does not vary very markedly, but the body and luminosity of the flame depend to a great extent on the nature of the charge and on the working conditions. The colours are intensified by very hot metal, large charges, heavy blast, and rapid working, and particularly by the presence of secondary constituents, such as zinc, lead, or arsenic, which liberate dense white fumes, and so increase the luminosity of the flame.

There are generally four main variations in the appearance of the flame from the acid-lined converter:—


At commencement
of blow,		Oxidation of secondary
constituents,		Dark reddish-brown
flame.
Burning of iron, sulphur,
and coal,		Accompanied by much
smoke.
Slagging stage,Iron-sulphide oxidation,Apple-green flame.
White metal stage,Copper oxidation in
presence of slag,		White-blue flame.
Blowing to blister
copper,		Sulphur oxidation,Thin red-purple flame.

The changes in composition of the charge during a converter blow have been traced by Mathewson, who assayed samples during the various stages; some of these results are indicated in Table xii. and in Fig. 66. For full record see Trans. Amer. Inst. Min. Engineers, 1907.

In general, of the constituents present in the matte, iron and sulphur are removed very readily, 96 per cent. of the former and 53 per cent. of the latter in the slagging stage of the blow, whilst the elimination of the injurious impurities is high, bismuth and arsenic being removed to the extent of upwards of 90 per cent., and of the antimony, selenium, and tellurium, from 40 to 70 per cent. are eliminated ([see p. 217]).

Working of a Typical Charge in Silica-lined Converter.—The Anaconda converter plant is now being operated with basic linings. The former practice at this works was representative of the best type of acid-lined working, and the following description, based upon this practice, is typical of the method in general use. There were in operation twelve converter stands of the dimensions previously given. Normal working was to convert the 45 per cent. copper matte to white metal, to pour off slag, blow to blister copper, and pour the resulting metal—in regular sequence.

The colour-changes in the flame during bessemerising are indicated in the colour-photographs reproduced in the frontispiece.

Fig. 66.—Composition of a Charge during Bessemerising Operation.

Fig. 67.—Pouring Slag, Anaconda.

Seven to 8 tons of matte at an average temperature of 900° C. are charged into the converter, which is in an upright position with the blast on (16 lbs. per square inch). The operation of charging occupies three minutes. A few lumps of coal are thrown in, a vigorous action commences, copious and heavy white fumes and smoke and a full red to red-brown flame being emitted. The converter is now turned slowly back, so as to bring the tuyeres more completely under the charge and ensure more rapid and efficient oxidation, and the blow proper then commences. The flame drops for a time, continuing to be of a red to red-purple colour for two to eight minutes, after which, green commences to show in the red smoky flame (A), indicating that the first or slag-forming period of the blow is beginning. The green colour becomes more prominent and continues for 40 to 45 minutes (B). A preliminary pouring off of slag is then usually made, owing partly to the danger of violent or even explosive interaction which might otherwise occur between matte and slag, and also with the object of keeping down the copper losses in the slag by removing the greater portion of the latter at as early a stage as possible. The blowing is then continued. Flashes of blue now occasionally appear in the flame, and gradually increase in number until the flame becomes blue-white (C), which indicates that most of the iron has been slagged off and that the white metal stage is reached. The blue-white colour of the flame is to be attributed to the production of copper silicate, owing to the tendency of the copper oxide formed by the air blast at this stage, to flux off, and to produce the silicate rather than attack the copper sulphide. This formation of copper-silicate is particularly liable to occur in the presence of much slag and at high temperatures, factors which are well known to encourage this selective combination, and which prevail at this stage.

The blowing up to white metal takes about one hour.

Slag is then poured off again, until an iron rabble held under the stream commences to show signs of “metal” which give an appearance of spots of grease on the blade. The charge is then usually “doped.” “Dope” consists of highly cupriferous scrap, cleanings, slags, residues, also some siliceous material, added partly for the purpose of cooling down the charge which tends to become overheated at this stage. The converter is turned up again and the blowing is resumed in order to convert the white metal to blister copper.

The main reaction which now proceeds is represented by the equation

Cu2S + 2Cu2O ➡ 6Cu + SO2.

This stage of the blow also occupies about one hour or more, according to circumstances. It commences with a vivid red flame accompanied by smoke, but this soon dies out and a thin purple, almost colourless, flame results, which continues practically unchanged for the remainder of the blow (D). The temperature of the white metal is to some extent judged by the appearance of the flame, a red-brown colour indicating the correct temperature. If the colour be too red, the metal is too cool, and coal is thrown in; if the tint be too orange, the temperature is too high, and dope is added. Constant punching of the tuyeres by long steel chisels is required during this stage of the blow, owing to the lessened heat production due to diminution of iron, and also to the marked tendency for the liberated copper to chill round the tuyeres. The end of the blow is most difficult to judge, and although the size and colour of the flame offer some criterion, the usual and most important guide is the emission of small shots of copper which no longer stick to the hood situated above the converter throat, but which rebound from it. This is the stage where the skill and judgment of the skimmer are most tried.

When the blow is considered satisfactory, the character of the metal is further tested by pouring a small quantity on to the floor—a rugged and uneven surface indicating satisfactory metal. If poured too soon, the copper is coarse and impure; if poured too late, heavy losses in the slag result, owing to excessive oxidation of the metal.

The copper is then poured into a ladle, and conveyed to the refining and casting furnaces.

The whole operation for a straight run occupies about two hours, but the time required in general naturally depends upon the rapidity of working, and particularly on the grade of matte, and the volume and pressure of the blast.

The slags during the early part of the blow generally carry about 2 per cent. of copper, after the white metal stage is passed, they are usually much richer, on account of the intensely oxidising atmosphere which prevails, and the decreasing quantity of protecting sulphur. These later slags often contain upwards of 20 per cent. of copper, and in consequence as much slag as possible is poured off during the early stages of the blow, and the quantity towards the close is kept at a minimum.

The subsequent treatment of the converter slag depends very much upon the conditions of work at the smelter; at Anaconda, the iron contents of this slag are very useful in the blast-furnace charge, as there is a shortage of suitable basic flux for the silica of the rather siliceous charges. The slag is poured from the converters into ladles, and conveyed to a slag-casting machine, consisting of a conveyor belt carrying cast-iron moulds which are sprayed with cold water, the slag being thus cast into cakes suitable for the blast-furnace charge.

At Tennessee, the pyritic-smelting slags are already too ferruginous for any addition of irony converter-slags in the blast-furnace charge to be desirable, and the only metallurgical treatment for which these are suited is that of recovering from them the large amount of copper which they carry. The molten converter slag is, therefore, poured directly into the blast-furnace settlers, and by this means, the slags are cleaned and the values recovered.

At the new Tooele Smelter, under Mathewson’s organisation, the molten converter slags are poured directly into the reverberatory furnaces, there being no blast-furnace or settler plant, and the cleaning and settling are thus very satisfactorily conducted.

Systems of Working: Acid-lined Converter.—The “normal” system of working—i.e., blowing a matte-charge first to white metal, then to blister copper—is not always practicable nor economically the best practice, and the system of operating the charges depends largely upon the working conditions, which are subject to much variation at different smelters. Even at the same plant, the procedure has to be varied according to the attendant circumstances.

Conditions which may influence the system of working include:—

As instances of the way in which some of these circumstances affect procedure, the following examples may be quoted.

(a) When working with matte of low grade, especially in small quantities, as formerly operated, the loss of heat by radiation and by that carried away in the large quantity of slag produced, is very considerable, whilst towards the later part of the blow, the amount of sulphide fuel diminishes to such an extent that the maintenance of the desired temperature is difficult. The bulk of the final copper-product of the operation is very small and the metal is therefore liable to chill. In such cases, the system of “doubling” is useful. This consists of blowing the matte to the white-metal stage, pouring off the slag and adding a further charge of matte. This, on the resumption of blowing, restores heat and yields a charge of white metal sufficient to maintain the required temperature for the last stage of the blow, as well as affording a convenient yield of metallic copper.

