Chapter C: THE MACDOUGAL TYPE.—The most important form of modern roaster (6)

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.

(_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.

=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—

(1) To preserve the steel shell and form a permanent
receptacle for the molten materials; by reason of
its refractory character.
(2) To prevent undue losses of heat from the materials;
by reason of its low conducting power.

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 SiO_{2}, 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—

(_a_) Large quantities of iron oxide are formed, which
require siliceous flux.
(_b_) Large quantities of highly ferruginous slag are
produced which carry copper values, and which also
demand special attention in operating.
(_c_) The quantity of copper obtained is comparatively
small, thus increasing the proportionate losses
and working difficulties.

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 SO_{2}. 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 am Charged No. 1 blast matte,│.. │ .. │46·08│0·15│ 31·50
11.54 am Blow commenced. │ │ │ │ │
12.04 pm Sample No. 1 blowing, │.. │ 10 │46·02│0·17│ 31·80
12.14 pm Sample No. 2 blowing, │.. │ 20 │51·46│0·18│ 35·80
12.18 pm Punched 3 minutes. │ │ │ │ │
12.24 pm Sample No. 3 blowing, │.. │ 30 │53·27│0·20│ 37·80
12.25 pm Punched 7 minutes. │ │ │ │ │
12.34 pm Sample No. 4 blowing, │.. │ 40 │56·29│0·21│ 40·80
12.40 pm Punched 2 minutes. │ │ │ │ │
12.44 pm Sample No. 5 blowing, │.. │ 50 │59·90│0·22│ 43·70
12.45 pm Punched 2 minutes. │ │ │ │ │
12.54 pm Sample No. 6 blowing, │ 1 │ 00 │62·67│0·23│ 44·90
12.55 pm Punched 6 minutes. │ │ │ │ │
1.04 pm Sample No. 7 blowing, │ 1 │ 10 │67·89│0·25│ 49·20
1.07 pm Punched 3 minutes. │ │ │ │ │
1.12 pm Blow stopped. │ │ │ │ │
1.13 pm Skimmed. │ │ │ │ │
1.14 pm Blow resumed. │ │ │ │ │
1.14 pm Sample No. 8 blowing, │ 1 │ 20 │73·97│0·27│ 54·90
1.17 pm Punched 2 minutes. │ │ │ │ │
1.21 pm Blow stopped. │ │ │ │ │
1.22 pm Skimmed. │ │ │ │ │
1.25 pm Blow resumed. │ │ │ │ │
1.25 pm Sample No. 9 blowing, │ 1 │ 30 │77·82│0·28│ 57·30
1.34 pm Sample No. 10 blowing, │ 1 │ 40 │74·16│0·26│ 54·30
1.44 pm Sample No. 11 blowing, │ 1 │ 50 │81·72│0·15│ 57·60
1.54 pm Sample No. 12 blowing, │ 2 │ 00 │98·50│0·78│107·70
2.02 pm Punched 1 minute. │ │ │ │ │
2.04 pm Sample No. 13 blowing, │ 2 │ 10 │98.57│0·40│ 81.60
2.05 pm Punched 2 minutes. │ │ │ │ │
2.08 pm Blow stopped, test for Cu.│ │ │ │ │
2.09 pm Blow resumed. │ │ │ │ │
2.10 pm Blow finished. │ │ │ │ │
Converted copper, │ 2 │ 16 │99.08│0·38│ 83·80
Total time punching, │.. │ 28 │ .. │ .. │ ..
Total time of blow, │ 2 │ 16 │ .. │ .. │ ..
Actual time of blow, │ 2 │ 09 │ .. │ .. │ ..
───────────────────────────────────┴───┴────┴─────┴────┴──────