(b) i. When working with a freshly-lined converter the charge is necessarily rather less than usual, owing to the smaller size of the cavity, and this results in a smaller yield of white metal, which is also colder. At the white-metal stage the slag is poured off, and the cavity having now become larger, owing to the fluxing action upon the lining, a fresh charge of hot matte is added, introducing fresh heat, further enlarging the cavity, and providing for a hot and plentiful supply of white metal for the blowing up to blister copper.

(c) ii. When the lining commences to wear thin, the converter may be retained solely for the purpose of blowing successive charges of white metal up to blister-copper, since owing to the very low iron content of white metal, there is little fluxing action on the lining during this stage, whilst the large quantity of white metal which can be operated in the enlarged cavity ensures a good supply of heat.

When linings burn through, the charge is transferred to another converter and the bessemerising finished there.

The management of the converters as thus indicated, and the distribution of the charges among the various converters are left to the head skimmer, who has control of the converter floor.

Working of the Basic-lined Converter.—The actual operations of bessemerising in the basic-lined converter differ but little from those where the silica lining is used. One important change has, however, been made, viz.: the introduction of the siliceous flux before the commencement of the blow. The lining having been heated up and “seasoned,” the charge of four or five ladles-full (30 to 40 tons) of matte is poured into the upright converter through the throat, 3 to 4 tons of siliceous flux, which must be well dried, are added, and the blast is turned on gently (at 5 lbs. pressure), whilst the converter is slowly turned back—these precautions being necessary in order to prevent excessive blowing out of the dry siliceous fines at the commencement of the work. When the silica is fairly well incorporated, the blast-pressure is increased to about 10 to 12 lbs. per square inch, the blowing is continued for 30 to 45 minutes, and after the silica has been fluxed by the iron oxides—which is tested by feeling the charge with an iron rod inserted through an opening in the breast—the converter is turned over and the slag poured off. A fresh charge of matte and a further quantity of siliceous ore are added and the blowing is resumed, these operations being repeated several times until the desired quantity of white metal has been accumulated, which is then blown up to blister copper in the usual manner. During the early stages of the blow, the operation is largely controlled by judging the quantity of iron remaining in the matte, from the appearance of small samples which are ladled out of the converter from time to time, and from this, the quantity of siliceous material required for the further fluxing is deduced. This material must be quite dry, so as to flux evenly and not form floaters. One of the advantages of the basic process is that siliceous ores containing values (the extraction of which may be profitable) which might not be suitable for use in siliceous linings, can be conveniently employed as flux in conjunction with the basic lining, though naturally the best work is done with flux containing a maximum of free silica. The character of the slag is not very different to that produced in the silica-lined converter, though it is usually lower in silica contents, and owing to the methods of frequent pouring, it is lower in copper values.

Special Features of Basic-lined Converter Work.—The basic-lined converter tends to lose heat by radiation and conduction more quickly than does the silica-lined vessel, due to the walls being thinner and the lining material a better conductor. Owing, however, to the use of larger charges, to the increased fuel value of the low-grade mattes, and to the larger blast-volume used, heat is retained sufficiently well for the successful operation of the bessemerising process. The temperature is, however, generally lower than that obtained when using the siliceous lining, and constant punching of the tuyeres is necessary—two men being required per shift for this work. The great advantages of the basic lining are connected chiefly with the fact that the frequent relining associated with the silica-lined converter is avoided, hence an extensive relining plant is not required, smaller building space and a lighter crane can be used. The use of basic linings further affords a means of extracting the copper and other values from siliceous ores which can be used as flux, but which might otherwise be difficult to treat, and it has made possible the cheaper treatment of low-grade mattes.

The disadvantages are chiefly those caused by

Converter Shop Organisation.—The introduction of the basic lining has, to a large extent, overcome the necessity for devoting so much shop space to the repair department, which formerly occupied a very considerable area. The converter stands are usually placed in alignment down one side of the building, the centre space is kept clear, and is commanded by the travelling crane for the conveyance of the ladles of matte, metal, or slag, to or from the converters. At Anaconda, the converters are charged from a train of matte-ladles mounted on bogies which run along a track behind the converters and situated some distance above them, the matte being poured down a launder which swings into position over the converter throat.

At Copperhill, Tenn., the converters are charged from ladles which are filled from the blast-furnace settlers situated at the other side of the furnace-building, whilst at the most modern large plant, at Tooele, Utah, the matte is run directly from the reverberatory furnaces to the converters along launders which are nearly 80 feet long and inclined at about 7 in 100. This method avoids all the handling of matte by cranes and ladles with the attendant troubles of skulls, breaks-down, spills, etc., and no difficulty has been found in keeping the channel free and open, nor in supplying matte at a sufficiently high temperature. At Anaconda and Tooele, the side of the converter-shop situated opposite to the converters is devoted to the refining and casting furnaces and to the slag-casting machines.

Modifications of Converter Practice.—(1) David’s Best Selecting Process.—David devised a special form of converter and suggested a method for conducting in the converter, instead of in the reverberatory furnace, the operations of the best “selecting process” on the principles of the old Welsh practice. The method embodied the converting of the matte somewhat beyond the white metal stage, by which means a small quantity of metallic copper was produced, in which the whole of the gold and silver values and most of the impurities collected, the remaining white metal being left tolerably pure. The metallic copper, thus obtained, was run into a side pocket in the lining and tapped from there, the rest of the pure white-metal was blown up to pure best-select copper.

Fig. 68.—General View of Converter Shop, Anaconda.

The method is, however, too specialised for ordinary commercial copper smelting, especially when electrolytic refining of the crude metal can be conveniently arranged for.

(2) The Haas Converter.—The Haas converter is spherical in form, and the tuyere holes through the lining are arranged at such an angle as to lessen the pressure required for the forcing of air through the metal. It is claimed for this form that it ensures better mixing of the materials and more even wear on the lining, by imparting a swirling motion to the bath.

References.

Douglas, James, “Treatment of Copper Matte in the Bessemer Converter.” Trans. Inst. Min. and Met., 1899, vol. viii., p. 1.

Baggaley, “A Brief Description of the Baggaley Process.”

Heywood, W. A., “The Baggaley Pyritic Conversion Process.” Eng. and Min. Journ., 1906, Mar. 24, p. 576.

Knudsen, E., “Pyrite Smelting by the Knudsen Method in Norway.” Mineral Industry, vol. xviii.

Moore, Redick R., “Copper Converters with Basic Linings.” Eng. and Min. Journ., 1910, June 25, p. 1317.

Editorial, “Improvements in Copper Smelting.” Eng. and Min. Journ., 1911, Mar. 4, p. 450.

Schreyer, Fr., “The Question of the Basic Bessemerising of Copper Mattes.” Metallurgie, 1909, vol. v., No. 6, p. 190.

“Improvements at the Washoe Smelter.” Mines and Minerals, 1910, April, vol. xxx., No. 9, p. 520.

Vail, R. H., “The Pierce and Smith Converter.” Eng. and Min. Journ., 1910, Mar. 12., p. 563.

Moore, Redick R., “Basic-lined Converter for Leady Copper Mattes.” Eng. and Min. Journ., 1910, Aug. 6, p. 263.

“Recent Practice in Copper Matte Converting.” Eng. and Min. Journ., 1910, Sept. 3, p. 460.

Neal, Carr B., “Further Data on the Basic Converter.” Eng. and Min. Journ., 1911, June 13, p. 964.

Keller, E., “A Study of the Elimination of Impurities from Copper Mattes in the Reverberatory and the Converter.” Mineral Industry, 1900, vol. ix., p. 240.

Mathewson, E. P., “The Relative Elimination of Iron, Sulphur, and Arsenic in Bessemerising Copper Mattes.” Bull. Amer. Inst. Min. Eng., 1907, Jan. 7, No. 13, p. 7.

Offerhaus, C., “Operation of an Anaconda Converter.” Eng. and Min. Journ., 1908, Oct. 17, p. 747.

Levy, D. M., “The Successive Stages in Bessemerising Copper Matte as indicated by the Converter Flame.” Trans. Inst. Min. and Met., 1910., vol. xx., p. 117.

Hixon, H., “Notes on Lead and Copper Converting.”

Semple, Clarence C., “Analyses of Converter Fume.” Eng. and Min. Journ., 1911, Mar. 11, p. 508.

Haas, Herbert, “The Vortex Copper Converter.” Eng. and Min. Journ., 1910, May 7, p. 972.

LECTURE IX.
The Purification and Refining of Crude Copper.

Preliminary Refining and Casting into Anodes— Electrolytic Refining—Bringing to Pitch, and Casting of Merchant Copper.

The further treatment of the converter metal depends to a large extent upon its composition, and the purpose for which it is intended. The matte-smelting operations on copper ores bring about the elimination of the greater part of the constituents accompanying the copper. The converter-grade matte may, however, in addition to the copper, iron and sulphur, also contain considerable proportions of easily reducible impurities of the ore, possessing a greater tendency to enter the matte than to be oxidised and eliminated in the slag. Such constituents may include gold and silver (practically all concentrated and retained in the cupriferous product), arsenic, antimony, bismuth, selenium and tellurium (retained to very considerable extent), as well as lead, zinc, nickel and cobalt (in much smaller proportions). The amount of these latter impurities ultimately retained in the converter matte depends very largely upon the proportions originally present in the ore, and upon the smelting conditions.

Under the strongly oxidising conditions of the Bessemer process the copper retains but small quantities of impurity, and those which do remain in ordinary converter metal may be broadly divided into two classes[17]—(a) those which are oxidisable with comparative ease, and (b) those which persist in the metal even under oxidising influences, unless treated by special means. The former include iron, sulphur, and zinc; the latter, arsenic, antimony, bismuth, selenium, tellurium, gold, and silver. Keller gives the following figures for the average elimination of the impurities in the converter:—

Iron,99per cent.
Sulphur,99"
Zinc, 99 "
Cobalt,99"
Bismuth,97"
Lead,96"
Arsenic,81"
Antimony,71"
Selenium,47"
Tellurium,  40"
Nickel,37"

Of the persistent elements, the retaining of the gold and silver in the converter-copper is a factor of much economic advantage, but the other impurities are curiously just those which are characterised by possessing most injurious effects on copper intended for electrical work—for which purpose most of the material is employed.

The demand for particularly pure metal in electrical and conductivity work therefore usually necessitates a further purification of the converter-copper (unless it be an exceptionally pure brand) and the production of metal specially free from the injurious constituents which persist to a small but sometimes very appreciable extent in the metal under the ordinary oxidising conditions. The presence of silver and gold in the copper may afford in many instances sufficiently good reason for a separating process independently of the market for the pure copper itself.

In modern practice, electrolytic methods are almost universally employed for the purification of the crude copper. By this means the large demands of the present day can be conveniently met, and the copper be obtained in a condition of remarkable purity. The frequent presence of gold and silver in the metal, and the convenience and completeness with which they are separated on electrolytic treatment of the copper are particularly advantageous features which recommend the adoption of electro-refining, and may in some cases be the reason for this procedure even though the metal might otherwise be already quite up to specification for electrical service. In the large majority of cases, these bullion-values constitute a welcome and independent bye-product, the returns from which may be set against the expenses of the refining operations on the copper, which might, in any case, be necessary.

The process may, therefore, be operated with one of the following objects—

Under the present industrial conditions, the electrolytic refineries are located at centres often at very considerable distance from the smelters. Situations for the refineries are chosen where the local conditions as regards power supply, technical resources, and particularly proximity to markets and distributing centres, allow of the operations being conducted under the most advantageous circumstances, and it is customary for smelters situated in the remoter mining districts to ship the crude copper to these custom refineries, instead of conducting the process themselves. At Anaconda, the well-equipped electrolytic refineries have been closed down, and the anode metal shipped to the Eastern refineries for treatment.

Preliminary Refining of Converter Copper and Casting into Anodes.—For modern electro-refining practice, the crude metal must be prepared into anodes, which are usually in the form of plates about 2 feet 6 inches × 3 feet by 2 inches thick. It is found that the metal as produced in the converters, on being cast into such plates, does not as a rule yield anodes which work satisfactorily in the tanks. This is largely owing to the impure and crude condition of the metal, which results in the production of plates which are spongy, coarse, and exceedingly rough and uneven on the surface. In consequence the direct employment of such metal would occasion irregularity and difficulties in the operation of the tanks, giving rise to short circuits, uneven wear, breaking off in large pieces, and similar troubles. Furthermore, the tank liquors and slimes become badly contaminated if large quantities of impurity be present in the anodes, and the deposition of good clean metal is thus greatly interfered with. All these reasons render it advisable that the converter-copper should, as a rule, undergo a preliminary furnace treatment before being cast into anodes.

The Anaconda practice is representative of the manner in which these preliminary refining and casting operations are conducted, except that the enormous scale and organisation of the operations are practically unique. The principles involved and the general method of operation are in all essentials those of the old Welsh furnace-refining process.

The Furnace.—The finished metal from the converters is teemed into ladles, and from these is poured directly into one of three casting furnaces. Two of the furnaces are in constant use, one of them engaged in refining, one being filled, and one in reserve or repair. Two of the furnaces are 14 feet × 22 feet 8 inches hearth dimensions, with a capacity of 95 tons; the third has a 14 feet × 28 feet hearth, and a capacity of 110 tons—the fire-boxes being 5 feet 6 inches × 7 feet. The furnace bottom is constructed of the local silica brick (which is claimed to be the finest in the world) laid down in four beds, the three lower being each 12 inches thick, whilst the working bed is constructed of 20-inch bricks; brick being found to be better than sand in this class of work. The bottom is curved to a depth of 2 feet. These furnaces are, in consequence of their different function, constructed on somewhat different principles to the reverberatory smelting furnaces.

Fig. 69.—Sectional Plan, Elevation, and Transverse Sections of
Refining and Anode-Casting Furnace, Anaconda (Peters).

Owing to the high conductivity of copper, and to the fact that the functions of the furnace are either largely as a medium for simple fusion or as a receptacle for molten metal, and further, that but little slag is produced, that no settling and separation of the fluid materials are required, and that there is no danger of dusting-losses, the furnace may conveniently be built with a deep hearth which need not be of very considerable length. The main requirements are refractoriness of the building materials, particularly careful construction so as to avoid breakouts, and very strong bracing indeed on account of the deep and heavy bath of material which is carried on the furnace hearth.

Operations—(a) Oxidation Stage.—The furnace is loosely filled with scrap copper which has accumulated round the works (8 to 12 tons), and converter metal (of composition say about 98·3 per cent. copper) is then poured in at the side door from ladles bringing it in quantities of about 5 tons at a time, as teemed from the converters. When the furnace is about half-filled, a blast of air at 90 lbs. pressure is injected through the metal by means of iron pipes, which at this stage just dip below the surface. These pipes are gradually eaten away by oxidation and slagging action, but as the end wears down, the pipe is pushed further in. The function of this air blast is to supply oxygen for the purpose of acting upon the small quantities of oxidisable impurity which remain in the metal after bessemerising, and which consist chiefly of iron and sulphur, in addition to the small quantities of metalloids. The oxygen partly acts directly on these constituents, but as already indicated, the scouring action is to a great extent performed by copper oxide which is produced and which is itself a powerful oxidising agent. The iron appears to be one of the first elements to be removed, and then a little sulphur, but this is chiefly eliminated after the iron has been oxidised. The interaction between the copper oxide and the sulphides liberates metallic copper and yields SO2, which bubbles up through the metal and gives to it an appearance of “boiling,” by which name this stage is known. Too rapid an oxidation during the early stages is dangerous if much sulphur be present, owing to the evolution of sulphur-dioxide assuming a degree of explosive vigour. Up to this point, the oxygen has been utilised in removing iron, sulphur, etc., which are eliminated as oxides, so that but little of the oxygen is retained in the metal, but after the boiling stage is passed, oxygen is actually absorbed, the copper now becoming oxidised, and the oxygen contents of the metal rapidly increase. As in the analogous instance of steel bessemerising, it appears essential to introduce some excess of oxygen into the metal in order to ensure the complete removal of the oxidisable impurities, so in copper-refining, an excess which amounts to about 0·7 per cent. of oxygen (equivalent to about 6 per cent. of Cu2O) must be introduced.