───────────────────────────────────┬────────┬────────────────────────
│ Time. │ Composition.
├───┬────┼──────┬─────┬─────┬─────
│ │ │Insol-│ │ │
│Hrs│Mins│ uble │ Fe. │ S. │ As.
───────────────────────────────────┼───┼────┼──────┼─────┼─────┼─────
│ │ │ % │ % │ % │ %
11.52 am Charged No. 1 blast matte,│.. │ .. │ 0·15 │24·30│24·70│0·22
11.54 am Blow commenced. │ │ │ │ │ │
12.04 pm Sample No. 1 blowing, │.. │ 10 │ 1·30 │23·70│22·95│0·07
12.14 pm Sample No. 2 blowing, │.. │ 20 │ 0·30 │20·50│23·10│0·06
12.18 pm Punched 3 minutes. │ │ │ │ │ │
12.24 pm Sample No. 3 blowing, │.. │ 30 │ 1·10 │18·70│22·15│0·06
12.25 pm Punched 7 minutes. │ │ │ │ │ │
12.34 pm Sample No. 4 blowing, │.. │ 40 │ 1·30 │16·20│21·85│0·06
12.40 pm Punched 2 minutes. │ │ │ │ │ │
12.44 pm Sample No. 5 blowing, │.. │ 50 │ 0·90 │13·70│21·95│0·06
12.45 pm Punched 2 minutes. │ │ │ │ │ │
12.54 pm Sample No. 6 blowing, │ 1 │ 00 │ 1·30 │11·40│21·35│0·06
12.55 pm Punched 6 minutes. │ │ │ │ │ │
1.04 pm Sample No. 7 blowing, │ 1 │ 10 │ 0·65 │ 7·60│21·15│0·05
1.07 pm Punched 3 minutes. │ │ │ │ │ │
1.12 pm Blow stopped. │ │ │ │ │ │
1.13 pm Skimmed. │ │ │ │ │ │
1.14 pm Blow resumed. │ │ │ │ │ │
1.14 pm Sample No. 8 blowing, │ 1 │ 20 │ 0·25 │ 3·40│20·10│0·05
1.17 pm Punched 2 minutes. │ │ │ │ │ │
1.21 pm Blow stopped. │ │ │ │ │ │
1.22 pm Skimmed. │ │ │ │ │ │
1.25 pm Blow resumed. │ │ │ │ │ │
1.25 pm Sample No. 9 blowing, │ 1 │ 30 │ 0·15 │ 0·90│19·60│0·04
1.34 pm Sample No. 10 blowing, │ 1 │ 40 │ 3·30 │ 2·60│16·60│0·04
1.44 pm Sample No. 11 blowing, │ 1 │ 50 │ 0·25 │ 0·20│15·35│0·04
1.54 pm Sample No. 12 blowing, │ 2 │ 00 │ 0·017│trace│ 0·78│0·050
2.02 pm Punched 1 minute. │ │ │ │ │ │
2.04 pm Sample No. 13 blowing, │ 2 │ 10 │ 0·052│ 0·01│ 0·78│0·033
2.05 pm Punched 2 minutes. │ │ │ │ │ │
2.08 pm Blow stopped, test for Cu.│ │ │ │ │ │
2.09 pm Blow resumed. │ │ │ │ │ │
2.10 pm Blow finished. │ │ │ │ │ │
Converted copper, │ 2 │ 16 │ 0·017│trace│ 0·01│0·033
Total time punching, │.. │ 28 │ .. │ .. │ .. │ ..
Total time of blow, │ 2 │ 16 │ .. │ .. │ .. │ ..
Actual time of blow, │ 2 │ 09 │ .. │ .. │ .. │ ..
───────────────────────────────────┴───┴────┴──────┴─────┴─────┴─────
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:—

┌─ ─┐
│Oxidation of secondary │ Dark reddish-brown
At commencement ─┤ constituents, ├─ flame.
of blow, │Burning of iron, sulphur,│ Accompanied by much
└─ and coal, ─┘ smoke.

Slagging stage, Iron-sulphide oxidation, Apple-green flame.
White metal stage, Copper oxidation in White-blue flame.
presence of slag,
Blowing to blister Sulphur oxidation, Thin red-purple flame.
copper,

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.

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

Cu_{2}S + 2Cu_{2}O ➡ 6Cu + SO_{2}.

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:—

(_a_) Grade of matte.
(_b_) Temperature of matte.
(_c_) Condition of converter lining.
(_d_) Rate of production of matte.
(_e_) Condition of affairs at the casting
and refining furnaces.

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

(_a_) The use of a material which is not perfectly suited for
constructional work, hence repairs occupy longer time.
(_b_) The risks of destroying the lining mechanically near
the tuyeres, owing to the extra punching required at
these points.
(_c_) The operation and manipulations requiring extreme care and
attention, owing to the tendency for the production of very
high temperature during the great evolution of heat in the
early stages of the blow, when large quantities of iron are
being oxidised.
(_d_) The tendency to losses, by the blowing out of the dry
siliceous ore, when first turning on the blast.

=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.

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, 99 per 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—

(_a_) Of purifying converter-copper.
(_b_) Of recovering from copper, the bullion-values which
have been collected in the metal.
(_c_) Of manufacturing pure copper, and recovering the gold
and silver as profitable bye-products.

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.

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.