In the refining practice as conducted by the Welsh process, much of this aëration took place during the slow melting down of the crude blister copper, and subsequently during the flapping operations with the rabble; but the use of the air-blast hastens this oxidation considerably, especially as the metal is now often directly poured into the furnace in a molten condition, so that oxidation during melting is not possible. It is essential to defer this final oxidation and elimination until it can be conducted at the refining furnaces rather than to attempt it in the converter, since the refining furnace allows of the operation being performed much more gradually and under better control, whereas if conducted in the converter, the necessarily vigorous action would occasion unduly heavy losses of copper in the slag and probably excessive oxidation of the metal.

During the aëration, the furnace contents are continually added to, by additional ladles-full of metal, and usually by the time the furnace is filled the air-blast has oxidised most of the impurities from the metal. These have entered the slag, and the copper has become “dry,” owing to the necessary super-aëration. If this stage has not been reached, the sample often shows “sprouting” (also known as “spewing” or “throwing a worm”), which is caused by the escape of SO2, and indicates that all the sulphur has not been eliminated. In that case the blowing is continued until small samples ladled out from the bath exhibit the characteristics of dry copper, viz.: the depression down the middle line of the ingot, brittleness of the metal, and a purplish brick-like fracture.

These preliminary operations may occupy some three or four hours or more.

(b) Poling and Bringing to Pitch.—The oxidation having proceeded to the required stage, the highly cupriferous slag is skimmed off, after being first thickened with ashes from the fire-grate, and the poling of the metal is then commenced. This operation is conducted by immersing poles of timber, three to six at a time, in the metal, holding them well under the surface and pushing them further in as the ends burn away. It is essential that the timber should be green and not dry, and preferably it should be hard wood, such as birch, beech, or oak. The poles are usually as long as possible, and are from 6 to 8 inches thick at the butt end. The function of the wood, particularly during the early stages is, to a great extent, mechanical, and any chemical changes effected are by indirect action.

The poling operation really consists of two stages, the first of which is the final elimination of SO2 retained by the metal, and the last, the actual reduction of the excess oxide and the “bringing of the metal to pitch.”

The green timber, when inserted into the copper, liberates large amounts of moisture and reducing gases which agitate the bath considerably and “shake” the gas out of the metal more or less mechanically, replacing, at the same time, some SO2 by CO and hydrocarbons which copper possesses the power of absorbing. When the SO2 has been satisfactorily eliminated, the reduction stage is arrived at, and this is conducted in a manner similar to the familiar poling operation of the Welsh process. The surface of the bath is completely covered over with a layer of coke, anthracite, or charcoal, and more poles are inserted. The exact mechanism of the operation has not yet been definitely traced, but the action of the wood at this stage is partly of a mechanical and partly of a chemical nature. The reducing gases liberated by the charring and destructive distillation of the wood have themselves a reducing action on the oxides which are dissolved in the dry copper, but an important feature of the action of these gases is the agitation and splashing which they occasion, thus bringing the molten metal into close and vigorous contact with the layer of reducing carbonaceous material maintained upon the surface of the bath.

Poles are inserted usually two or three at a time, and samples are constantly taken and examined for surface indications and for fracture. This preliminary refining operation usually has for its chief object the preparation of a fairly pure metal which will yield a sound, clean, and even anode casting, and which is not required at this stage to pass the rigid mechanical tests essential for the market product. In this case it is therefore usual to carry the poling operation only to such a degree that the samples ladled out and cast into small ingots solidify with the even, smooth surface desired and which is characteristic of “tough-pitch copper”—irrespective of any special mechanical properties of the metal. If the test is satisfactory, the metal is ready for casting.

The poling occupies some hours, and usually from 40 to 50 poles of wood are used up before the metal is in a suitable condition for casting. During these operations the coal fire in the grate is manipulated in a manner best suited to the various stages of the process; there may thus be an oxidising flame during the early part of the refining, but the flame must be of a reducing character whilst poling is in progress.

Casting.—Until comparatively recent years, the size of the refining furnace has been necessarily limited to small dimensions, owing to the difficulty in emptying the furnace of large charges. The practice, as conducted hitherto, has been based on the familiar method of the old Welsh process, viz., that of ladling out the metal by small hand ladles. This involves so much hand labour, and requires such a long period of time for its operation as to make practically impossible any attempt to deal with large quantities of metal, or to lead to any considerable increase in furnace capacity. The chief difficulties to be overcome when operating on large charges of copper by this hand-lading method are those of maintaining the metal at the correct pitch during the lengthy period of ladling; whilst the large amount of time during which the finished copper has to remain within prevents the furnace being used for its chief purpose, that of refining more metal.

The method of hand ladling was employed for so many years on account of the difficulties of controlling the stream of metal and of tapping the furnace in the usual way—i.e., through a tap-hole at the lowest point of the bath. These difficulties were due to the very high working temperatures, to the great weight of metal behind the stream, which forced it out under great pressure, and to the high melting point, conductivity, and tendency to chill of the copper, which was apt to cause setting of metal in the tap-hole, and led to the latter becoming rapidly closed up and useless. Regulation of the stream of metal to a gentle flow was impossible under such conditions.

With the introduction of casting machines by Walker, and the improvements in the methods of tapping by the adoption of the vertical tapping-slot, these difficulties have been removed, and the casting of 100 tons of metal from one of the modern large casting furnaces presents, to experienced workers, little practical difficulty.

The modern casting machine brings a series of moulds continually under the supply of metal which issues from a large ladle fed continuously from the tapping-slot of the furnace.

The method of tapping now used is to allow the copper to gently run out of the furnace, by gradually lowering the level of a temporary retaining wall which is constructed in a narrow vertical slot in the tapping side of the furnace.

Fig. 70.—Indicating Tilting and Pouring Mechanism
for Ladle of Casting and Refining Furnace.

This slot, which is

-shaped in plan, extends from the lowest point of the hearth to well above the highest possible level of the liquid metal. It is about 3 feet high and 4½ inches wide, and whilst the furnace is working it is kept rammed with a mixture of loam and anthracite, this filling being supported by a series of short transverse bars, 16 inches long and 1 inch square in section, which are set 3 inches apart and rest upon lugs fixed to the iron plates which strengthen the furnace-wall. During the operation of casting, this hard filling can be readily cut away as required and the level of the dam thus gradually lowered at will, permitting the gentle and continuous overflow of the molten metal. The stream is also regulated by inserting a pole of wood in the opening, should the flow become too rapid, and by this means it is kept under absolute control. The molten metal flows along a spout which feeds a small suspended ladle of about 800 lbs. capacity, the supply being so regulated that this ladle is filled sufficiently slowly as not to get ahead of the moulds. The ladle is supported hydraulically, and is pivoted so that it can be brought forward and tilted for pouring, and then lowered and moved a slight distance backwards, to allow the next mould to come into position.

On tilting the ladle, the metal flows gently and without splashing through a three-hole grid in the front—which keeps back slag or cinders—and runs into the mould, which is rapidly filled. In order to prevent the metal overflowing in the mould, and also to rapidly cool that portion which forms the lugs of the anode-plate, a hollow water-cooled block 2 feet 6 inches long and of 6 inches square section, situated opposite the ladle, is brought forward hydraulically into such a position that it rests on the mould just against the edge of the lugs.

At many smelters the circular form of anode casting machine introduced by A. L. Walker is employed. This apparatus consists of a horizontal wheel which can be rotated slowly, carrying a series of arms at the end of which the moulds are supported, so that they form a broken ring. By the rotation of the machine, one mould after another can be brought under the ladle and filled. The moulds are pivoted, so as to allow of tilting, and when the metal has set, the ingot is thus dropped into a cold water bosh, whence it is carried to the yards by a conveyor. At Anaconda, the casting machine consists of a series of moulds carried on a platform conveyor which is operated hydraulically—the moulds are attached by bolting them on to the belt through lugs fixed underneath. The moulds are constructed of 1 inch cast-iron, and allow of the production of anode ingots 2 feet 6 inches × 3 feet by 2 inches thick, provided with lugs at the corners of one end for the purpose of supporting the plates in the tanks.

Each mould holds about 560 lbs. of metal, and when the anode has been cast, the ladle is dropped back into position and the mould is moved forward by means of the conveyor belt. After traversing a distance equal to three times its own length, the ingot becomes fairly solid, and at a point corresponding to this position the conveyor base inclines slightly upwards. The cake is sprayed gently during its passage over a distance of about 8 feet, the conveyor belt then passes over a pulley-wheel, and when in a vertical position, the anode is forced out of the mould by a crowbar and falls into a water bosh, from which it is carried by another conveyor on to a platform. Here it is wheeled to stacks, examined for flaws, and weighed. Sample anodes are placed on one side, and the others are packed for shipment to the Eastern refineries.

Fig. 71.—Walker’s Anode Casting Machine.

Fig. 72.—General View of Tank-room of Electrolytic
Refinery, Perth Amboy, N.J.

The furnace deals with one charge (usually of 100 tons, but occasionally much more) per eight-hour shift, and the casting machine yields 25 tons of anodes per hour.

Samples weighing from 4 to 6 ozs. are taken three times per shift from the stream of copper running into the moulds, by batting the metal into water with a wooden paddle. This method checks very well with drillings taken from the anode plates, the chief discrepancy feared having been with respect to silver contents, owing to the tendency of this metal to segregate. The assay of the anode metal at Anaconda averages copper 99·3 per cent., silver 80 ozs. per ton, and gold 0·5 oz. per ton.

Electro-Refining.—Electrolytic refining was introduced on a commercial scale by Elkington at Pembray in 1865, and with the general adoption of the dynamo for the production of power, dating from about 1870, the process was greatly developed. Most of the copper now placed on the market has passed through the electrolytic refinery.

System of Working.—The method of arranging the electrodes in the depositing tanks which is usually adopted at the great refineries at the present day, is that known as the parallel or multiple system.

In this method of working, the anodes are all connected to one pole of the circuit, and the cathodes, situated between them, are all connected to the other. In this way, each tank comprises in reality one large anode and one large cathode, and the voltage as measured between any two neighbouring electrodes will be the same. The system thus allows of currents at low voltage being employed, since the voltage is a factor of the number of electrodes in series, and in consequence danger of short circuiting is lessened. This allows of plates being placed closer together in the tank, with less danger from this source of trouble.

A large number of tanks are employed at the refineries, and they are usually arranged in series, the anode plates of one vat being connected to the cathode plates of the neighbouring one, the current thus passing from one vat to the other through the entire system.

Various other methods of arranging the electrodes have been favoured from time to time, and of these the series-system is the most important, this being still in use at several large refineries, though it has been generally superseded by the multiple method.

The plan underlying the series method was that of avoiding the trouble and expense of preparing and working with the special cathode sheets of pure copper as are necessitated by the multiple system. In the series-method, each anode was made to serve as a depositing surface for the pure cathode copper produced by the operation, so that as impure anode copper was dissolved away on one side of the “anode”-plate, pure copper was gradually deposited upon the other side. This system appeared, therefore, to have several marked advantages to recommend it, but in practical operation many difficulties in working and several serious disadvantages were encountered. The chief points in favour of the series-system are—

(a) Smaller first cost of the installation, particularly in the matter of electrical connections, since the multiple system requires heavy leads running along each side of the tank, as well as close attention to the providing of good contacts, in order to connect all the anodes and all the cathodes together with a minimum of current leakage. In the series-method, the plates are readily connected one to the other.

(b) The great saving of the cost of preparation and arranging for specially pure cathode plates, this constituting a very important factor in the costs of the multiple process.

(c) The output of metal per vat is greater.

On the other hand, the disadvantages of the system, except under special conditions, are very serious.

(a) More scrap is produced and requires re-treating, owing to the difficulty of separating the new deposit from the remaining portions of old anode, which often adhere very firmly.

(b) Higher voltage through the tanks is required, owing to the large number of electrodes in series in the bath. Hence the danger of current-leakage and short circuiting is greatly increased, especially when impure anodes are used, since they tend to produce conducting layers of mud on the bottom of the vat.

(c) The anode plates have to be made particularly smooth and even on the surface, since in order to lessen the voltage required, the plates are brought as close together as possible, in consequence of which, any excrescences upon the surface greatly increase the danger of short circuiting. Preliminary furnace-refining and special straightening of the anodes are therefore essential in connection with the series method.

(d) Special tanks are required, as the protecting lead liner cannot be employed, since the danger of current-leakage through it is increased, owing to the higher voltages required. Hence special acid-resisting material, such as slate, is necessary, the expense of which is considerable.

(e) The cost of stripping the cathodes is high, and the operation is often difficult.

(f) The cost of maintaining the plant is greater.

The special advantages of the multiple system are that—

It is, however, general to carry out this preliminary refining, which yields sounder anodes, keeps the electrolyte purer, and promotes the more regular working of the electrodes and electrolyte—although some smelters still cast the anodes direct from converter metal.

Summarising, it is found generally that—

Outline of the Process.—The electro-refining industry is a highly specialised one, and the methods of putting the comparatively simple underlying principles into practical operation have assumed great complexity and diversity in detail, concerning which Ulke has collected and published much valuable information.

The following details have more particular reference to the multiple system of working, as being the most representative of the electro-refining methods in general use.

The outlines of the process constitute the passing of direct electric current through tanks containing acidified solutions of copper sulphate, employing plates of crude copper as anodes, and depositing pure metal upon cathode-plates of specially-refined thin sheets of pure copper. The precious metals and most of the impurities of the anode metal are liberated as small insoluble particles which gradually settle to the bottom of the tanks in the form of mud, soluble constituents, such as iron and zinc, first passing into solution.

General Conditions—Anodes.—The usual dimensions of the anode-plates are 3 feet high by 2 feet 6 inches wide and about 2 inches thick; they are generally cast with lugs, so as to allow of suspension in the tanks. The anode metal is usually brought by a preparatory operation, to as high a state of purity as is economically practicable—

(a) The necessity for the employment of solid and even anodes has already been indicated; it allows of closer suspension of the electrodes, lessens the liability of sprouting and unevenness on the deposits and the irregular wear and breaking up of the anode-plates before they are sufficiently worn away.

(b) The more free the metal is from iron, sulphur, zinc, nickel, etc., the purer remains the electrolyte, since these elements pass into solution at a greater speed than does the copper itself, and, gradually concentrating in the tank liquors, render them more and more impure—the purity of the metal deposited at the cathode being in consequence decreased. The preliminary refining and bringing up to pitch of the metal before casting into anodes, as already described, thus has for its object the preparation of electrodes in a suitable mechanical as well as chemical condition. The copper content is rarely less than 98 per cent. and is often more than 99 per cent. The gold and silver contents are not affected by this preliminary treatment, nor are, to any great extent, the proportions of arsenic, antimony, bismuth, etc.

The size of the anode-plate varies somewhat at different refineries, the usual standard dimensions being indicated above; the size depends to a large extent upon the facilities for handling the electrodes and on the circuit system operated. There is a tendency at several works possessing suitable facilities, to increase the size of the electrodes.

The Cathodes.—The copper is deposited from the electrolyte upon cathode sheets, which are usually thin plates of pure copper corresponding in size to the anodes. As these sheets cannot be conveniently provided with suitable lugs for suspension, they are usually made of somewhat greater length than the anodes, so as to allow of bending over the cross-conductors; otherwise they are furnished with metallic clips for attachment to these bars.

These cathode sheets are prepared by depositing layers of pure metal upon plates of refined copper of suitable surface dimensions and of about ¼ inch thickness. Each side of these plates, which are specially smoothed, is first slightly oiled so as to allow of the subsequent convenient stripping of the sheet when made, and it is then well coated with graphite in order to present a conducting surface on which deposition can proceed. The cathode-sheets are deposited either in the regular tanks of the refinery or in vats specially devoted to the purpose, using in that case, pure electrolyte, and working in the usual manner. On attaining a thickness of about 1⁄25 inch, the sheet is stripped off, cleaned and clipped ready for use as a cathode.

The Electrolyte.—The electrolyte is essentially an acid solution of copper sulphate. The average proportions are from 15 to 20 per cent. of copper sulphate crystals, and from 5 to 10 per cent. of sulphuric acid—the usual density of the solution ranging from 1·12 to 1·25. The liquid under ordinary conditions of working, remains reasonably pure for a considerable time. It tends, however, to decrease in acidity and to increase in copper contents, partly owing to the presence of cuprous oxide in the metal, which passes into solution independently of the indirect transference of metallic copper from anode to cathode. The composition of the tank liquors must, therefore, be frequently checked. The gold and silver values do not pass into solution under ordinary working conditions, and the addition of a small quantity of common-salt or of hydrochloric acid to the vat effectually prevents any silver from remaining in solution in the liquors.

A considerable proportion of the arsenic in the tough-pitch anode-copper, existing as arsenate, is deposited with the mud residues, it being insoluble and non-conducting. Arsenic in a reduced condition is, however, soluble, and may gradually concentrate in the liquors and contaminate the cathode copper, unless suitable precautions are taken. Some of the reduced arsenic, moreover, tends to produce a slimy arsenite of copper, which, though insoluble, exists in a colloidal non-settling form. Addition of ammonium sulphate to the electrolyte prevents this formation, whilst combined aëration and heating promote the precipitation of arsenic as insoluble arsenates which settle with the tank slimes.

Some of the antimony and bismuth tend to first pass into solution, but for the most part they are precipitated as insoluble basic salts. Under suitable conditions with respect to the acidity and copper contents of the electrolyte, there is little tendency for deposition of these impurities with the copper, but deviation from the correct composition is liable to cause contamination of the deposited metal. These impurities, when in solution, tend to be oxidised by aëration, and this operation greatly encourages their precipitation with the mud.

Iron readily dissolves in the electrolyte, forming soluble ferrous sulphate which tends to gradually accumulate in the solution. This contamination spoils the quality of the deposited metal, and interferes with the process of deposition, decreasing the conductivity of the bath and thus necessitating higher voltage. Aëration of the solution, especially when warmed, leads to the formation of basic ferric sulphates which are insoluble, and which therefore accumulate at the bottom of the tank.

Selenium and tellurium, which when present most probably exist as insoluble selenides and tellurides of copper or silver, are also precipitated, and thus do not find their way into the deposited metal.

The copper itself is deposited from the electrolyte on to the cathode-sheets by the action of the current, whilst at the anodes, the metal passes into solution, and the other constituents are either dissolved or precipitated. It follows that in an undisturbed solution, the liquid near to the cathode becomes gradually impoverished in copper, resulting in a decrease in the rate of deposition and necessitating greater electrical pressure, whilst in the neighbourhood of the anode, the liquor is proportionately stronger in copper and less acid in character. Should these conditions continue to any great extent, the working of the bath is seriously interfered with, since diffusion proceeds too slowly for uniformity to be restored, and in order to secure uniform composition of the electrolyte, it must be maintained in gentle motion by some system of circulation. This agitation and mixing is assisted by the aëration of the bath for the purpose of hastening the oxidation and precipitation of several of the impurities—this being effected by blowing through the liquid a gentle supply of air.

Temperature.—The electrical resistance of the solution decreases as the temperature rises, and in practice the bath is maintained at a uniform temperature of 45° to 50° C. By this means a useful increase in conductivity is obtained, the strength of the deposited copper being at the same time greatly augmented.

Electrical Conditions.—The electrical factors which mainly control the working of the electrolytic process are those of—

(a) Current Density.—The quantity of metal deposited from an electrolyte is proportional to the current which passes, and to the electro-chemical equivalent of the particular metal. Thus a current of one ampere will deposit from copper sulphate solution, 1·1832 grammes of copper per hour, and the total quantity deposited in any given time is determined by the product of the current, the time, and this electro-chemical equivalent (which is determined experimentally).

In practical operation, a factor which is amongst the most important of those governing the working of a plant, is the current density, or current per unit of area of depositing surface, since from this factor the rate of deposition upon the cathode plates is determined, and from it the power requirements, accommodation, etc., for the plant are fixed. The current density is subject to wide variation, but, as a general rule, it ranges from 8 to 18 amperes per square foot of plate-area. Its value is largely dependent upon the speed of working, the cost of power, and the composition of the anode metal, the electrolyte and the desired product, etc.

In general, high current-density possesses many advantages, resulting from the fact that it occasions a more rapid deposition. It causes a proportionately greater output, consequently the stock of metal held back in the tanks is reduced, and hence there is less capital locked up in the form of metal undergoing treatment, and less plant and accommodation are required for the same output.

The current density permissible is, however, limited by the composition of the electrodes and the solution. High current-density causes rapid dissolution of the anodes, and if the plates are not particularly free from impurity, the electrolyte rapidly becomes contaminated, since its dissolving power on the impurities becomes greater with increased electrolytic action, and this affords less opportunity for the precipitation and settling of the injurious constituents. In consequence, the cathode copper is contaminated through this mechanical inclusion of impurities, whilst electro-deposition of some of these materials may also be encouraged. The presence of much silver in the anodes causes the rapid breaking-up of the plates, especially if the current density be high, and thus the separation of the values in the slimes is not so efficiently managed. With high values in the anode copper, it is necessary to reduce the current density to 8 or 10 amperes per square foot, whereas with purer metal a density of as much as 16 to 20 may be conveniently employed.

(b) Voltage.—Electrical pressure is required in order to force the depositing current through the electrolyte against the resistances in the circuit. The voltage required depends upon the current density, the composition and temperature of the electrolyte, the composition of the anodes, and also upon the general conditions of working. These being constant, the voltage necessary is largely a factor of the number of electrodes in series in the tank and of the distance apart of the plates. Under ordinary circumstances this voltage varies from 0·1 to about 0·3 volt. High voltage is to be avoided, owing to the danger of short-circuiting, especially in cases where the accumulation of mud in the tanks, or the impregnation of metallic salts in the tank walls, or the growth of excrescences upon the plates, lead to the passage of the current through these conductors rather than through the electrolyte solution itself. Short circuiting naturally diminishes the output of the plant.

These electrical factors which form the basis of the power requirements of the refinery, call for careful observation during the progress of the operations in order to ensure successful working and a high efficiency of the plant.

The current from the dynamos is brought by heavy leads and is distributed through the sets of tanks in the manner best suited to the installation. At one of the newer works, dynamos producing about 6,000 amperes at 120 to 150 volts, supply sufficient current to operate a set of 400 vats working on the lines just indicated.

Fig. 73.—Indicating Methods of Suspending and
Connecting Electrodes (Perth Amboy, N.J.).

The Depositing Tanks.—The tanks are usually constructed of wood, such as strong pitch pine, and they are lead-lined. The cross-section is usually such as will allow a space of about 3 inches between the edges of the electrode and the wall, and a 6-inch space from the tank-bottom to the lower edge of the plate. The length of the vats varies considerably, according to the desired output and to convenience of working—10 to 15 feet being average dimensions. This size of vat will hold 15 to 25 anodes together with a corresponding number of cathodes (16 to 26). The tanks are arranged across the building in a number of rows which are usually stepped down in stages of about 2 to 3 inches each, so as to assist as much as possible the circulation of the electrolyte through the system by gravity, and the vats are set in pairs with aisle-ways between ([see Fig. 72]).

Leads run along each side of the tank, the current being conveyed to all the anodes at once by resting one lug of each plate upon the lead which runs along one side of the tank, carefully insulating the other lug from the conductor situated at the opposite side, this being used for connecting up the cathodes. The cathode-sheets are suspended from metallic cross-bars, which rest upon their own conducting lead, and are carefully insulated from the anode lead at the other side. The solutions are heated by means of steam coils.

Distribution of the Electrolyte.—The necessary circulation of the electrolyte is effected as much as possible by the natural action of gravity. The tanks of the top row in the depositing-house receive a constant supply of fresh solution from upper distributing vats, whilst old electrolyte is drawn off from near the bottom of the tanks, and flows over to those on the next and lower level. Fresh solution thus enters at the top of the tank, old solution is drawn off from below, and thus a uniform density and composition is maintained. From the tanks situated at the lowest level, the solution passes to a well, and from there is pumped up to the store-tanks or, when necessary, to the purifying tanks; air-pressure pumps being often employed for this work.

Fig. 74.—Indicating Connections for Circulation of Electrolyte (Barnett).

In course of time—and under the modern system of working with moderately pure anodes, this period is of considerable duration—the gradual accumulation in the electrolyte, of the small quantities of impurity which are dissolved from the anode, may render the liquid so impure, that a danger arises of contamination of the cathode copper to such a degree that it becomes unfit for conductivity work. It then becomes necessary to purify the solution. In present-day practice, this continued accumulation of impurity in the electrolyte is prevented by continuously withdrawing, for separate purification, a certain proportion of the electrolyte from the circuit—replacing it by a fresh supply of pure solution from the store-tanks. Constant regeneration, purification and circulation are thus effected, whilst uniform composition is maintained.

After considerable use, the electrolyte solution gradually tends to increase in copper contents, and the first stages in the scheme of treatment for the old solution is to recover this excess of copper, which is effected in tanks known as “liberating tanks.” These are similar in general features to the refining vats, except that lead plates are employed instead of the copper anodes, so that the excess metal is deposited without any addition of copper being made to the solution, from the anodes. In due course, the desired composition in the electrolyte is once more attained.

When the solutions have become too impure for further use in the tanks, the bulk of the copper sulphate is recovered by evaporation in large pans, followed by crystallisation in somewhat shallow vats of large dimensions. The crude blue vitriol is further purified by repeated crystallisation, and any copper which still remains in the solution is then precipitated on scrap iron, the cement copper being worked through the furnaces again. Excess acid is also often recovered on further evaporation of the liquors, and is employed in the subsequent treatment of the slimes.

Working.—In the large modern refineries, the anodes are carried to and from the tank-house by cars, and at the tank-room are suspended from frames which are conveyed over the baths by means of overhead electrical cranes of about 10 tons lifting capacity. These rectangular frames correspond in size to the dimensions of the tanks, and are constructed of steel girders. Under the longer sides of this frame a series of hooks project, upon which the lugs of the anodes rest, and the hooks are placed at distances corresponding to the eventual position of the plates in the tank, so that the whole series of anodes can be dumped into position at one operation.

The cathodes are placed in a second rack, and likewise brought into position, between the anodes. The solution is then turned into the tank, the current started, and the refining proceeds, with a steady flow of liquid circulating through the system. The operations of changing electrodes, cleaning and reloading occupy about one hour, and, but for this manipulation, the process under normal working is continuous. In ordinary practice, about 20 to 25 lbs. of copper are deposited daily on each cathode. Constant examination is made as to the electrical conditions, and the composition, temperature, and density of the solutions.

The anodes usually remain in the bath for a period of about six weeks, and they are then removed from the tank, scrubbed, and sent back to the furnaces to be remelted and re-cast into fresh anodes, the quantity of such anode scrap under good working conditions amounting to about 9 or 10 per cent. of the original metal.

Fig. 75.—Tank-house, showing Anode Crane (Ulke).

The cathodes remain in the tanks for about one week, by which time a deposit of from 150 to 170 lbs. of pure metal has been obtained upon each. The practice of frequently replacing the cathodes possesses, among other advantages, those of maintaining a more even current density over the plates, of preventing the growth of excrescences and the irregular dissolution of the anodes, and of lessening the danger of breakdown of the somewhat slenderly suspended cathodes, by putting less weight on the supports. The removal in one operation of the entire batch of cathodes from the bath is effected by means of the suspended hook-frame, as employed in charging. The plates are rinsed, the top edges are cut off and returned with the anode scrap, whilst the pure electrolytic copper passes to the refining and casting furnaces, where it is prepared for the market.

Collection of the Slimes.—Depending upon the working conditions of the refinery, but usually at intervals of three months, the precipitated slimes are collected and the tanks are cleaned out. The quantity of slime deposited is generally not very large, from 15 to 25 lbs. per tank being a not unusual yield. The current and the supply of solution are cut off, the plates removed, the contents of the tank allowed to settle, the liquid siphoned off to within about 6 inches of the bottom, and the residues are swilled out through a trap at the bottom of the tank. The sludge passes through a sieve that separates the lumps of anode copper which have broken off and fallen to the bottom of the tank, the slime then passes to the special refinery for treatment. The processes adopted for recovering the gold and silver from this residue are highly specialised, and belong properly to the technology of refining of the precious metals.

Modifications of Electrolytic Refining.—Great success has not yet attended the attempts which have been made to employ copper matte in the form of anodes in electrolytic refining processes, and the method is not in operation at any of the great modern works. Marchésé, Hoepfner, Siemens-Halske, Keith, and others have introduced processes, but their practical operation is attended with very great difficulty and but little commercial success. Matte is exceedingly brittle and it readily breaks up, it is a bad conductor and necessitates the use of high voltage, the solutions become very foul, and the processes require very special apparatus and equipment.

Methods for the production, by electrolytic processes, of pure copper in forms ready for service, such as wires or tubes, have been introduced successfully by a number of workers, including Elmore, Thomerson, and Cowper-Coles. Several of these methods are now in apparently successful commercial operation, and the published results of the working of the processes and of tests on the deposited materials offer considerable promise for their future industrial application for special purposes, if not for general use. The attaining of the necessary compactness, toughness, and strength of the metallic product is aided by the employment of pressure during deposition, as by burnishers, or by very rapid rotation of the depositing surfaces in the solutions. Details of these processes and products may be found from the references subsequently given.

Bringing up to Pitch and Casting the Merchant Copper.—The final stages in the smelting process from ore to market-metal are those of “fining,” toughening, and casting the cathode copper, the object of these operations being to impart to the metal the chemical composition and mechanical and physical properties which are required in order to fit it for the market, and also to prepare it into a suitable form for service. In addition to cathode copper, other forms of the metal, if of suitable composition, are also treated with this object.

For conductivity copper, however, these final operations are conducted on metal from which practically all the impurities have been removed, but which is not sufficiently tough and homogeneous or which is not in a suitable shape for immediate industrial use. The toughening operation consists almost entirely of adjusting the percentage of oxides in the metal, partly in order to overcome the influence of any traces of injurious impurity that might remain, but mainly to exercise the functions previously indicated, of imparting by its more or less direct action upon the metal, a definite toughening and strengthening effect. The mechanism of the action is not perfectly understood, but the recent work referred to in [Lecture II., p. 28], affords useful evidence as to its possible mode of action.

The actual refining operation and the furnace employed for the process are exactly similar to those used in preparing the metal to ensure the casting of sound ingots, as already described. The operations consist of a preliminary aëration, by means of which any oxidisable impurity still remaining in the metal is oxidised out, mainly through the action of copper oxide which is formed during the process in some considerable excess.

After the copper has become “dry” or over-oxidised, which condition is characterised by brittleness, depressed surface, and brick-like purple-red fracture of the metal, it is reduced by poling and timbering operations to a definite point, viz.: until a sample ingot of the metal indicates a maximum of toughness, accompanied by level surface and bright salmon-coloured silky fracture—it is then of “tough-pitch” quality.

The furnace employed for the refining has already been described. One of the main features in which it differs from the ordinary modern reverberatory smelting-furnace is that owing to the exceedingly high heat-conducting power of metallic copper, and to the absence of an insulating layer of non-conducting slag, there is little danger of much chilling action occurring on the hearth of the furnace; the temperature may, indeed, often become too high rather than too low. In consequence, it is not so usual to construct the furnace with a very massive hearth foundation as for smelting, but to build it upon a vault or upon a series of piers. With this type of foundation, the very considerable, but practically unavoidable, absorption of metal in the hearth-material is reduced to a minimum. It is usual to work a charge consisting of scrap and oxide in the furnace before the regular smelting campaign begins in order to “season” the hearth. This procedure allows the primary absorption of copper by the hearth-material, and assists its consolidation, whilst the action of the oxide promotes a surface glazing which lessens the tendency for further absorption of copper, and gives a good surface to the working bed. As has been already stated, the hearth is generally built of brickwork rather than of sand. The furnace is constructed to hold from 80 up to 200 tons of metal. The method of working differs mainly from that previously described, in that instead of pouring molten copper into the furnace, as is usual with converter-metal, the cathode plates must be charged in a different manner.

In order to deal with such a large quantity of charge in this bulky form, without occupying so much time as to make the whole operation too protracted, it is usual to employ some form of charging machine rather than to use hand labour for the operation. In some cases a small melting furnace is employed solely for the purpose of preparing the metal in a molten form for feeding into the refining furnace. The type of cathode-charger most used is very similar in operation to the Welman charger for steel furnaces, and by its means, 100 tons of material can be charged per hour.

Operation.—The refining and toughening process is conducted in the six stages of:—

(a) The charging is sometimes conducted in stages, this being indeed unavoidable when very large quantities of material are worked with, the bulk of which, when solid, would more than fill the furnace. Two-thirds or three-quarters of the material may be put in at first and just melted down slowly, after which the remainder is added.

Owing to the not infrequent presence of sulphur in the furnace coals, and to its ready affinity for copper, resulting in undesirable consequences for the commercial metal, contamination by this element is usually prevented, as much as possible, by giving to the cathodes a wash of lime previous to charging.

(b) The melting is generally conducted somewhat slowly, so as to allow some oxidation of the metal during this stage, which may occupy some twelve hours. Skimming of slag as it forms, and subsequent blowing of the copper towards the end of the melting stage are frequently resorted to.

(c) The slag which accumulates, sometimes in considerable quantity, is skimmed off as occasion requires. When converter metal of such purity as not to need electrolytic refining is treated directly in the furnace, much of this slag is converter-slag introduced from the ladle, and requires to be skimmed off at an early stage. In the usual process of melting cathode-copper, slag is produced from the last traces of iron which may have remained in the metal. In order to render it sufficiently viscid to be pulled out by the skimmer, ashes from the fire-grate are thrown upon and rabbled into the slag. This skimming may continue for some time, and a very rich coppery slag is pulled off, from which the metal values are subsequently recovered.

(d) The oxidation of the small quantities of impurity still remaining in the metal is completed by the operation of airing, as already described, and the action is continued in order to produce a small excess of oxide until the copper is “dry.” The time occupied for this airing is now not very protracted, since most of the impurities have been previously removed from the metal.

(e) The copper is then brought up to pitch by “poling” in the manner previously indicated, except that at this final stage, the testing of the metal and the adjusting of the oxygen proportion are conducted with much greater precision than was necessary for the simple production of the sound anode plates. In the present instance, the character of the metal and its value as a commercial article largely depend upon the care and accuracy with which the correct “pitch” is reached and is maintained in the bath during the entire period of casting of the metal. The poling for the “shaking out” of the gases is rarely necessary with cathode metal, and the addition of the cover of carbonaceous material for the purpose of effecting the reduction of the oxides to the desired extent, is made either at the commencement of poling or else shortly afterwards. After some time, a series of small samples is taken at intervals, by means of ladles, and the surface of the ingot is examined. The depression characteristic of dry copper gradually becomes less marked, the brick-like fracture appears finer and finer until it becomes silky, whilst the colour eventually turns to a very delicate salmon-pink. Meanwhile the mechanical properties have gradually improved, signs of brittleness disappear, and somewhat larger samples of the metal, which are now taken and tested, are characterised by a very marked toughness and strength. This is the moment at which the poling must cease. The residual copper-oxide has now reached the proportion which was necessary for the imparting of the best mechanical properties, and the metal is tough-pitch. The skill of the workman is now exercised to the highest degree, in maintaining the metal in this condition during the whole of the subsequent casting period. Oxidation must be avoided in order to prevent a reversion to dry copper, whilst any further reducing action removes some of the necessary oxide, and results in “over-poling.” The metal would then become brittle again, coarsely fibrous and possibly somewhat spongy in fracture and very pale in colour, whilst in setting it would show a ridge upon the surface. In that case it would be necessary to “air” the metal again until it became dry, and then to pole it back to the “tough-pitch” stage.

The copper, when of correct pitch, is therefore removed from the furnace and cast at once; this being readily conducted through the tapping slot, the level of which is gradually lowered. The metal then flows down the spout to the ladle, and is poured into the moulds attached to some form of mechanical casting machine; the ingots being finally dropped into a water-bosh, weighed, sampled, and stacked, and are then in a condition ready for the market.

Phosphorus is sometimes employed for giving soundness to the castings, being added to the bath in small quantities in the form of phosphor-copper containing about 10 per cent. of the non-metal. Although very little of this phosphorus is retained by the metal, being mostly eliminated as oxide, special caution is required in employing it for high-grade conductivity copper, since the effect of very small quantities has a deleterious influence upon the conducting properties.

Silicon also is used for a similar purpose, and causes a considerable increase in toughness.

(f) When intended for conductivity work, the metal is cast into the form of “wire-bars” of very varied shape and size, according to requirements; thus the 100-lb. bars are about 3 feet long by about 3 inches square section, the 500-lb. bars 7 feet long by about 4½ inches square. Furnace samples weighing about 1 lb. are drawn down gradually to about ⅛-inch wire, and are tested for conductivity, as well as for strength and toughness, occasional analysis being also undertaken, whilst samples of the wire-bars in market form are similarly examined.

a b

Fig. 76.—Microstructure of Commercial Copper containing Oxygen (Hofman).

a. Calumet and Hecla copper after
60 minutes’ poling.		b. Calumet and Hecla “dry” copper
before poling.
0·22 per cent. oxygen = 1·98 per cent. Cu2O.  0·64 per cent. oxygen = 5·76 per cent. Cu2O.

Compare with [Fig. 6, p. 28].
(By permission of American Institution of Mining Engineers.)

References.

Ulke, T., “Modern Electrolytic Copper Refining.” (With complete Bibliography).

Peters, E. D., “Modern Copper Smelting” (1905). (Chapter xviii. by M. Barnett.).

Mineral Industry. Annual Review.

Cowper Coles, S., “An Electrolytic Process for the Production of Copper Wire.” Proc. B’ham. Met. Soc., 1908–9, p. 5.

Schnabel and Louis, “Handbook of Metallurgy,” pp. 327–369.

Wraith, W., “Sampling Copper Anodes at Anaconda.” Trans. Amer. Inst. Min. Eng., March, 1910.

“The De Lamar Electro-Refinery at Chrome, New Jersey.” Eng. and Min. Journ., Jan. 13, 1906, p. 73.

Flinn, F. B., “Electrolytic Copper.” The Metal Industry, April, 1910, p. 112, vol. ii., No. 3.

Greenawalt, W. E., “The Greenawalt Electrolytic Process.” Eng. and Min. Journ., Nov. 26, 1910, p. 1062.

See also References to Johnson (p. 34), Hofman and others (p. 50), Keller (pp. 50, 215), and Peters (p. 80).